UNITED STATES
SECURITIES AND EXCHANGE COMMISSION
Washington, D.C. 20549

_______________________

FORM 40-F/A
(Amendment No. 1)

[X] Registration statement pursuant to Section 12 of the Securities Exchange Act of 1934

or

[   ] Annual report pursuant to Section 13(a) or 15(d) of the Securities Exchange Act of 1934

For the fiscal year ended _______________________ Commission File Number 001-38179

_______________________

Kirkland Lake Gold Ltd.
(Exact name of Registrant as specified in its charter)

Ontario 1000 Not Applicable
(Province or other jurisdiction of (Primary Standard Industrial Classification (I.R.S. Employer
incorporation or organization) Code Number) Identification Number)

200 Bay Street, Suite 3120
Toronto, Ontario M5J 2J1
Canada
(416) 840-7884
(Address and telephone number of Registrant’s principal executive offices)

_______________________

Registered Agent Solutions, Inc.
99 Washington Avenue
Suite 1008
Albany, NY 12260
(888) 705-7274
(Name, address (including zip code) and telephone number (including
area code) of agent for service in the United States)

_______________________

Securities registered or to be registered pursuant to Section 12(b) of the Act:

Title of each class Name of each exchange on which registered
   
Common Shares, no par value New York Stock Exchange

Securities registered pursuant to Section 12(g) of the Act: None.

Securities for which there is a reporting obligation pursuant to Section 15(d) of the Act: None

For annual reports, indicate by check mark the information filed with this Form:

[   ] Annual information form [   ]Audited annual financial statements

Indicate the number of outstanding shares of each of the registrant’s classes of capital or common stock as of the close of the period covered by the annual report: N/A

Indicate by check mark whether the registrant: (1) has filed all reports required to be filed by Section 13 or 15(d) of the Exchange Act during the preceding 12 months (or for such shorter period that the registrant was required to file such reports); and (2) has been subject to such filing requirements for the past 90 days. [   ] Yes      [   ] No

Indicate by check mark by filing the information contained in this Form is also thereby furnishing the information to the Commission pursuant to Rule 12g3-2(b) under the Securities Exchange Act of 1934 (the “Exchange Act”). If “Yes” is marked, indicate the file number assigned to the Registrant in connection with such Rule. [   ] Yes      [X] No

Indicate by check mark whether the registrant is an emerging growth company as defined in Rule 12b-2 of the Exchange Act.


Emerging growth company [X]

If an emerging growth company that prepares its financial statements in accordance with U.S. GAAP, indicate by check mark if the registrant has elected not to use the extended transition period for complying with any new or revised financial accounting standards† provided pursuant to Section 13(a) of the Exchange Act. [   ]

The term “new or revised financial accounting standard” refers to any update issued by the Financial Accounting Standards Board to its Accounting Standards Codification after April 5, 2012.


EXPLANATORY NOTE

Kirkland Lake Gold Ltd. (the “Company”, the “Registrant”) is a Canadian issuer eligible to file its registration statement pursuant to Section 12 of the Securities Exchange Act of 1934, as amended (the “Exchange Act”), on Form 40-F pursuant to the multi-jurisdictional disclosure system of the Exchange Act. The Company is a “foreign private issuer” as defined in Rule 3b-4 under the Exchange Act. Equity securities of the Company are accordingly exempt from Sections 14(a), 14(b), 14(c), 14(f) and 16 of the Exchange Act pursuant to Rule 3a12-3.

The Company filed a Registration Statement on Form 40-F on August 4, 2017 (the “Original Form 40-F”). The Company is filing this Amendment No. 1 for the sole purpose of filing exhibits that were too large to be filed with the Original Form 40-F.

4


SIGNATURES

Pursuant to the requirements of the Exchange Act, the Registrant certifies that it meets all of the requirements for filing on Form 40-F and has duly caused this Registration Statement to be signed on its behalf by the undersigned, thereunto duly authorized.

KIRKLAND LAKE GOLD LTD.
   
   
By: /s/ Jennifer Wagner
  Name: Jennifer Wagner
  Title: Corporate Secretary

Date: August 4, 2017

5


EXHIBIT INDEX

The following documents are being filed with the Commission as Exhibits to this Registration Statement:

Exhibit Description
   
99.1* Annual Audited Consolidated Financial Statements for Kirkland Lake Gold Ltd. as at December 31, 2016, December 31, 2015, April 30, 2015 and April 30, 2014 and the Year Ended December 31, 2016, the Eight-Month Period Ended December 31, 2015 and Year Ended April 30, 2015*
   
99.2* Management's Discussion and Analysis for the year ended December 31, 2016*
   
99.3* Annual Information Form dated March 30, 2017*
   
99.4* Certification of Refiled Annual Financial Statements by the CEO dated August 1 2017*
   
99.5* Certification of Refiled Annual Financial Statements by the CFO dated August 1 2017*
   
99.6* Indenture dated July 19, 2012*
   
99.7* Supplemental Indenture dated November 7, 2012*
   
99.8* Arrangement Agreement dated November 16, 2015*
   
99.9* News Release dated January 11, 2016*
   
99.10* News Release dated January 18, 2016*
   
99.11* News Release dated January 26, 2016*
   
99.12* Articles of Arrangement dated January 26, 2016*
   
99.13* News Release dated February 2, 2016*
   
99.14* News Release dated February 12, 2016*
   
99.15* Material Change Report dated February 17, 2016*
   
99.16* News Release dated February 26, 2016*
   
99.17* News Release dated February 29, 2016*
   
99.18* News Release dated March 4, 2016*
   
99.19* Consolidated Financial Statements for the years ended December 31, 2015 and 2014*
   
99.20* Management’s Discussion and Analysis for the years ended December 31, 2015 and 2014*
   
99.21* Confirmation of Notice of Record and Meeting Dates dated March 16, 2016*
   
99.22* News Release dated March 21, 2016*
   
99.23 Technical Report for the Maud Creek Gold Project, Northern Territory Australia dated March 21, 2016
   
99.24 Technical Report for the Stawell Gold Mine, Victoria, Australia dated March 16, 2016
   
99.25 Report on the Mineral Resources & Minerals Reserves of the Northern Territory Operations, Northern Territory, Australia dated March 21, 2016
   
99.26** Report on the Mineral Resources & Mineral Reserves of the Fosterville Gold Mine, Victoria, Australia dated March 21, 2016**
   
99.27* Annual Information Form for the year ended December 31, 2015*
   
99.28* Certification of Annual Filings in connection with filing of Annual Information Form by CEO March 21, 2016*
   
99.29* Certification of Annual Filings in connection with filing of Annual Information Form by CFO March 21, 2016*

6



Exhibit Description
   
99.30* Material Change Report dated March 21, 2016*
   
99.31* Revised Confirmation of Notice of Record and Meeting Dates dated March 28, 2016*
   
99.32* News Release dated March 30, 2016*
   
99.33* News Release dated April 4, 2016*
   
99.34* News Release dated April 6, 2016*
   
99.35* News Release dated April 12, 2016*
   
99.36* Notice of Annual General Meeting of Shareholders dated April 7, 2016*
   
99.37* Management Information Circular dated April 7, 2016*
   
99.38* Form of Proxy dated April 22, 2016*
   
99.39* News Release dated April 26, 2016*
   
99.40* News Release dated April 29, 2016*
   
99.41* Condensed Interim Consolidated Financial Statements for the three months ended March 31, 2016 and 2015*
   
99.42* Management’s Discussion and Analysis for the three months ended March 31, 2016 and 2015*
   
99.43* Certification of Interim Filings by CEO April 29, 2016*
   
99.44* Certification of Interim Filings by CFO April 29, 2016*
   
99.45* News Release dated May 9, 2016*
   
99.46* News Release dated May16, 2016*
   
99.47** Technical Report and Preliminary Economic Assessment of the Maud Creek Gold Project, Northern Territory, Australia dated May 16, 2016**
   
99.48* News Release dated May 18, 2016*
   
99.49** Amended Technical Report and Preliminary Economic Assessment of the Maud Creek Gold Project, Northern Territory, Australia dated May 18, 2016**
   
99.50* Material Change Report dated May 18, 2016*
   
99.51* News Release dated May 26, 2016*
   
99.52* Report of voting results dated May 26, 2016*
   
99.53* News release dated June 27, 2016*
   
99.54* News release dated July 12, 2016*
   
99.55* News release dated July 29, 2016*
   
99.56* Management’s Discussion and Analysis for the three and six months ended June 30, 2016*
   
99.57* Condensed Interim Consolidated Financial Statements for the three and six months ended June 30, 2016 and 2015*
   
99.58* Certification of Interim Filings by CEO dated July 29, 2016,*
   
99.59* Certification of Interim Filings by CFO dated July 29, 2016*
   
99.60* News release dated August 3, 2016*
   
99.61* News release dated August 22, 2016*
   
99.62* News release dated September 14, 2016*

7



Exhibit Description
   
99.63* News release dated September 20, 2016*
   
99.64* News release dated September 29, 2016*
   
99.65* Form of Voting and Support Agreement dated September 29, 2016 re Kirkland Lake Gold Inc.*
   
99.66* Form of Voting and Support Agreement dated September 29, 2016 re Kirkland Lake Gold Inc.*
   
99.67* Form of Voting and Support Agreement dated September 29, 2016 re Newmarket Gold Inc.*
   
99.68* Arrangement Agreement dated September 29, 2016*
   
99.69* Material Change Report dated October 4, 2016*
   
99.70* Confirmation of Notice of Record and Meeting Dates dated October 12, 2016*
   
99.71* News release dated October 13, 2016*
   
99.72* Revised Confirmation of Notice of Record and Meeting Dates dated October 13, 2016*
   
99.73* Certificate of Officer dated October 31, 2016*
   
99.74* Notice of Special Meeting of Shareholder of Newmarket Gold Inc. dated October 28, 2016*
   
99.75* Joint Management Information Circular Concerning an Arrangement Involving Kirkland Lake Gold Inc. and Newmarket Gold Inc. dated October 28, 2016*
   
99.76* Annual Information Form of Kirkland Lake Gold Inc. dated March 10, 2016*
   
99.77* Audited Financial statements of Kirkland Lake Gold Inc. for the stub year ended December 31, 2015 and the year ended April 30, 2015*
   
99.78* Management’s Discussion and Analysis of Kirkland Lake Gold Inc. for the eight month (stub) year ended December 31, 2015*
   
99.79* Unaudited Condensed Consolidated Interim Financial Statements of Kirkland Lake Gold Inc. as at and for the three and six month period ended June 30, 2016 and July 31, 2015*
   
99.80* Management’s Discussion and Analysis of Kirkland Lake Gold Inc. for the three and six months ended June 30, 2016*
   
99.81* Management Information Circular of Kirkland Lake Gold Inc. dated May 16, 2016*
   
99.82* Management Information Circular of Kirkland Lake Gold Inc. dated December 15, 2015*
   
99.83* Management information circular of Kirkland Lake Gold Inc. dated September 23, 2015*
   
99.84* Material Change Report of Kirkland Lake Gold Inc. dated October 3, 2016*
   
99.85* Material Change Report of Kirkland Lake Gold Inc. dated January 27, 2016*
   
99.86* News Release dated August 2, 2017*
   
99.87* Joint Management Information Circular of Newmarket Gold Inc. and Crocodile Gold Corp. dated June 2, 2015*
   
99.88* News Release dated October 31, 2016*
   
99.89* Form of proxy dated October 31, 2016*
   
99.90* News Release dated November 3, 2016*
   
99.91* Management’s Discussion and Analysis for the three and nine months ended September 30, 2016*
   
99.92* Condensed Interim Consolidated Financial Statements for the three and nine months ended September 30, 2016 and 2015*
   
99.93* Certification of Interim Filings by CEO dated November 3, 2016*

8



Exhibit Description
   
99.94* Certification of Interim Filings by CFO dated November 3, 2016*
   
99.95* News Release dated November 8, 2016*
   
99.96* News Release dated November 9, 2016*
   
99.97* News Release dated November 11, 2016*
   
99.98* News Release dated November 25, 2016*
   
99.99* Report of voting results dated November 25, 2016*
   
99.100* Letter of Transmittal for Registered Holders of Common Shares of Newmarket Gold Inc. dated November 29, 2016*
   
99.101* News Release dated November 30, 2016*
   
99.102* Articles of Amendment dated November 30, 2016*
   
99.103* Notice of Change in Corporate Structure Pursuant to Section 4.9 of National Instrument 51-102 dated December 2, 2016*
   
99.104* News Release dated November 30, 2016*
   
99.105* Second Supplemental Indenture dated as of November 30, 2016*
   
99.106* Material Change Report dated December 2, 2016*
   
99.107* News Release dated December 6, 2016*
   
99.108* News Release dated December 12, 2016*
   
99.109* News Release dated December 23, 2016*
   
99.110* News Release dated January 3, 2017*
   
99.111* Report of Exempt Distribution dated January 3, 2017*
   
99.112* News Release dated January 9, 2017*
   
99.113* News Release dated January 17, 2017*
   
99.114* News Release dated January 19, 2017*
   
99.115* News Release dated January 30, 2017*
   
99.116* News Release dated February 27, 2017*
   
99.117* News Release dated March 6, 2017*
   
99.118* Confirmation of Notice of Record and Meeting Dates dated March 10, 2017*
   
99.119* News Release dated March 28, 2017*
   
99.120* News Release dated March 29, 2017*
   
99.121* Third Supplemental Indenture dated March 13, 2017*
   
99.122*** Report on the Mineral Resources & Mineral Reserves of the Northern Territory Operations, Northern Territory, Australia dated March 30, 2017***
   
99.123*** Macassa Property, Ontario, Canada Updated NI 43-101 Technical Report dated March 30, 2017***
   
99.124*** Holt-Holloway Property, Ontario, Canada Updated NI 43-101 Technical Report dated March 30, 2017***
   
99.125**** Report on the Mineral Resources & Mineral Reserves of the Stawell Gold Mine, Victoria, Australia dated March 30, 2017****
   
99.126**** Hislop Property, Ontario, Canada Updated NI 43-101 Technical Report dated March 30, 2017****

9



Exhibit Description
   
99.127**** Report on the Mineral Resources & Minerals Reserves of the Fosterville Gold Mine, Victoria, Australia dated March 30, 2017****
   
99.128**** Taylor Property, Ontario, Canada Updated NI 43-101 Technical Report dated March 30, 2017****
   
99.129* News Release dated March 30, 2017*
   
99.130* Voting Instruction Form dated April 11, 2017*
   
99.131* Notice of Annual General Meeting of Shareholders dated April 7, 2017*
   
99.132* Management Information Circular dated April 7, 2017*
   
99.133* Form of Proxy dated April 11, 2017*
   
99.134* Kirkland Lake Gold Ltd. Long Term Incentive Plan dated April 7, 2017*
   
99.135* Kirkland Lake Gold Ltd. Deferred Share Unit Plan dated April 7, 2017*
   
99.136* Code of Conduct dated April 11, 2017*
   
99.137* News Release dated April 12, 2017*
   
99.138* Form of Proxy dated April 12, 2017*
   
99.139* News Release dated April 24, 2017*
   
99.140* News Release dated May 3, 2017*
   
99.141* News Release dated May 4, 2017*
   
99.142* Condensed Consolidated Interim Financial Statements for the three months ended March 31, 2017 and 2016*
   
99.143* Management’s Discussion and Analysis for the three months ended March 31, 2017 and 2016*
   
99.144* Certification of Interim Filings by CEO dated May 4, 2017*
   
99.145* Certification of Interim Filings by CFO dated May 4, 2017*
   
99.146* Report of voting results dated May 4, 2017*
   
99.147* News Release dated May 5, 2017*
   
99.148* News Release dated May 15, 2017*
   
99.149* News Release dated May 23, 2017*
   
99.150* Annual Report 2016*
   
99.151* News Release dated June 19, 2017*
   
99.152* News Release dated June 21, 2017*
   
99.153* News Release dated June 27, 2017*
   
99.154* News Release dated June 28, 2017*
   
99.155* News Release dated July 9, 2017*
   
99.156* News Release dated July 27, 2017*
   
99.157* Condensed Consolidated Interim Financial Statements for the three and six months ended June 30, 2017 and 2016*
   
99.158* Management’s Discussion and Analysis for the three and six months ended June 30, 2017 and 2016*
   
99.159* Certification of Interim Filings by CEO dated August 1, 2017*
   
99.160* Certification of Interim Filings by CFO dated August 1, 2017*

10



Exhibit Description
   
99.161* Consent of Jason Keily*
   
99.162* Consent of Peter Fairfield*
   
99.163* Consent of SRK Consulting (Australia) Pty Ltd.*
   
99.164* Consent of David Schonfeldt*
   
99.165* Consent of Danny Kentwell*
   
99.166* Consent of Justine Tracey*
   
99.167* Consent of Mark Edwards*
   
99.168* Consent of Wayne Chapman*
   
99.169* Consent of Murray Smith*
   
99.170* Consent of Troy Fuller*
   
99.171* Consent of Ion Hann*
   
99.172* Consent of Mining Plus PTY Ltd.*
   
99.173* Consent of Simon Walsh*
   
99.174* Consent of Pierre Rocque*
   
99.175* Consent of Douglas Carter*
   
99.176* Consent of John Winterbottom*
   
99.177* Consent of Ian Holland*
   
99.178* Consent of Glenn R. Clark*
   
99.179* Consent of Glenn R. Clark & Associates*
   
99.180* Consent of Stewart Carmichael*
   
99.181* Consent of Christopher Stewart*
   
99.182* Consent of Keyvan Salehi*
   
99.183* Consent of Dean Basile*
   
99.184* Consent of Phil Bremner*
   
99.185* Consent MiningOne Pty*
   
99.186* Consent of Simon Hitchman*
   
99.187* Consent of Stuart Hutchin*
   
99.188* Consent of GMP Securities L.P.*
   
99.189* Consent of CIBC World Markets Inc.*
   
99.190* Consent of RBC Dominion Securities Inc.*
   
99.191* Consent of Maxit Capital LP*
   
99.192* Consent of PricewaterhouseCoopers LLP*
   
99.193* Consent of KPMG LLP*

* previously filed with the Original Form 40-F
** to be filed with Amendment No. 2 to this Registration Statement on Form 40-F
*** to be filed with Amendment No. 3 to this Registration Statement on Form 40-F
**** to be filed with Amendment No. 4 to this Registration Statement on Form 40-F

11



Kirkland Lake Gold Ltd.: Exhibit 99.23 - Filed by newsfilecorp.com

Technical Report
 
Mineral Resources of the
Maud Creek Gold Project,
Northern Territory, Australia
   
   
   
Prepared for:  
Newmarket Gold Inc
 
   
   
Prepared by:  
   
SRK Consulting (Australia) Pty Ltd
 
ABN 56 074 271 720
Level 1, 10 Richardson St,
West Perth Western Australia 6005
 
 
SRK Project Number: CGC001
 
 
Qualified Persons:
 
Peter Fairfield, BEng (Mining), FAusIMM (No: 106754), CP (Mining), Principal Consultant
 
Danny Kentwell, MSc Mathematics & Planning (Geostatistics), FAusIMM, Principal Consultant
 
Date of Report: 21 March 2016

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SRK Consulting Page ii

Date and Signature Page
 
 
SRK Project Number: CGC001
 
 
SRK Consulting (Australasia) Pty Ltd
 
Level 1, 10 Richardson St,
West Perth Western Australia 6005, Australia
 
 
 
Newmarket Gold Inc
 
Maud Creek Gold Project
Northern Territory, Australia
 

Project Manager: Peter Fairfield
   
   
Date of Report: 21 March 2016
   
   
Effective Date: 15 March 2016
   

Signature Qualified Persons:

Peter Fairfield, BEng (Mining), FAusIMM (No: 106754), CP (Mining), Principal Consultant

Danny Kentwell, MSc Mathematics & Planning (Geostatistics), FAusIMM, Principal Consultant

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SRK Consulting Page iii

Important Notice

This technical report has been prepared as a National Instrument 43-101 Technical Report, as prescribed in Canadian Securities Administrators’ National Instrument 43-101, Standards of Disclosure for Mineral Projects (NI 43-101) for Newmarket Gold Inc.). The data, information, estimates; conclusions and recommendations contained herein, as prepared and presented by the Authors, are consistent with:

Information available at the time of preparation;
   
Data supplied by outside sources, which has been verified by the authors as applicable; and
   
The assumptions, conditions and qualifications set forth in this technical report.

CAUTIONARY NOTE WITH RESPECT TO FORWARD LOOKING INFORMATION

This document contains forward-looking information as defined in applicable securities laws. Forward looking information includes, but is not limited to, statements with respect to the future production, costs and expenses of the project; the other economic parameters of the project, as set out in this technical report, including; the success and continuation of exploration activities, including drilling; estimates of mineral reserves and mineral resources; the future price of gold; government regulations and permitting timelines; requirements for additional capital; environmental risks; and general business and economic conditions. Often, but not always, forward-looking information can be identified by the use of words such as plans, expects, is expected, budget, scheduled, estimates, continues, forecasts, projects, predicts, intends, anticipates or believes, or variations of, or the negatives of, such words and phrases, or statements that certain actions, events or results may, could, would, should, might or will be taken, occur or be achieved. Forward-looking information involves known and unknown risks, uncertainties and other factors which may cause the actual results, performance or achievements to be materially different from any of the future results, performance or achievements expressed or implied by the forward-looking information. These risks, uncertainties and other factors include, but are not limited to, the assumptions underlying the production estimates not being realized, decrease of future gold prices, cost of labour, supplies, fuel and equipment rising, the availability of financing on attractive terms, actual results of current exploration, changes in project parameters, exchange rate fluctuations, delays and costs inherent to consulting and accommodating rights of local communities, title risks, regulatory risks and uncertainties with respect to obtaining necessary permits or delays in obtaining same, and other risks involved in the gold production, development and exploration industry, as well as those risk factors discussed in Newmarket Gold Inc.’s latest Annual Information Form and its other SEDAR filings from time to time. Forward-looking information is based on a number of assumptions which may prove to be incorrect, including, but not limited to, the availability of financing for Newmarket Gold Inc.’s production, development and exploration activities; the timelines for Newmarket Gold Inc.’s exploration and development activities on the property; the availability of certain consumables and services; assumptions made in mineral resource and mineral reserve estimates, including geological interpretation grade, recovery rates, price assumption, and operational costs; and general business and economic conditions. All forward-looking information herein is qualified by this cautionary statement. Accordingly, readers should not place undue reliance on forward-looking information. Newmarket Gold Inc. and the authors of this technical report undertake no obligation to update publicly or otherwise revise any forward-looking information whether as a result of new information or future events or otherwise, except as may be required by applicable law.

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SRK Consulting Page iv

NON-IFRS MEASURES

This technical report contains certain non-International Financial Reporting Standards measures. Such measures have non standardized meaning under International Financial Reporting Standards and may not be comparable to similar measures used by other issuers.

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SRK Consulting Page v

Table of Qualified Persons and Contributors

Section Description Nominated QP Contributors
  Executive Summary Peter Fairfield All
2 Introduction Peter Fairfield  
3 Reliance on Experts Property Description and Peter Fairfield
4 Location Peter Fairfield
5 Accessibility, Climate, Local Resources, Infrastructure and Physiography Peter Fairfield
6 History Danny Kentwell  
7 Geological Setting and Mineralization Danny Kentwell
8 Deposit Types Danny Kentwell  
9 Exploration Danny Kentwell  
10 Drilling Danny Kentwell Kirsty Sheerin
11 Sampling Preparation, Analysis and Security Danny Kentwell Kirsty Sheerin
12 Data Verification Danny Kentwell Kirsty Sheerin
13 Mineral Processing and Metallurgical Testing Peter Fairfield Simon Walsh
14 Mineral Resource Estimates Danny Kentwell  
15 Mineral Reserve Estimates Peter Fairfield  
16 Mining Methods Peter Fairfield  
17 Recovery Methods Peter Fairfield Simon Walsh
18 Project Infrastructure Peter Fairfield  
19 Market Studies and Contracts Peter Fairfield Simon Walsh
20 Environmental Studies, Permitting and Social, or Community Impact Peter Fairfield Lisa Chandler, Ken Redwood
21 Capital and Operating Costs Peter Fairfield  
22 Economic Analysis Peter Fairfield  
23 Adjacent Properties Danny Kentwell All
24 Other Relevant Data and Information Danny Kentwell All
25 Interpretation and Conclusions Danny Kentwell All
26 Recommendations Danny Kentwell All
27 References Danny Kentwell All

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List of Abbreviations

Abbreviation Meaning
2D two dimensional
3D three dimensional
AAS atomic absorption spectroscopy
ALS ALS Minerals
AMC AMC Consultants Pty Ltd
Amdel Amdel Limited Mineral Services Laboratory
ANFO ammonium nitrate-fuel oil
ASL above sea level
ATCF after tax cash flow
Au gold
Au gold equivalent
AUD Australian dollar
BAppSc Bachelor of Applied Science
BCom Bachelor of Commerce
BD bulk density
BEng Bachelor of Engineering
BSc Bachelor of Science
Cambrian Cambrian Mining Limited
CIM Canadian Institute of Mining, Metallurgy and Petroleum
CIP carbon-in -pulp
(CP) Chartered Professional of The Australasian Institute of Mining and Metallurgy
CRF cemented rock fill
dmt dry metric tonne
DTM digital terrain model
EM electromagnetic
EPA Environmental Protection Agency
EVC’s Ecological Vegetation Classes
FAR fresh air rise
FAusIMM Fellow of The Australasian Institute of Mining and Metallurgy
GDip Graduate Diploma
GEF Gold and Exploration Finance Company of Australia
GMA/WMC Gold Mines of Australia/Western Mining Corporation
GPS global positioning system
GST goods and services tax
g/t grams per tonne
HBr hydrobromic acid
HCl hydrochloric acid
HR hydraulic radius
ICP - AES inductively couple plasma atomic emission spectroscopy
ID2 inverse distance squared
ID3 inverse distance cubed
IP induced polarisation
IRR internal rate of return
JORC Joint Ore Reserves Committee

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Abbreviation Meaning
kg kilogram
kL kilolitre
km kilometer
Koz kilo ounces
kt kilotonne
ktpa kilotonnes per annum
Ktpm kilotonnes per month
kV kilovolt
kVA kilovolt ampere
kW kilowatt
kWh kilowatt hour
L litres
LHD load-haul-dump
LOM/ LoM life of mine
L/s litres per second
M million
Ma million years
Newmarket Gold Newmarket Gold Inc
MAusIMM(CP) Member of The Australasian Institute of Mining and Metallurgy
mg/kg milligrams per kilogram
mg/L milligrams per litre
mH meters high
ML million litres
mm millimeters
MMI mobile metal ion
Moz million ounces
mRL meters reduced level
MRSD Act Mineral Resources (Sustainable Development) Act 1990
Mtpa million tonnes per annum
m3 cubic meters
m3/s cubic meter per second
m3/s/KW cubic meter per second per kilowatt
MVA megawatt ampere
mW meters wide
MW megawatt
NI 43-101 National Instrument 43-101
NPV net present value
OH & S Occupational Health and Safety
Onsite Onsite Laboratory Services
ozs ounces
PEA Preliminary Economic Assessment
PhD Doctor of Philosophy
Planet Planet Resources Group NL
QA/QC quality assurance/quality control
QP Qualified Person
RAR return air raise

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Abbreviation Meaning
RC reverse circulation
ROM run-of-mine
SD standard deviation
SRK SRK Consulting (Australasia) Pty Ltd
t tonnes
tpa tonnes per annum
t/mth tonnes per month
TSF tailings storage facility
TSX Toronto Stock Exchange
UCS unconfined compressive strength
USD US dollars
V volt

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SRK Consulting Page ix

Executive Summary

Introduction

SRK Consulting (Australasia) Pty Ltd (SRK) was engaged by Newmarket Gold Inc. (Newmarket Gold) to undertake a study on the Maud Creek Gold Project (Maud Creek or the Project) and prepare a Technical Report to support the release of the updated Mineral Resource estimate.

In early July 2015, Newmarket Gold merged with Crocodile Gold Corp. (Crocodile Gold) to form a new Canadian, Toronto Stock Exchange listed gold mining company named Newmarket Gold that has 100% ownership of the Maud Creek Project.

This Technical Report documents the review and assessment of the project’s geology, exploration, mineral resource, geotechnical and metallurgical aspects prepared by SRK. It was prepared following the guidelines of the Canadian Securities Administrators’ National Instrument (NI) 43-101 and Form 43-101 F1.

The mineral resource statement reported herein was prepared in conformity with generally accepted Canadian Institute of Mining, Metallurgy, and Petroleum’s (CIM) Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines.

Scope

The scope of this study was to review the geological model, update the Mineral Resource estimate, review the geotechnical and metallurgical data and aspects of the deposit to support future studies

Based on the Mineral Resource review and findings of the geotechnical and metallurgical review work, SRK recommends that the Preliminary Economic Assessment (PEA) of the Project continue.

The basis of the PEA is processing options that enable the Oxide mineralisation to be considered in a potential mine plan and a processing route and sale of a gold rich concentrate.

The previous Technical Report, Bremner, P and Edwards, M 2012. Report on the Mineral Resource and Mineral Reserve of the Maud Creek Gold Project, excluded Oxide mineralisation, assumed processing including a Bacterial Oxidation plant (BiOX) and sale of gold Dore.

The PEA is considering a stand-alone processing plant and associated infrastructure at Maud Creek and processing of mineralization at the Union Reefs processing plant. Both options would be designed to produce a gold-rich concentrate that would be transported to the Port of Darwin for shipping to overseas markets.

Property Description and Location

The Maud Creek Gold project (Maud Creek or the Project) is located within the Pine Creek region of the Northern Territory of Australia, 20 kilometers north-east of Katherine. Previous mining activities at Maud Creek have been limited to open pit mining during 2000 when the owner was AngloGold.

The project comprises a total of 23 mineral titles (all granted), and the deposit is located wholly within tenement ML30260 which is held 100% by Newmarket Gold. Maud Creek is located at latitude 14°26’41” south and longitude 132°27’10” east.

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Accessibility, Climate, Local Resources, Infrastructure and Physiography

Access is gained to the Project from Darwin by travelling south for some 314 road kilometers along the sealed Stuart Highway to the town of Katherine.

Darwin has a population in excess of 129,000 and is the capital city of the Northern Territory. It is the administrative centre of the Northern Territory government and a major transportation hub, with an international airport and deep-water port and the Adelaide to Darwin transcontinental railway terminating at the East Arm port.

Katherine is a regional centre with a population of approximately 9,800 and enjoys excellent infrastructure, services and communications. This is the closest centre of population to the Maud Creek project. The regional mining communities of Pine Creek (with a population of 450) and Adelaide River (population of 200) support the Burnside, Maud Creek and Moline gold projects.

The major land use is grazing on native pastures and traditional Indigenous uses with some horticulture, grazing on modified pastures and nature conservation. The region has undergone some clearing (approximately 167,000 ha) for these developments the vegetation of the Maud Creek area consists largely of woodlands and open woodlands (predominant species –Eucalypts) that have been degraded by the impacts of cattle, buffalo and wild donkeys. No rare, threatened or endangered species have been identified in the area.

In the Maud Creek area, the terrain is flat lying to undulating. Ephemeral streams transect the project area and drain into the westward flowing Katherine River that flows all year.

The Top End of the Northern Territory has a tropical monsoon climate characterized by two distinct seasonal patterns: the ‘wet’ monsoon and the ‘dry’ seasons. The wet season generally occurs from November through to April and the dry season between May and October. Almost all rainfall occurs during the wet season, mostly between December and March, and the total rainfall decreases with distance from the coast.

History

Gold was initially discovered in the Maud Creek area in 1890 and a small plant was set up but ultimately abandoned in 1891. This is now called the Chlorite Hills and O’Shea’s area.

The area was re-looked at from 1932-34 when 400 tonnes of ore -produced 540 ounces of gold. Mining was from about 20 shallow shafts and small holes that were 6 - 12 meters deep with horizontal workings from 15 30 meters in length in the Chlorite Hills and O’Shea’s area.

Interest in the Maud Creek area was rekindled in the 1960s during an assessment of the mineral potential of the Top End of the Northern Territory. This study was prompted by the discovery of-significant uranium mineralization in the nearby South Alligator River valley in the mid 1950s.

The Maud Creek project was owned by a number of companies until the acquisition of the Project by GBS Gold in December 2006. Substantial drilling, in the order of 66,000-90,000 meters of RC and diamond drilling, is reported on the Project area during the period 1966 – 2006, oriented toward gold exploration.

AngloGold acquired rights to mine the oxide zone of the Main Zone deposit at Maud Creek and treat the ore at the Union Reefs plant. Mining operations were conducted during 2000. A total of 173,581 tonnes at 3.32 g/t Au produced 18,527 ounces. Ore was trucked from Maud Creek to the Union Reefs mill.

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An agreement to acquire a number of properties, including the Maud Creek property, was entered into on June 19 2009 from GBS Gold International Inc. (GBS Gold) (in liquidation). GBS Gold operated the Tom’s Gully and Brock’s Creek underground gold mines, mined several open pit gold deposits and operated two gold processing plants, one at Tom’s Gully, the other at Union Reefs, near Pine Creek, Northern Territory, until September 2008, when administrators were appointed.

On November 6 2009, the mining tenements including the Maud Creek Property were registered in the name of Crocodile Gold Australia Pty Ltd, a subsidiary of Crocodile Gold, which became Newmarket Gold in July 2015.

Geological Setting and Mineralization

The Maud Creek Gold deposit is located in the south-eastern part of the Pine Creek Geosyncline, within the Gold Creek Fault Zone, which forms the contact between mafic tuffs of the Dorothy Volcanics to the east and sedimentary rocks of the Tollis Formation to the west.

The Tollis Formation is the youngest member of the Finnis River Group, has limited aerial extent and consists of a succession of interbedded mudstone, slate, metagreywacke and minor felsic volcaniclastic shales. The Dorothy Volcanic Member consists of volcanic tuff with minor interbedded zones of sediments

The north-south trending Gold Creek Fault Zone and primary Maud Creek mineralised zones dip steeply to the east. The deposit is roughly bound to the east by the Maud Creek Dolerite, which also exhibits mineralization at the tuff/dolerite contact and to the north by a small andesite body located at the contact between the sandstone and tuff (Maud Creek Contact Fault). To the south of the deposit a major east-west structure with sinistral strike-slip movement has been interpreted. Eight faults have been identified in the Maud Creek deposit area and generally exhibit reverse movement, with limited offsets in the range of meters.

The Maud Creek Contact Fault is filled with quartz stockwork veins; three vein lodes have been modelled for the Maud Creek deposit; the primary contact vein, upper contact vein and lower contact vein. The primary vein is strongly associated with the sandstone/tuff contact but does not strictly follow the boundary. Therefore, it has been modelled as an ‘overprinting’ volume onto the sediment, tuff, dolerite and andesite lithology wireframes. Mineralization in the east at the tuff/dolerite contact generally form steeply dipping discrete lenses with limited continuity.

Outside of and adjacent to the Maud Creek Contact Fault mineralization are many intercepts carrying similar grades to those within the vein itself; these extend up to 25 meters into the hangingwall and to a lesser extent into the footwall. In addition, a greater than 0.1 g/t Au halo can be observed up to 50 meters into the hanging wall and occasionally in the footwall.

Deposit Types

A variety of genetic models have been postulated for the formation of gold deposits in the Pine Creek Geosyncline. Gold and base metal mineralization is commonly associated with granite intrusions and are often been classified as high temperature contact aureole deposits. A secondary host rock control has also been suggested due to the association of gold mineralization with carbonaceous metasedimentary rocks. More recently, authors have argued that gold mineralization is structurally controlled; occurring in brittle- ductile structures at the greenschist-amphibole facies boundary and hence has an epigenetic origin.

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Accepting that gold deposits of the Northern Territory have a structurally controlled mesothermal setting, then on the basis of host rock and mineral association they can be divided into seven types:-

Gold quartz veins, lodes, sheeted veins, stockworks, saddle reefs (Pine Creek Orogen);
   
Gold-ironstone bodies (Tennant Inlier);
   
Gold in iron rich sediments (Pine Creek Orogen, Tanami);
   
Polymetallic deposits (Iron Blow, Mt Bonnie);
   
Gold-PGE deposits (South Alligator River area);
   
Uranium-gold deposits (Pine Creek Orogen, Murphy Inlier); and
   
Placer deposits.

Of these types, Maud Creek aligns with the gold-quartz veins, lodes, sheeted veins, stockwork deposit type. Five main types of mineralization have previously been recognized within the Pine Creek Orogen. These include:

Sheeted and stockwork quartz vein systems located along major anticlinal hinges;
   

Sediment-hosted stratiform gold mineralization and quartz-sulphide- vein-hosted stratabound gold mineralization in cherty ironstone and carbonaceous mudstone;

   

Stratiform, massive to banded, sulphide-silicate-carbonate mineralization ;

   

Sediment-hosted stratiform and stratabound gold mineralization in cherty, dolomitic and sulphidic shales; and

   
Sheeted or stockwork quartz-feldspar-sulphide veins.

Of these mineralization types, Maud Creek is consistent with stockwork quartz-feldspar-sulphide veining hosted at the contact of either sandstone/tuff or tuff/dolerite units.

Mineral Resource Estimates

The Mineral Resources are stated here for the Maud Creek deposit with an effective date of 15 March 2016

The Maud Creek deposit consists of open pit and underground resources presented in Table ES-1 and ES-2. All relevant diamond drillhole samples, available as of April 2015 for the Maud Creek deposit were used to inform the estimate. The estimation methodology utilised was Ordinary Kriging (OK) to estimate gold and arsenic using hard domain boundaries.

Table ES-1:      Maud Creek Gold Project Open Pit Mineral Resource Summary

Mineral Resource
Category
Inventory
(Kt)
Gold Grade
(g/t)
Contained Metal
(KOz Au)
Measured 1 067 5.59 192
Indicated 1 100 2.14 76
Measured and Indicated 2 167 3.84 268
Inferred 531 1.41 24

It should be pointed out the mineral resource estimate is categorized as Measured, Indicated and Inferred as defined by the CIM guidelines for resource reporting. Mineral resources do not demonstrate economic viability, and there is no certainty that these mineral resources will be converted into mineable reserves once economic considerations are applied. The Measured, Indicated and Inferred mineral resource estimate has been prepared in compliance with the standards of NI 43 – 101 by Danny Kentwell, FAusIMM.

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Notes to Table ES-1:

1.

CIM definitions followed for classification of Measured, Indicated, and Inferred Mineral Resources.

2.

Mineral Resources estimated as of 15 March 2016.

3.

Mineral Resources stated according to CIM guidelines and include Mineral Reserves.

4.

Totals may appear different from the sum of their components due to rounding.

5.

Reported at a 1.5 g/t cut- off grade.

6.

The open pit mineral resource is exclusive of the underground mineral resource.

7.

The Mineral Resource estimation was performed by Danny Kentwell FAusIMM fulltime employee of SRK Consulting, who is a Qualified Person under NI 43- 101.

Table ES-2:      Maud Creek Gold Project Underground Mineral Resource Summary

Mineral Resource
Category
Inventory
(Kt)
Gold Grade
(g/t)
Contained Metal
(KOz Au)
Measured - - -
Indicated 4 330 3.28 456
Measured and Indicated 4 330 3.28 456
Inferred 1 450 2.65 124

It should be pointed out the mineral resource estimate is categorized as Indicated and Inferred as defined by the CIM guidelines for resource reporting. Mineral resources do not demonstrate economic viability, and there is no certainty that these mineral resources will be converted into mineable reserves once economic considerations are applied. The Measured, Indicated and Inferred mineral resource estimate has been prepared in compliance with the standards of NI 43 – 101 by Danny Kentwell, FAusIMM.

Notes to Table ES-2:

8.

CIM definitions followed for classification of Measured, Indicated, and Inferred Mineral Resources.

9.

Mineral Resources estimated as of 15 March 2016 .

10.

Mineral Resources stated according to CIM guidelines and include Mineral Reserves.

11.

Totals may appear different from the sum of their components due to rounding.

12.

Reported at a 1.5 g/t cut- off grade.

13.

The underground mineral resource is exclusive of the open pit mineral resource.

14.

The Mineral Resource estimation was performed by Danny Kentwell FAusIMM fulltime employee of SRK Consulting, who is a Qualified Person under NI 43- 101.

In SRK’s opinion, based on the depth and distribution of the mineralization open pit and underground mining could be viable options for extraction.

In assessing the criteria for reasonable prospects of economic extraction both open pit and underground scenarios were considered. With respect the scattered lower grade mineralization contained within the near surface a simple pit optimisation using the optimistic parameters at twice the current gold spot price did not generate a pit of practical size on the eastern domains. All material in the eastern domains is not considered to have reasonable prospects of economic extraction and does not appear in the Mineral Resource.

With respect to the underground potential the grade is reasonably consistent down to approximately 650 mRL below which it drops significantly. All material below 650 mRL is not considered to have reasonable prospects of economic extraction and does not appear in the Resource.

Mining

Based on the geological review the deposit has the potential to be exploited by conventional open pit and underground mining methods.

The final underground mining method and extraction sequence would be determined by the availability of pastefill. Pastefill will be available if a processing plant was to be constructed at Maud Creek. If mineralization is trucked to Union Reefs, a cemented aggregate fill method would be utilised.

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The Union Reefs option provides a processing stream for the oxide material within the Mineral Resource that is not available under the stand-alone plant option and will influence the mine design.

Geotechnical

In conjunction with the Mineral Resource review, a comprehensive review of the available data and analysis to determine open pit and underground geotechnical design guidelines.

An assessment of overall slope angles and underground mining parameters has been undertaken using geological and geotechnical drilling data supplied by Newmarket Gold. The analysis provides good early-stage design guidelines of the geotechnical properties of the rock mass. The typical geotechnical conditions on site can be summarised as follows:

The Hangingwall Tuffs are typically massive but may be locally bedded. Hangingwall tuffs are also affected by the numerous shears present in the Hangingwall, resulting in reduced strength, increased fracture frequency and graphitic and/or chloritic alteration of the rock mass.

 

 

The Footwall Sediments consist of low to medium strength thinly bedded or laminated mudstone and siltstone, and medium to thickly bedded sandstone. Zones of intense shearing with chlorite and graphite alteration occurring in the 5 to 10 meters below the mineralized zone where the sediments are commonly black, highly graphitic and/or chloritic, very weak and fissile.

 

 

The competency of the mineralized zone can be expected to be variable with competent, partially silicified mineralized zones separated by zones of intensely sheared rock.

 

 

The distribution of the various fault configurations is not understood at this stage and this should be one of the main focus for subsequent field investigations.

The absence of suitable data has led to low-confidence in the geotechnical conditions. Additional data is required to improve confidence and refine decisions on mining methods and the mine design. The mining method studies are linked to the decision on the location of the processing plant and the availability of pastefill.

Metallurgy and Recovery Methods

An extensive program of metallurgical testing was carried out from 1994 through to 2006 at reputable and suitably experienced laboratories. Testing was undertaken at both batch and pilot scale and including variability testing. Part of the focus of testing was on downstream oxidation processes on the refractory and preg-robbing Maud Creek mineralization, such as bio-oxidation (Biox) and the GEOCOAT® process. Direct cyanidation leaching of mineralization and concentrates was tested on the fresh (sulphide) mineralization with poor results and was eliminated as a potential processing route. A number of engineering studies were undertaken in conjunction with this test work.

Metallurgical testing has shown the mineralization to be moderately hard and abrasive, to have variable levels of gravity gold recovery, refractory and preg-robbing in nature but responsive to simple flotation techniques – demonstrating high gold recoveries in excess of 95%. Total recovery is consistently high irrespective of gravity recovery. The flotation concentrate has sufficient grade to be classified as a gold concentrate for the purposes of importation into China (> 40 g/t). It is noted that part of the gold is associated to arsenopyrite and as a result, arsenic grades in the concentrate are elevated at approximately 3.6% .

The Maud Creek mineralization has been subject to extensive metallurgical testing.

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Downstream processing options for refractory gold mineralization and concentrates were considered in earlier studies. The direct smelting of flotation concentrates option, involving concentrate being dewatered, bagged, stored in shipping containers and transported by road, rail then ship to China.

Project Infrastructure

At present there is no mining infrastructure onsite at Maud Creek. The site is serviced by local infrastructure including the Stuart Highway and services from the regional centre of Katherine. Katherine provides access to grid power, medical facilities, grid power and mobile telecommunications services.

Interpretation and Conclusions

The modelling of vein and grade volumes for the 2015 estimate takes a very different approach to the 2012 model. The 2015 model incorporates a detailed structural, vein and lithological model in the construction of the various estimation domains. This was a deliberate decision to address potential deficiencies in the very linear grade only approach previously used. Concerns had been expressed in some previous reports that insufficient attention had been paid to the geology and that the previous models may have diluted a high grade, geologically controlled core to the main zone thereby creating a model that underestimated grade and overestimated tonnages at economic cutoffs.

The differences between the 2012 and 2015 modelling approach are detailed below:

The 2015 model uses pure geology to define the main and minor vein domains. The 2012 used grade only. Consequently the 2015 vein model contains considerably lower tonnage and slightly elevated grade in comparison.

 

 

The 2015 model uses grade halos to capture both high and low grade outside the geological veins. This captures low grade material that was not modelled in 2012 which may be of value in an open pit scenario.

 

 

The 2015 model uses orientation controls on the grade halos derived by the combined fault / lithology contact model resulting in multiple orientations and fattening around fault and contact intersections.

The 2015 model is considered by SRK to be more robust in terms of its geological basis and this has led to a slightly higher grades but a reduction in contained gold. Only further drilling can define true connectivity of the mineralization in widely spaced areas.

Discussion of the risks and opportunities in the Mineral Resource model is presented in Table ES-3.

Table ES-3:      Mineral Resource Model Risks and Opportunities

Project Element Economic Risk Level Comment Opportunity
Database – Exploration data Low Historical and recent data have been re-collated and re-validated for this Mineral Resource estimate.
Assaying Low QAQC for recent and older assaying shows no material issues. Arsenic assaying has incomplete coverage. Additional assaying for Arsenic may be beneficial depending on the processing method.
Surveying Low Both collar surveys and downhole surveys completed to a high level of accuracy for recent drilling. Representative collars resurveyed for older drilling with no significant discrepancies.

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Project Element Economic Risk Level Comment Opportunity
Geology Low Detailed logging and interpretation together with evidence from both regional structural features and detailed in pit mapping informs the geological understanding Additional drilling may be able to add detail to the interaction of structures controlling mineralization at depth.
Geological modelling Low A detailed structural and lithological model has been built and incorporated into the estimation domain construction. Additional drilling may be able to add detail to the interaction of structures controlling mineralization at depth.
Resource Estimation Low Ordinary kriging cross checked and validated with theoretical grade tonnage curves and alternative search parameters has been used. The project may benefit from simulation studies or non-linear estimates if detailed studies at selective mining unit block sizes are required in the future.

Recommendations

Based on the Mineral Resource review and findings of the geotechnical and metallurgical review work, SRK recommends that the Preliminary Economic Assessment (PEA) of the Project continue.

Infill drilling in the parts of the resource currently classified as Indicated and Inferred would enable an upgrade of the Mineral Resource Classification. An approximate meterage and cost to complete this from surface down to 850 mRL is provided in Table ES-4 assuming the drilling takes place from surface. A number of sections of the geological model remain open down dip with good grades seen in the last hole down dip. Extension drilling is recommended to test these areas. Metres and costs to complete these are shown in Table ES-4 assuming drilling from surface. Costs are based on RC collars and 50m diamond drill tails.

Table ES-4:      Recommended drilling

Target Current
exploration
status
Potential End
of 2016 Status
Description Drilling
(Meters)
Total Cost
(AUD)
Infill Drilling Indicated and/or Inferred Measured and/or Indicated Increase confidence in estimated Mineral Resource 9,200 770,000
Extension drilling Down dip or along strike from current Mineral Resource Indicated and/or Inferred Close off or extend Mineral Resource volumes 2,200 200,000

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Table of Contents

  Important Notice iii
  Table of Qualified Persons and Contributors v
  List of Abbreviations vi
  Executive Summary ix
     
2 Introduction 1
     
  2.1 Scope of Work 1
  2.2 Work Programme 1
  2.3 Basis of Technical Report 1
  2.4 Qualifications of SRK and SRK Team 2
  2.5 Acknowledgement 2
  2.6 Declaration 2
       
3 Reliance on Other Experts 4
     
4 Property Description and Location 5
     
  4.1 Property Location 5
  4.2 Land Tenure 6
  4.3 Underlying Agreements 8
    4.3.1 Royalties 8
    4.3.2 Farm-out Agreement 9
    4.3.3 Farm-in Agreement 10
  4.4 Environmental Liability 11
  4.5 Legislation and Permitting 12
       
5 Accessibility, Climate, Local Resources, Infrastructure and Physiography 15
     
  5.1 Accessibility 15
  5.2 Land Use 15
  5.3 Topography 16
  5.4 Climate 16
  5.5 Infrastructure and Local Resources 17
       
6 History   18
       
  6.1 Introduction 18
  6.2 Ownership and Exploration Work 18
       
7 Geological Setting and Mineralization 23
     
  7.1 Regional Geology 23
  7.2 Property Geology 26
    7.2.1 Property Mineralization 27
  7.3 Deposit Mineralization 28
       
8 Deposit Types 29
     
  8.1 Deposit Models 29

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  8.2 Structural Models 31
       
9 Exploration 34
     
  9.1 VTEM Airborne Survey 34
  9.2 Stream sediment survey 38
  9.3 Soil sampling surveys 40
  9.4 Rock Chip Sampling 54
       
10 Drilling 56
     
  10.1 2011 Drilling program 57
    10.1.1 Surveying 59
    10.1.2 Core Recovery 59
  10.2 Sampling prior to 2011 59
    10.2.1 Surveying 59
    10.2.2 Core recovery 60
         
11 Sample Preparation, Analysis, and Security 62
     
  11.1 Sampling Techniques 62
    11.1.1 Reverse Circulation Sampling for the 2011 drilling program: 62
    11.1.2 Diamond Sampling for the 2011 drilling program: 63
    11.1.3 Sampling prior to 2011 64
  11.2 Data Sampling and Distribution 64
  11.3 Testing Laboratories 64
    11.3.1 2011 Drilling program 64
    11.3.2 Sampling prior to 2011 64
  11.4 Sample Preparation 65
    11.4.1 2011 Drilling program 65
    11.4.2 Sampling prior to 2011 67
  11.5 Sample Analysis 68
    11.5.1 2011 Drilling program 68
    11.5.2 Sampling prior to 2011 68
  11.6 Laboratory Reviews 69
  11.7 Assay Quality Assurance and Quality Control 69
    11.7.1 Standard Reference Material 69
    11.7.2 Blank Material 72
    11.7.3 Duplicate Assay Statistics 73
    11.7.4 Internal laboratory Repeats 76
    11.7.5 Inter- laboratory Repeats 78
    11.7.6 Sampling prior to 2011 80
  11.8 Sample Transport and Security 82
    11.8.1 2011 Newmarket Gold drilling program 82
    11.8.2 Sampling prior to 2011 82
  11.9 Conclusions 82

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    11.9.1 2011 Newmarket Gold drilling program 82
    11.9.2 Sampling prior to 2011 83
         
12 Data Verification 84
     
  12.1 Site Visit 86
  12.2 2011 Newmarket Gold soil sampling program 88
  12.3 Sampling prior to 2011 88
       
13 Mineral Processing and Metallurgical Testing 91
     
  13.1 Metallurgical Testing 91
    13.1.1 Comminution 91
    13.1.2 Gravity Gold Recovery 94
    13.1.3 Flotation 96
    13.1.4 Tailings and concentrate dewatering 103
  13.2 Future Test work 103
    13.2.1 Delineation of Mineralization Oxidation Extent 104
  13.3 Geometallurgy 105
       
14 Mineral Resource Estimate 112
     
  14.1 Introduction 112
  14.2 Lithology and Structural Model 112
  14.3 Vein Model 115
  14.4 Grade Halo Models 117
    14.4.1 Assumptions on non-continuity of grade adjacent to vein 119
  14.5 Domaining 121
  14.6 Compositing 121
  14.7 Metallurgical Samples 122
  14.8 Block Model Definition 124
  14.9 Grade Interpolations 124
  14.10 Declustering 125
  14.11 Outliers 126
  14.12 Drillhole Types 129
  14.13 Summary Statistics 132
  14.14 Variography 132
  14.15 Estimation 135
    14.15.1 Gold and Arsenic 135
    14.15.2 Other elements 136
  14.16 Density 137
  14.17 Validation 138
  14.18 Classification 143
  14.19 Mineral Resource Tonnage and Grade 144
       
15 Mineral Reserve Estimate 146
     
16 Mining Methods 147

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  16.1 Geotechnical 147
    16.1.1 Level of Confidence 147
    16.1.2 Project Risks and Opportunities 148
    16.1.3 Review of Geotechnical data 148
    16.1.4 Rock mass characterisation 148
    16.1.5 Underground Design Considerations 149
    16.1.6 Open pit Slope Stability Analysis 151
         
17 Recovery Methods 153
     
18 Project Infrastructure 154
     
  18.1 Site Access 154
  18.2 Power Supply 154
  18.3 Water 155
       
19 Market Studies and Contracts 156
     
20 Environmental Studies, Permitting, and Social or Community Impact 157
     
  20.1 Environment and Social Aspects and Impacts 157
    20.1.1 Social and Economic Context 157
    20.1.2 Surface Water 162
    20.1.3 Groundwater 164
    20.1.4 Native Flora and Vegetation 165
    20.1.5 Native Fauna and Habitats 168
    20.1.6 Air Quality, Noise & Vibration 170
    20.1.7 Conservation Areas 170
    20.1.8 Heritage Values 171
    20.1.9 Mine Closure and Revegetation 174
    20.1.10 Potential Impacts 174
  20.2 Regulatory Approvals 177
    20.2.1 Mineral Titles Act 177
    20.2.2 Mining Management Act 177
    20.2.3 Waste Management and Pollution Control Act 2009 178
    20.2.4 Water Act 1992 178
    20.2.5 Aboriginal Land Rights (Northern Territory) Act 1976 (Cwlth) 178
    20.2.6 Native Title Act 1993 (Cwlth) 179
    20.2.7 Heritage Act 2011 179
    20.2.8 Northern Territory Aboriginal Sacred Sites Act 1989 179
  20.3 Waste Rock and Ore Geochemistry 181
    20.3.1 Available Data 181
    20.3.2 Sample Representivity 184
    20.3.3 Geochemical Characteristics 184
  20.4 Mineralised Waste Management 187
       
21 Capital and Operating Costs 188

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22 Economic Analysis 189
     
23 Adjacent Properties 190
     
24 Other Relevant Data and Information 193
     
25 Interpretation and Conclusions 194
     
26 Recommendations 196
     
27 References 198

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List of Tables

Table 2-1: Site Visits 2
Table 4-1: Summary of Mineral Titles Maud Creek Deposit 6
Table 9-1: VTEM conductor prioritization – Maud Creek survey area 36
Table 9-2: Comparative statistics between 1997 and 2012 soil survey results 39
Table 9-3: Comparative statistics for Maud Creek and Bons Rush Area – soil sample results 51
Table 10-1: Drill statistics for the Maud Creek deposit 56
Table 10-2: Historical drilling be previous tenement holders 56
Table 10-3: Parameters used for infill diamond drilling at Maud Creek 57
Table 10-4: Maud Creek 2011 drilling data 58
Table 10-5: Core recovery of drillholes prior to 2011 61
Table 11-1: Summary of QAQC reports written for Maud Creek 64
Table 11-2: Standard ST202/5355 Compliance table for Maud Creek Newmarket Gold 2011 drilling program. 70
Table 11-3: Maud Creek Certified Laboratory Standards 70
Table 11-4: Duplicate analysis table for the 2011 drilling program 74
Table 11-5: Duplicate correlation table for the 2011 drilling program: 74
Table 11-6: Duplication R Table for the 2011 drilling program 74
Table 11-7: Repeat analysis table for the 2011 drilling program: 77
Table 11-8: Repeat correlation table for the 2011 drilling program: 77
Table 11-9: Repeat R Table for the 2011 drilling program 77
Table 11-10: Inter-laboratory repeat analysis table NTEL: ALS for the 2011 drilling program: 78
Table 11-11: Inter-laboratory repeat correlation table NTEL: ALS for the 2011 drilling program: 79
Table 11-12: NTEL: ALS Inter-laboratory repeat R Table for the 2011 drilling program 79
Table 12-1: Summary of QAQC Reports completed for Maud Creek 85
Table 12-2: Holes and intervals checked 87
Table 13-1: Crushing Test results from Report A5161 92
Table 13-2: Crushing Test Results from Report A6443 (Oxides Excluded) 92
Table 13-3: Grinding Test Results from Reports A5161 and A6076 93
Table 13-4: Summary of Comminution Results from J MacIntyre Metallurgical Evaluation (Sept ‘98) 93
Table 13-5: Comminution Results from Report A6443 94
Table 13-6: Flotation Results from Ammtec Report A5161 99
Table 13-7: Flotation Results from Ammtec Report A6076 99
Table 13-8: Classification of Zones in Maud Creek Deposit assumptions 105
Table 13-9: Historical Metallurgical Reports Used to Construct Geometallurgical Database 110
Table 14-1: Maud Creek Lithology Model Groupings 113
Table 14-2: Influence of the composite length on gold mean grade 122
Table 14-3: Comparison of arsenic assay from drilling only and including metallurgy samples 123
Table 14-4: Block model properties 124
Table 14-5: Declustering results in the vein domains 125

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Table 14-6: Declustering results in the western domains 125
Table 14-7: Declustering results in the eastern domains 126
Table 14-8: Outlier treatment in veins domains 128
Table 14-9: Outlier treatment in western domains 129
Table 14-10: Outlier treatments in eastern domains 129
Table 14-11: Comparison of mean gold grade (g/t) within each domain 132
Table 14-12: Variogram model characteristics for Au 133
Table 14-13: Variogram model characteristics for Arsenic 134
Table 14-14: Neighbourhood parameters – Au estimated by OK 135
Table 14-15: Neighbourhood parameters – Au estimated by OK – 0P1WEST soft boundary 136
Table 14-16: Neighbourhood parameters – AS estimated by OK 136
Table 14-17: Neighbourhood parameters 137
Table 14-18: Density statistics by oxide state all data 138
Table 14-19: Density statistics by oxide state in all estimation domains 138
Table 14-20: Density statistics by estimation domain within all estimation domains in Fresh rock 138
Table 14-21: Density statistics by lithology in Fresh rock all data 138
Table 14-22: Density statistics by lithology in estimation domains in in Fresh rock 138
Table 14-23: Comparison between declustered top cut Au composite and Au estimates per domain at zero cut-off 140
Table 14-24: Pit optimisation parameters for evaluation of Mineral Resource classification 143
Table 14-25: Open pit Mineral Resource above 950 mRL at 0.5 g/t Au cut-off – base case 145
Table 14-26: Open Pit Mineral Resource above 950 mRL at 1.0 g/t Au cut-off – for comparison only 145
Table 14-27: Underground Mineral Resource below 950 mRL at 1.5 g/t Au cut-off – base case 145
Table 14-28: Underground Mineral Resource below 950 mRL at 2.0 g/t Au cut-off – for comparison only . 145
Table 16-1: Qualitative risk assessment of study components 148
Table 16-2: Empirical Stope Design Summary 150
Table 16-3: Ground Support Requirements 150
Table 16-4: Preliminary Slope Design Parameters 152
Table 18-1: Areal extent of catchment areas 155
Table 20-1: Preliminary aspects and impacts analysis 175
Table 20-2: Sampled materials (GCA, 1997) 183
Table 20-3: Summary of Available Results, Detailed Geochemical Characterisation Waste Rock and Ore 186
Table 25-1: Mineral Resource Model Risks and Opportunities 195
Table 26-1: Recommended drilling 196

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List of Figures

Figure 4- 1: Newmarket Gold’s general location – Northern Territory gold properties 5
Figure 4- 2: Maud Creek Gold Project Tenements 6
Figure 4- 3: Agreements for Maud Creek Royalty 7
Figure 4- 4: Agreements for Maud Creek Royalty with Biddlecombe 8
Figure 4- 5: Agreements for Maud Creek Farm-in with PNX Metals 10
Figure 4- 6: Existing disturbance at Maud Creek (May 2007) (from URS, 2008) 12
Figure 4- 7: Simplified process diagram – NT environmental assessments 14
Figure 6- 1: 1997 airborne magnetic and radiometric survey 20
Figure 6- 2: Maud Creek area pits and prospects (Independent Engineers) 20
Figure 6- 3: Maud Creek regional geology and structural interpretation by Independent Engineers 2005 21
Figure 7- 1: Pine Creek Orogen (northern orange zone) within the Northern Territory 23
Figure 7- 2: Summary Stratigraphic Chart of the Pine Creek Orogen Source: Ahmad & Hollis, 2013 25
Figure 7- 3: Location Map of the Maud Creek Deposit 27
Figure 8- 1: Structural – stratigraphic model for Newmarket Gold deposits 31
Figure 8- 2: Pine Creek Regional structural interpretation 33
Figure 9- 1: Maud Creek Property with VTEM flight lines 34
Figure 9- 2: Strong VTEM Conductors on a 35 hertz conductor base map 35
Figure 9- 3: Merged 1997 and 2012 Aeromagnetic data with strong VTEM conductors (Card, 2012) 36
Figure 9- 4: Maud Creek Area – 2012 stream sediment survey – selected gold anomalous areas 40
Figure 9- 5: Location of soil anomalies requiring further work 45
Figure 9- 6: Red Queen – Chessman Area – Soil anomaly designation 46
Figure 9- 7: Chlorite Hills and O’Shea’s soil anomaly designations 48
Figure 9- 8: Pussy Shear zone soil anomalies 49
Figure 9- 9: Chlorite Hills and Kittens (Droves) soil anomaly areas 50
Figure 9- 10: Au (ppb) for ionic leach soil results at Chlorite Hills, Maud Creek 52
Figure 9- 11: Ionic leach Au (ppb) in soil results, Chessman Area, Maud Creek 54
Figure 9- 12: Maud Creek area gold in 2012 rock chip sample results 55
Figure 10-1: Plan view of the RC and DDH Drilling completed in 2011 57
Figure 10-2: Long section looking west showing grade shells and 2011 diamond drillholes 58
Figure 10-3: Changes in Azimuth with Depth for drilling prior to 2011 60
Figure 10-4: Changes in Dip with Depth for drilling prior to 2011. 60
Figure 11-1: Reverse Circulation Sampling flow chart 63
Figure 11-2: Diamond Drilling Sampling flow chart 65
Figure 11-3: Laboratory sampling flow chart 67
Figure 11-4: Standard ST202/5355 Compliance chart for Maud Creek Newmarket Gold 2011 drilling program. 70
Figure 11-5: Standard OxH19 71
Figure 11-6: Standard OxE20 71
Figure 11-7: Standard OxE21 72

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Figure 11-8:

Standard OxG22

72
Figure 11-9:

Compliance chart for blanks used in the 2011 Maud Creek drilling program.

73
Figure 11-10:

Duplicate correlation plot; range <0.2 g/t for 2011 drilling program

75
Figure 11-11:

Relative difference plot original vs lab repeats for pre 2011 Maud Creek drill data

76
Figure 11-12:

Repeat correlation plot; range <20 g/t for 2011 drilling program

78
Figure 11-13:

NTEL: ALS Inter-laboratory repeats for all ranges NTEL FA25: ALS AA25 for the 2011 drilling program

80
Figure 11-14:

ALS: Assay Corp Inter-laboratory check analysis for pre 2011 Maud Creek drilling

81
Figure 11-15:

ALS Re-split check assays for pre 2011 Maud Creek drilling

81
Figure 12-1:

Precision plot original vs resample

89
Figure 12-2:

Precision plot fire assay vs screen fire assay

90
Figure 13-1:

Recovery by Grind Size (Extracted from Ammtec Reports A4997 & A9617)

98
Figure 13-2:

Recovery vs. grade (Ammtec tests A5161 and A6067)

100
Figure 13-3:

1997 Pilot Plant Survey Results

101
Figure 13-4:

Cross Section of Maud Creek Deposit

105
Figure 13-5:

Long Section View Facing South- East Showing Silver Data Compared to Ore Domains

107
Figure 13-6:

Long Section View Facing South- East Showing Arsenic Data Compared to Ore Domains

107
Figure 13-7:

Long Section View Facing South- East Showing Bismuth Data Compared to Ore Domains

108
Figure 13-8:

Long Section View Facing South-East Showing Carbon, Carbonate Carbon, CO3, Organic Carbon and Total Carbon Data Compared to Ore Domains

108
Figure 13-9:

Long Section View Facing South- East Showing Sulphide Sulphur, Sulphate Sulphur and Total Sulphur Data Compared to Ore Domains

109
Figure 13-10:

Long Section View Facing South- East Showing Antimony Data Compared to Ore Domains.

109
Figure 14-1:

Maud Creek Deposit lithology model showing primary units, cover not shown

113
Figure 14-2:

Interpreted fault architecture of the Maud Creek Deposit

114
Figure 14-3:

Cross section view looking north (northing 9008mN)

115
Figure 14-4:

Cross-section view looking north of the three modelled veins (primary, upper and lower), contact fault (dark blue) and two of the eight faults modelled (light blue)

116
Figure 14-5:

Vein architecture of the Maud Creek Deposit illustrating three primary veins and indicating a south-eastward plunge

117
Figure 14-6:

Cross section illustrating 0.75 g/t grade shell (yellow), primary contact vein (red) and fault architecture (blue)

118
Figure 14-7:

Cross section looking north (9250 mN) exhibiting low grade hole MCP088 through the fault contact

119
Figure 14-8:

Section 9150N (20 m grid)

120
Figure 14-9:

Section 9175 (20 m grid)

121
Figure 14-10:

Composite length histogram

122
Figure 14-11:

MAIN – Quantile- Quantile plot comparing arsenic from drilling only and including metallurgy samples (ASSMET)

123
Figure 14-12:

MINOR – Arsenic - Left: Plan view of data, Right: Cumulative histogram, Red: Arsenic from drilling only, Green Arsenic from the metallurgical samples

124
Figure 14-13:

MAIN – Left and Middle Au histogram: from 0 to 20 g/t and from 20 to 1000 g/t and Right: Variogram cloud, Red and blue: outliers

127
Figure 14-14: MAIN – Au samples location – Left: XoY view, Right: XoZ view 127

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Figure 14-15:

AU samples collected in the RC drillholes, Red: RCGC, Blue: RC samples above 1,084 mRL and Green: remaining RC

130
Figure 14-16:

Comparison of weighted and top-cut Au mean grade between RC and RCGC above 1084 mRL

131
Figure 14-17:

MINOR – Au distribution – Left: RC above 1,084 m, Right: RCGC

131
Figure 14-18:

Comparison of unweighted Au mean grade per drilling type

132
Figure 14-19:

MAIN variogram fitting – left: Au, Right: As

134
Figure 14-20:

Regression slope histogram for Au estimates (left) and AS estimates (right)

139
Figure 14-21:

Regression slope histogram for Main Au Local check estimate

140
Figure 14-22:

MAIN – Au Global OK – Swath plots

142
Figure 14-23:

MAIN –gold – Grade tonnage comparison Reasonable Prospects of economic extraction

142
Figure 14-24:

Long section facing west displaying measured blocks for all domains

143
Figure 14-25:

Long section facing west displaying indicated blocks for all domains

144
Figure 14-26:

Long section facing west displaying inferred blocks for all domains

144
Figure 16-1:

Plan view of existing pit showing the geotechnical domains used for rock mass characterisation

149
Figure 18-1:

Site Access Roads

154
Figure 18-2:

Catchment areas in Maud Creek site

155
Figure 20-1:

Katherine municipality local government area (ABS, 2011 census)

157
Figure 20-2:

Katherine region (shown in darker tan)

158
Figure 20-3:

Katherine LGA population, by age group (2011 census)

158
Figure 20-4:

Gross regional product, by industry – Katherine NT (2012)

159
Figure 20-5:

Employment by industry sector – Katherine LGA (ABS data)

160
Figure 20-6:

Unemployment rates – Katherine LGA (ABS data)

160
Figure 20-7:

Labour force participation rate, Katherine LGA (ABS data)

161
Figure 20-8:

Estimated flood extents in absence of engineering controls

163
Figure 20-9:

Groundwater aquifers near Katherine (from Katherine land use plan, 2014)

164
Figure 20-10: 

Distribution of Tephrosia humifusa (Atlas of Living Australia, accessed 15/01/2016)

165
Figure 20-11:

Conservation significant flora

166
Figure 20-12:

Vegetation communities

167
Figure 20-13:

Conservation significant fauna

169
Figure 20-14:

Locations of parks and reserves near Maud Creek

171
Figure 20-15:

Aboriginal and non- Aboriginal heritage sites (Figure 11.1 from URS 2008)

173
Figure 20-16: 

Restricted work areas (excerpt from AAPA certificate, June 2007)

181
Figure 20-17:

Drillhole database S assay data

182
Figure 20-18:

Geochemical classification of waste rock samples using NPR

185
Figure 23-1:

Pine Creek Gold Project Tenements

190
Figure 23-2:

Union Reefs Gold Project Tenements

191
Figure 23-3:

Burnside Gold Project Tenements

191
Figure 23-4:

Moline Gold Project Tenements

192
Figure 23-5:

Maud Creek & Yeuralba Gold Project Tenements

192

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2

Introduction

SRK Consulting (Australasia) Pty Ltd (SRK) was engaged by Newmarket Gold Inc. (Newmarket Gold) to undertake a study on the Maud Creek Gold Project (Maud Creek or the Project) and prepare a Technical Report to support the release of the updated Mineral Resource estimate.

In early July 2015, Newmarket Gold merged with Crocodile Gold Corp. (Crocodile Gold) to form a new Canadian; Toronto Stock Exchange listed gold mining company named Newmarket Gold that has 100% ownership of the Maud Creek Project.

This Technical Report documents the review and assessment of the project’s geology, exploration and mineral resource aspects prepared by SRK. It was prepared following the guidelines of the Canadian Securities Administrators’ National Instrument (NI) 43-101 and Form 43-101 F1.

The mineral resource statement reported herein was prepared in conformity with generally accepted Canadian Institute of Mining, Metallurgy, and Petroleum’s (CIM) Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines.

2.1

Scope of Work

The agreed Scope of Work was prepared by SRK in response to a specific request by Newmarket Gold, a review of documents provided by Newmarket Gold relating to the Project and subsequent discussions between Newmarket Gold and SRK.

The Scope of Work called for preparation of an NI 43-101 Technical Report on the Maud Creek Gold Project in compliance with NI 43-101 and Form 43-101 F1 guidelines.

The technical aspects of the scope of work included a review and update of the geological model and release of a Mineral Resource estimate, review of the geotechnical aspects and review of the metallurgical characteristics to support future technical work.

2.2

Work Programme

The Technical Report was assembled in Melbourne during the months of April 2015 to March 2016.

The Mineral Resource Statement reported herein was prepared in conformity with generally accepted CIM Exploration Best Practices and Estimation of Mineral Resource and Mineral Reserves Best Practices guidelines. This Technical Report was prepared following the guidelines of the Canadian Securities Administrators National Instrument 43-101 and Form 43-101 F1.

2.3

Basis of Technical Report

The purpose of this Technical Report is to present the geological review of the Maud Creek Gold Project. This report is based on information provided by Newmarket Gold to SRK and verified during site visits conducted in 2015 and any additional information provided by Newmarket Gold throughout the course of SRK’s investigations. The Qualified Persons have reviewed all relevant information and determined it to be adequate for the purposes of the Technical Report. The Qualified Persons do not disclaim any responsibility for this information. SRK has no reason to doubt the reliability of the information provided by Newmarket Gold.

This Technical Report is based on the following sources of information:

  Discussions with Newmarket Gold personnel;
     
  Inspection of Newmarket Gold’s Maud Creek Gold Project; and
     
  Additional information and studies provided by Newmarket Gold.

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2.4

Qualifications of SRK and SRK Team

The SRK Group comprises over 1,400 professionals, offering expertise in a wide range of Resource engineering disciplines. The SRK Group’s independence is ensured by the fact that it holds no equity in any project and that its ownership rests solely with its staff. This fact permits SRK to provide its clients with conflict-free and objective recommendations on crucial judgment issues. SRK has a demonstrated track record in undertaking independent assessments of Mineral Resources and Mineral Reserves, project evaluations and audits, technical reports and independent feasibility evaluations to bankable standards on behalf of exploration and mining companies and financial institutions worldwide. The SRK Group has also worked with a large number of major international mining companies and their projects, providing mining industry consultancy service inputs.

The compilation of this Technical Report was completed by Peter Fairfield, Principal Consultant (Project Evaluation), BEng (Mining), FAusIMM (No 106754) CP (Mining). By virtue of his education, membership to a recognised professional association and relevant work experience, Peter Fairfield is an independent Qualified Person (QP) as defined by NI 43-101.

Danny Kentwell, Principal Consultant (Resource Evaluation), MSc Mathematics and Planning (Geostatistics), FAusIMM, undertook a review of the Mineral Resources and geological aspects of the project and contributed to the relevant sections in this Technical Report. By virtue of his education, membership to a recognised professional association and relevant work experience, Danny Kentwell is an independent QP as defined by NI 43-101.

Table 2-1:      Site Visits

QP Position Employer Last Site Visit
Date
Purpose of Visit
Peter Fairfield Mining Principal Consultant SRK 13 Aug 2014 Site Inspection
Simon Walsh Processing Principal Consultant Simulus 13 Aug 2014 Site Inspection
Rodney Brown Geology Principal Consultant SRK 13 Aug 2014 Site Inspection
Louie Human Geotechnical Principal Consultant SRK 3-7 Aug 2015 Geotechnical Logging
Tristan Cook Geotechnical Consultant SRK 3-7 Aug 2015 Geotechnical Logging
Kirsty Sheerin Geology Consultant SRK 3-7 Aug 2015 Geological Logging

2.5

Acknowledgement

SRK would like to acknowledge the support and collaboration provided by Newmarket Gold personnel for this assignment. Their collaboration was greatly appreciated and instrumental to the success of this project.

2.6

Declaration

SRK’s opinion contained herein is based on information collected by SRK throughout the course of SRK’s investigations, which in turn reflect various technical and economic conditions at the time of writing. Given the nature of the mining business, these conditions can change significantly over relatively short periods of time. Consequently, actual results may be significantly more or less favourable.

This report may include technical information that requires subsequent calculations to derive sub-totals, totals and weighted averages.

Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, SRK does not consider them to be material.

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SRK is not an insider, associate or an affiliate of Newmarket Gold, and neither SRK nor any affiliate has acted as advisor to Newmarket Gold, its subsidiaries or its affiliates in connection with this project. The results of the technical review by SRK are not dependent on any prior agreements concerning the conclusions to be reached, nor are there any undisclosed understandings concerning any future business dealings.

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3

Reliance on Other Experts

This report has been prepared for Newmarket Gold and is based, in part, as specifically set forth below, on the review, analysis, interpretation and conclusions derived from information which has been provided or made available by Newmarket Gold, augmented by direct field examination and discussion with former employees, current employees of Newmarket Gold.

The Qualified persons have reviewed such technical information and determined it to be adequate for the purposes of this Technical Report. The Qualified Persons do not disclaim any responsibility for this information.

SRK have not performed any sampling or assaying, detailed geological mapping, excavated any trenches, drilled any holes or carried out any independent exploration work.

SRK did undertake geological and geotechnical logging of specific drillholes to assist in validating project assumptions.

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4

Property Description and Location

The Maud Creek Deposit of Newmarket Gold described within this Technical Report is located within the Pine Creek region of the Northern Territory of Australia (Figure 4-1). There are other projects managed and owned by Newmarket Gold in the Northern Territory which are discussed further in Section 23, they include:

 

The Union Reefs Gold Project and processing plant (Figure 4-2), located approximately 170 km south-southeast of Darwin accessible by the Stuart Highway, 18 km north- northeast of the Township of Pine Creek;

   

 

 

The Pine Creek Gold Project;

   

 

 

The Burnside Gold & Base Metals Project:

   

 

 

The Moline Gold and Base Metals Project; and

   

 

 

The Yeuralba Gold and Base Metals Project.

Figure 4-1:      Newmarket Gold’s general location – Northern Territory gold properties

4.1

Property Location

The Maud Creek project comprises a total of 23 mineral titles (all granted) covering a total area of 29,489 ha (294.89 km2), as follows summarised in Figure 4-1.

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Table 4-1:      Summary of Mineral Titles Maud Creek Deposit

Licence Type Number Area (km2)
Exploration Licence
Exploration Licence (EL) 2 280.26
Sub-total 2 280.26
Mineral Leases
Mineral Lease (ML) 2 12.25
Sub-total 2 12.25
Total 4 292.51

The Project is located 20.7 km east northeast of the regional administrative centre of Katherine (population 9,800) and is just east of the Township of Katherine. The deposit and proposed infrastructure is located on ML30260 which was granted 14 April 2014 and will expire on the 13 April 2024.

Geographically, the Project is centred about 6.7 km straight line distance northeast of the Stuart Highway, 287 km southeast of Darwin (population 129,100), the capital city of the Northern Territory (population- 233,300), at Latitude 14°26’41”S Longitude 132°27’10” and UTM (AMG) coordinates (WGS 84, Zone 53L) 225407mE and 8401561mS, elevation 131 mASL.

Figure 4-2:      Maud Creek Gold Project Tenements

4.2

Land Tenure

The Maud Creek Project area lies within land traditionally owned by the Jawoyn people, who continue to exercise their traditional cultural attachment to the Katherine region as the owners and co-managers (with the NT Parks and Wildlife Commission) of the Nitmiluk National Park. The project lies on freehold land (NT Portion 4192, a subdivision of Portion 4159), outside lands administered by the Jawoyn Aboriginal Land Trust and the proposed mining operations area does not intersect any other Aboriginal land trust parcels.

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The key mining tenement required for implementation of the Maud Creek project is ML 30260. Newmarket Gold was granted tenure over ML30260 on 14 April 2014. The current tenure expires on 13 April 2024, at which time the tenement holder has the option to renew its holdings, providing it has adhered to tenement conditions and to reporting and expenditure obligations.

There are no registered or determined native title claims over the Project area.

A land use agreement is in place for the titles shown in Figure 4-3. This is termed the Michell Compensation Agreement, which was original signed in 1992 between Michell, Biddlecombe and Trescabe to compensate the landholder for being deprived of the use of the surface of land, any damage to the property through exploration activities and being deprived of land improvements. This agreement has been assigned and accepted by Newmarket Gold.

There is an agreement between Newmarket Gold- and the estate of Robert Biddlecombe relating to titles EL7775, EL8018 and MCN’s 4218 4225, (Figure 4-4), inclusive relating to a royalty payment for the mining of gold on these tenements. This is discussed further in Section 22.2.5.

Figure 4-3:      Agreements for Maud Creek Royalty

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Figure 4-4:      Agreements for Maud Creek Royalty with Biddlecombe

4.3

Underlying Agreements

   
4.3.1

Royalties

The following is a summary of the agreement and royalties, provided to SRK by Newmarket Gold, further detail is presented in Section 22.2.5.

Government Royalty – payable to the Northern Territory under the Mineral Royalty Act (NT). The royalty rate is 20% of the net value of a saleable mineral commodity (in this case the gold concentrate) sold (or removed without sale) from a production unit (i.e. ML 30260) in a royalty year. The net value for the production is calculated by the following formula:

Net Value = Gross Realization – (operating costs + capital recognition deduction + eligible exploration expenditure + additional deductions)

Harmony Royalty – payable to Harmony Gold Operations Limited pursuant to a Deed of Assignment and Assumption dated 2 September 2009. Applies to all of ML 30260. The royalty rate is 1% of the value of all Gold as defined in the agreement (i.e. the Perth Mint price for gold with no deductions). The royalty is not payable before 250,000 ounces of gold produced. Note also the Decision to Mine payment of AUD2M (indexed to CPI).

Virotec Royalty – payable to Mt Carrington Mines Pty Ltd pursuant to a Deed of Assignment and Assumption dated 2 September 2009. Applies only to that part of ML 30260 that was formerly within MCNs 4218 to 4225. Royalty rate is AUD5.00 per ounce with respect to 80% of the gold produced.

Biddlecombe Conglomerate Royalty – payable to Robert Biddlecombe (estate) pursuant to a Deed of Assignment and Assumption dated 2 September 2009. Applies only to that part of ML 30260 that was formerly within MCNs 4218 to 4225. Royalty rate is 1% of the gross value received as sale proceeds of all mineralization, metals, minerals and other products, after payment of the expenses incurred in smelting and refining charges.

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Note that the above summaries are based on a plain reading of the documents.

NT Build Levy

Although it is not a royalty, for the purposes of the financial modelling of the Maud Creek project the construction works associated with the Maud Creek project may be subject to a levy under the Construction Industry Long Service Leave and Benefits Act (NT) (Levy Act).

In summary, the Levy Act imposes a levy on Construction Work in the Territory where the costs of construction work, commenced after 7 April 2014, are AUD1 million or more. The levy must be paid, prior to construction works commencing, by the person for whom the work is to be done (i.e. Newmarket Gold). The definition of Construction Work applies to civil works and works for buildings and structures that form part of the land, including a range of repair and maintenance works with respect to such civil works or buildings and structures.

The levy rate for construction works that commence after 7 April 2014 is 0.1% of the costs of the construction work. The Levy Act specifies that the costs of Construction Work is the total contract prices for all the construction contracts in relation to the work.

4.3.2

Farm-out Agreement

Farm-out agreements provide for third parties to explore on mineral titles, which are not owned 100% or substantially controlled by Newmarket Gold. The following discusses agreements relevant to the Maud Creek Project.

On November 6, 2013, Thundelarra Exploration Limited Uranium Exploration (Thundelarra) withdrew from a joint venture agreement with Newmarket Gold. Thundelarra was replaced by Rockland Resources Pty Ltd (Rockland) as party to the joint venture agreement, a 100% owned subsidiary of Oz Uranium Pty Ltd. Rockland was then replaced as a party to the agreement with Oz Uranium Exploration Agreement for the Pine Creek Tenements. Rockland Resources Pty Ltd (Rockland), a wholly-owned subsidiary of Oz Uranium, and Crocodile Gold formed a joint venture on November 6 2013, in regards to uranium exploration and development on the Maud Creek, Burnside, Cosmo, Pine Creek, Union Reefs and Moline projects. Rockland has a minimum expenditure commitment of AUD1 million over the next four years. Rockland has the rights to apply for a mining tenement in its own right as long as it does not conflict with Newmarket Gold’s operations.

Over the past 12 months Rockland has been active in the Pine Creek region. They have conducted regional scale geophysical surveys and reviews (VTEM) as well as geochemical analysis, structural mapping and drilling in and around their currently identified uranium deposits. While one prospect is close to the Cosmo Mine (Fleur de Lys) no work has been conducted by Rockland on MLN993.

Land and Mining Property Swap Agreement - 2008

Land & Mining Property Exchange Deed (unregistered) and Land Use Deed dated 2008 (unregistered).

Parties involved:

 

GBS Gold Australia (Land Holdings) Pty Ltd and Terra Gold Mining Pty Ltd (previously Terra Gold Mining Limited); and

   

 

 

Teelow Nominees Pty Ltd and Michael Daniel Teelow.

BACKGROUND

This agreement relates to the transfer of the Moline Project from Michael Teelow to GBS Gold in exchange for the transfer of the Maud Creek farm and NT portion 4192 from GBS to Michael Teelow. The titles transferred to GBS included MLN 41 and 1059, EL’s 23605, 22966, 22967, 22968, 22970, 24262 and 24127 and MLA 24173.

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PARTICULARS

As part of the property swap agreement Teelow, the lease owner, has the first right of refusal on any land sale at the Maud Creek farm. The tenement owner also has the Right of way over the property which allows the tenement owner the rights to establish easement over the farm. Newmarket Gold exercised this right and has established two easements over the farm for future access to the mine.

4.3.3

Farm-in Agreement

In 2014, Phoenix Copper Pty Ltd (now PNX Metals) entered into a Farm-in agreement with Newmarket Gold. The Heads of Agreement was signed in August 2014 and was completed in December 2014. The Farm-in agreement relates to exploration activities on the Burnside Exploration Licenses as well as at the Chessman (close to Maud Creek) and Moline projects.

The Farm-in Tenements include the Maud Creek Project including exploration licenses EL25054 and EL28902, and mineral lease ML30293.

The Farm-in agreement will allow Phoenix to earn up to 51% through the spending of AUD2 million on exploration activities over a two year period. They can then earn a further 39% by spending an additional AUD2 million for another two year period. This will potentially take their ownership of these projects to 90%. Newmarket Gold retains a claw-back right to precious metal discoveries. While this agreement is not over the Cosmo Mine area, it covers the exploration licenses that surround the Cosmo Deposit.

PNX Metals has been active since signing the Heads of Agreement in August 2014.

Figure 4-5:     Agreements for Maud Creek Farm-in with PNX Metals

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4.4

Environmental Liability

In addition to any environmental impacts that would arise in connection with new mining or mineral processing activities, Newmarket Gold would generally be liable for the management and eventual rehabilitation of legacy impacts present at the Maud Creek site at the time that Newmarket Gold took ownership of the Project.

Exploration for a range of commodities (copper, molybdenum, uranium, gold) has occurred intermittently at Maud Creek since approximately 1890. Small scale gold mining and a limited amount of processing is reported to have occurred at the site in 1890 – 1891 and again in 1932 – 1934 (Mining One, 2013).

The most recent mining at Maud Creek occurred in 2000, when Katherine Mining NL conducted open cut mining for gold. Ore was treated offsite. In the order of 9 ha of disturbed land (comprising 2.7 ha associated with the pit void, 1.6 ha associated with the former ROM pad and approximately 4.7 ha occupied by a waste rock dump) remain from previous mining. Minor disturbance related to support infrastructure (access tracks, relocatable offices) also remains (Figure 4-5).

Vegetation mapping conducted in 2007 as part of baseline environmental studies for an environmental impact assessment of the proposed Terra Gold mining project at Maud Creek mapped an area of approximately 14 ha as cleared for mining and the Terra Gold EIA reported that Approximately 96 ha of savannah woodland vegetation has been cleared in the Maud Creek project area for pastoral development and to support exploration and historical mining activity (Crawford and Metcalfe, 2007; URS, 2008). Exploration disturbance (drillholes) arising during exploration activities subsequent to Crocodile Gold’s acquisition of the Maud Creek tenements in 2009 is reported to have been rehabilitated (Mining One, 2013).

In addition to direct clearing, historic mining activity is likely to have contributed to the establishment and spread of weeds, which are reported to be abundant and well established at Maud Creek, with dense infestations, especially along access roads and drainage lines and in disturbed areas. Some of the weeds recorded in the Project area are declared weeds under the NT Weed Management Act. Landholders are required to make a reasonable attempt to control and prevent the spread of declared weed species (Department of Land Resource Management, 2015).

Legacy features from previous mining at Maud Creek (notably the waste rock dump and pit void) may represent a potential source of acid or metalliferous drainage, however the limited surface and groundwater quality data for the site do not so far show the presence of significant impacts: water quality downstream of the site and in water storages at the site is generally in the range of values recorded upstream.

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Figure 4-6:      Existing disturbance at Maud Creek (May 2007) (from URS, 2008)

Mining activities in the Northern Territory are fully bonded. That is, the NT government requires lodgement of a security to cover 100% of the estimated cost of rehabilitating any disturbance proposed by the tenement holder under its Mining Management Plan (MMP), together with the cost of rehabilitating any pre-existing disturbance on the tenement. Additionally, from 2013, tenement holders are required to make annual contributions to the Mine Rehabilitation Fund (MRF) established under the Mining Management Act. The MRF payments are a non-refundable annual levy of 1% of the total calculated rehabilitation cost applied to each mining operation.

SRK has not independently estimated the cost of rehabilitating existing disturbance within the Maud Creek tenements as part of this initial review.

Permitting of the Project would require the preparation of a Mining Management Plan (MMP), including a cost estimation of mine rehabilitation and closure works. The NT Government has developed an Excel spreadsheet for estimating the security deposit to be lodged to cover mine rehabilitation and closure.

4.5

Legislation and Permitting

In the Northern Territory environmental impact assessment and subsequent authorisation and regulation of the implementation of mining and related support activities is chiefly administered until three Acts:

  The Environmental Assessment Act 1982,
     
  The Mining Management Act 2001, and
     
  The Waste Management and Pollution Control Act 2009.

The Mineral Titles Act 2010 also exerts a considerable influence on regulation of environmental aspects of mining activities in that it provides for a number of significant exemptions to licensing provisions under other Acts that would otherwise apply.

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If a mining proposal has the potential to give rise to significant adverse impacts on a Matter of National Environmental Significance then it may also require referral to and assessment by the federal Department of the Environment (DotE) under the Environment Protection and Biodiversity Conservation Act 1999 (EPBC Act). A simplified flow chart showing the environmental impact assessment process is presented in Figure 4-6.

Although there are some differences in the duration and fine detail of the Public Environmental Review (PER) and Environmental Impact Statement (EIS) pathways, the overall processes are similar. The red arrow on the figure shows the point at which Terra Gold’s environmental permitting was terminated by the proponent.

The Maud Creek Project currently contemplated by Newmarket Gold has not been referred to the Northern Territory Environmental Protection Authority (NTEPA) or DotE for assessment. In 2006 Terra Gold referred a proposal for mining and processing of mineralization from Maud Creek to the NTEPA for assessment. The EPA determined that Terra Gold’s Maud Creek proposal should be assessed via the EIS pathway. A draft EIS report was issued for public review and comment in early 2008, but the proposal was ultimately withdrawn without completing the assessment process. Terra Gold also referred its Maud Creek proposal to the Commonwealth in 2006. In early 2007, the federal government determined that the Project was not a ‘controlled action’ under the EPBC Act.

It is likely that any future assessment of the Maud Creek project would follow a similar assessment path. Although much of the information produced for Terra Gold’s EIS would still be relevant, it would almost certainly be necessary to recommence the Project’s environmental assessment from the NOI stage (i.e., it would not be possible to re-activate the project assessment starting at the public exhibition phase). There is no fixed statutory timeline for the EIS process. A minimum of 18 to 24 months is typically required to progress from the NOI stage to completion of the EPA assessment.

Once the EIS process has been completed, permitting of operational aspects of the Project would be administered primarily under the Mining Management Act and the Waste Management and Pollution Control Act. Apart from the approvals required under these Acts, a range of authorisations or implementation conditions may arise under the following legislation:

 

Water Act 1992;

   

 

 

Heritage Act 2011;

   

 

 

Northern Territory Aboriginal Sacred Sites Act 1989;

   

 

 

Planning Act 1999;

   

 

 

Dangerous Good Act 1998;

   

 

 

Transport of Dangerous Goods by Road and Rail (National Uniform Legislation) Act 2011;

   

 

 

Territory Parks and Wildlife Conservation Act 2000;

   

 

 

Native Title Act 1993 (Cth); and

   

 

 

Aboriginal Land Rights (Northern Territory) Act 1976 (Cth).

Additional information on permitting of specific environmental and heritage aspects of the Project is presented in Section 20.2

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Figure 4-7:      Simplified process diagram – NT environmental assessments

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5

Accessibility, Climate, Local Resources, Infrastructure and Physiography

The Northern Territory is the least populated of all areas in Australia. It encapsulates a total area of 1.35 million square kilometers and accounts for 20% of the whole country; however, just 233,300 (ABS March 2012) or 1% of Australia's population reside there.

The Territory varies considerably in topography, climate, and infrastructure. The Top End, where the Northern Territory Gold Properties are located, is home to the Aboriginal Arnhem Land which includes the Kakadu National Park. The region is dry between April and September, and wet between October and March. During the wet season everything is green and there is no dust; however, the humidity and temperatures are high and access off road is difficult.

The centre is extremely arid, with greatly varying temperatures and is known as the Red Centre named because red is the predominant color found in the soil.

Darwin, Capital of the Northern Territory, lies on the coast to the north and provides the majority of infrastructure support and services for the mining industry. The Stuart Highway, which virtually bisects the country, is the main road that leads from Darwin to Alice Springs then on to Adelaide in South Australia.

5.1

Accessibility

Access is gained to the Project from Darwin by travelling south for some 314 road kilometers along the sealed Stuart Highway to the town of Katherine.

The Stuart Highway, the area’s major thoroughfare, and the Adelaide-to-Darwin transcontinental railway line bisect Australia in a north-south sense and provide access to the Maud Creek Property. The Project site is approximately 30km from the town of Katherine.

The Union Reefs processing plant, owned by Newmarket Gold is located approximately 185 km southeast of Darwin, 15 km north of the town of Pine Creek.

5.2

Land Use

Major land uses are traditional Indigenous uses, nature conservation (including parts of Kakadu National Park and World Heritage Area and Litchfield National Park), urban and other intensive uses and grazing. Approximately 85,000 hectares have been cleared.

The region has undergone some localized clearing and the major land uses are grazing, nature conservation (including parts of Kakadu National Park and World Heritage Area and Litchfield National Park), traditional Indigenous uses and other intensive uses including horticulture.

The Daly Basin Bioregion consists of gently undulating plains and scattered low plateau remnants and has a tropical monsoonal climate with distinct wet and dry seasons and high temperatures throughout the year. Dominant vegetation is tropical eucalypt woodlands/grasslands and eucalypt open forests. Smaller patches of eucalypt woodlands and melaleuca forests and woodlands are present.

The major land use is grazing on native pastures and traditional Indigenous uses with some horticulture, grazing on modified pastures and nature conservation. The region has undergone some clearing (approximately 167,000 ha) for these developments the vegetation of the Maud Creek area consists largely of woodlands and open woodlands (predominant species – Eucalypts) that have been degraded by the impacts of cattle, buffalo and wild donkeys. No rare, threatened or endangered species have been identified in the area.

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5.3

Topography

Generally the topography of the Property area is flat, locally gently undulating.

In the Maud Creek area, the terrain is flat lying to undulating. Ephemeral streams transect the project area and drain into the westward flowing Katherine River that flows all year. Land units occurring within the Maud Creek area include:

 

Rugged terrain with slopes 15 to 40% with shallow or skeletal soils;

   

 

 

Hilly terrain with slopes 5 to 15%, rocky and boulder strewn with shallow and skeletal soils;

   

 

 

Gently undulating crests and upper slopes to 5% with shallow rocky soils;

   

 

 

Undulating terrain with slopes 5 to 10% with grey and brown clays; and

   

 

 

Major creeks and gullied tributaries.


5.4

Climate

The Top End of the Northern Territory has a tropical monsoon climate characterized by two distinct seasonal patterns: the ‘wet’ monsoon and the ‘dry’ seasons. The wet season generally occurs from November through to April and the dry season between May and October. Almost all rainfall occurs during the wet season, mostly between December and March, and the total rainfall decreases with distance from the coast.

The mean daily maximum temperature, as recorded at Darwin on the northern coastline, is 31°C in the coolest months of June to August and 33°C in the hottest months of October and November. The mean daily minimum temperature in Darwin range from approximately 19°C (dry season) to 25°C (wet season). The average annual rainfall at Darwin is 1,713 mm.

The mean daily maximum temperature, as recorded at Katherine, is 31°C in the coolest months of June to August and 38°C in the hottest months of October and November. The mean daily minimum temperatures at Katherine range from approximately 13°C (dry season) to 24°C (wet season). The average annual rainfall at Katherine is 971 mm.

During the wet season, high intensity rainfall events are common, resulting in local flash flooding of ephemeral streams and watercourses. Mining operations are continuous throughout the year; however, increased ore stockpiling is undertaken in the lead up to the wet season thereby offsetting the reduced mining movements over that period. Experience has shown that it is best to shut down ore hauling during periods of extreme rainfall as damage to haul roads by large trucks may occur quickly.

The annual evaporation rate remains high throughout most of the Northern Territory, ranging from 2,400 mm to 4,000 mm per annum. Monthly evaporation exceeds rainfall for eight months of the year at the coast increasing to the whole year inland. It remains relatively high even during the wet season.

Climate gradually moves from seasonally wet tropical in the north to arid in the south, with corresponding changes in landscape, with areas of rocky escarpment and plateau which break a low relief in the north and rocky ridges in the south.

The Northern Territory has a diversity of vegetation that is maintained by its variety of climate and soils. Natural vegetation of the Properties is typical of savannahs of the northern part of Australia, dominated by Eucalypt species with a grassy understorey dominated by sorghum species. The Northern Territory is the only area in Australia that does not have conspicuous temperate flora.

In the north, the vegetation is typically tropical savannah (eucalypt woodland and eucalypt open woodland with a grassy understory). This landscape experiences dramatic seasonal changes with intense growth in the wet season (summer) and widespread fires in the dry season (winter). Famous worldwide for the tropical wetlands and rugged sandstone escarpments of Kakadu National Park, the wetlands are of importance for conservation, providing breeding areas, habitat and refuge for important wildlife populations.

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From the north, a transition area moves from eucalypt woodlands into areas of melaleuca and acacia forests and woodlands and south into the spinifex (hummock grasslands), Mitchell grass (tussock grasslands) and acacia woodlands and shrublands. The vegetation increases in diversity around Alice Springs with areas of mulga, mallee, chenopods, hummock grasslands, small pockets of eucalypt woodlands and salt lakes.

5.5

Infrastructure and Local Resources

Darwin has a population in excess of 129,000 and is the capital city of the Northern Territory. It is the administrative centre of the Northern Territory government and a major transportation hub, with an international airport and deep-water port and the Adelaide to Darwin transcontinental railway terminating at the East Arm port. As the largest city in the Northern Territory, Darwin also has excellent schools, hospitals, and retail, commercial and light industrial services.

A considerable proportion of consumer and other goods reaching the Northern Territory are brought by road from Queensland or South Australia. The Stuart, Arnhem, Kakadu, Barkley and Victoria Highways ensure high service levels to the Darwin region from the Australian capitals and other regional centres.

Despite its low population, the area between Darwin and Katherine in the Northern Territory is well serviced with infrastructure. Significant mining operations have been developed in the area over the past 30 years, with gold ore mining and processing operations conducted within or in close proximity to the project areas at Cosmo Howley, Brocks Creek, Pine Creek, Mount Todd and Union Reefs.

Katherine is a regional centre with a population of approximately 9,800 and enjoys excellent infrastructure, services and communications. This is the closest centre of population to the Maud Creek project.

The regional mining communities of Pine Creek (with a population of 450) and Adelaide River (population of 200) support the Burnside, Maud Creek and Moline gold projects.

The Arnhem Highway to the east southeast of Darwin provides a communication link to the Kakadu National Park and Jabiru, a town of 1,135, which provides accommodation for the uranium mines in the vicinity. Accommodation and services are available along the highway, primarily for the tourist trade.

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6

History

   
6.1

Introduction

Gold was initially discovered in the Maud Creek area in 1890 and a small plant was set up but ultimately abandoned in 1891. This is now called the Chlorite Hills and O’Shea’s area.

The area was re-looked at from 1932-34 when 400 tonnes of ore -produced 540 ounces of gold. Mining was from about 20 shallow shafts and small holes that were 6 - 12 meters deep with horizontal workings from 15 30 meters in length in the Chlorite Hills and O’Shea’s area.

Interest in the Maud Creek area was rekindled in the 1960s during an assessment of the mineral potential of the Top End of the Northern Territory. This study was prompted by the discovery of-significant uranium mineralization in the nearby South Alligator River valley in the mid 1950s.

The Northern Territory Geological Survey carried out IP surveys, soil sampling and petrographic investigations in the late 1970s as part of an assessment of an extension to the nearby township of Katherine.

6.2

Ownership and Exploration Work

Between 1966 and 1973 several companies including Western Nuclear Australia and Magnum Exploration NL explored the area for copper, gold and uranium. IP surveys and drilling of siliceous and gossanous breccias intersected low, albeit anomalous, concentrations of copper and molybdenum and numerous pyritic zones.

In 1973 Magnum Exploration NL (EL147) explored the breccia in the Red Queen/Chessmen area as part of a copper-uranium search. They considered the breccia to be similar to the Rum Jungle occurrence. They drilled 7 holes into the breccia and met with pyritic material with low copper values. They also dug trenches, and obtained anomalous copper and molybdenum values. They did not assay for gold.

In 1985 the Minerals and Exploration and Development Group (MEDG) of CSR Ltd explored the Maud Creek area. Stream sediment sampling returned a 1.3 ppb BLEG gold result about 1.3 km to the west of the old 19th century workings on Maud Creek (now called Chlorite Hills and O’Shea’s).

Placer Exploration purchased MEDG in 1987 and followed up on the BLEG anomaly with rock chip sampling and drilling which subsequently resulted in the discovery of the Gold Creek Zone.

Placer sold the deposit in 1992 to Kalmet Resources NL.

CSR Limited in 1986 held the Peckham Hill EL4874 that covered the Chessmen/ Red Queen prospect (located to the northwest of the Gold Creek Zone). They recognized that the breccia in the area previously mapped by the Bureau of Mines displayed epithermal textures. They conducted exploration programs comprising rock chip sampling, soil sampling, trenching and drilling at Red Queen and along strike. Some 10 km of strike of the breccia/veins were rock chipped and anomalous gold in a gossan sample was reported at Red Queen up to 6.63 g/t Au and 1.02% arsenic (AMG 8406034mN 221372E). The samples were also anomalous in copper, antimony, mercury and thallium.

Another gold anomalous breccia was located 1 km NE of Red Queen and this assayed 0.23 g/t Au. The remainder of the siliceous breccia gave values below 0.07 g/t Au. The north Chessmen location gave an anomalous stream sediment value (14.3 ppb gold at AMG 8406800N 331300E).

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Soil sampling over the Red Queen prospect produced a peak value of 0.33 g/t Au at the western edge of the volcanic-sediment contact. The anomalous breccia also gave positive soil values. Another zone was detected in the north with a value of 0.32 g/t Au in sediments, but without obvious structural association.

Trenching by CSR was carried out in the Red Queen area to test the soil anomalies numbered T1 to T7. T1 coincided with a Magnum Exploration (1973) trench and anomalous gold, arsenic and copper were reported with the best gold value at 6.72 g/t Au from quartz veins in mafic fragmentals. T2, also in an adjacent Magnum trench gave a peak value of 0.61 g/t Au. Trench T3, located 30 m to the SW, met with a carbonated zone with a maximum value of 1 m @ 0.42 g/t Au. T4, on a soil anomaly, met with 2 m @ 0.33 g/t Au. T5 and T6 did not explain the soil anomaly. T7 met with a best result of 2 m @ 0.46 g/t Au.

Percussion drilling was carried out totalling 1,210 m in 9 holes (CMPDH series). - While promising mafic lithologies and chalecedony quartz-carbonate alteration were met with, the results were sub economic with the best values falling in the range 0.1 - to 0.4 g/t Au over intervals up to 19 m.

Between 1989 1990 Placer re-established the CSR grid at Red Queen/Chessmen and conducted soil sampling (412 samples) on a 25 m grid. Samples were assayed for gold, copper, lead, zinc and arsenic. They reported two zones of gold anomalism, one over quartz veins and sheared chert and the other with the western contact zone. The anomalies displayed little correlation with the CSR soil anomalies. Elevated base metal values appeared to correlate with the hematite stained volcanics.

Rock chip sampling and mapping comprised 29 samples and assays up to 4.1 g/t Au were obtained from hematite float in the western contact zone.

Two lines of IP, dipole-dipole, were carried out. A chargeability anomaly was attributed to black cherts. Ground magnetics were carried out over two airborne magnetic anomalies.

Five RC holes were drilled at Red Queen for 576 m. (RQP) Hole RQP5 drilled under a soil anomaly near the IP chargeability anomaly met with 10 m @ 0.95 g/t Au from 46 m and 8 m @ 0.97 g/t Au from 60 m. The associated lithology was black cherts. In general, attempts to correlate surface geology with the drilling proved confusing.

The Chessmen/Red Queen area was reduced to three MCNs and held as a second priority resource area following the discovery of the Main Zone gold deposit at Gold Creek. The trenches and drillholes were rehabilitated.

The Maud Creek project was owned by a number of companies until the acquisition of the project by GBS Gold in December 2006. Substantial drilling, in the order of 66,000- 90,000 meters of RC and diamond drilling, is reported on the project area during the period 1966 – 2006, oriented toward gold exploration.

During 1985 and 1986, CSR Limited (CSR) explored the area in an attempt to locate gold mineralization in the Lower Proterozoic dolerites. Work included 25 meter line spacing airborne magnetic and radiometric surveys, stream sediment sampling, soil sampling, rock chip sampling, petrographic sampling and trenching.

Between 1993 and 1997, Kalmet Resources NL (Kalmet) completed a series of drilling programs at Maud Creek. Metallurgical testing of five high-grade RC samples was completed and an environmental impact study was commissioned. Metallurgical studies showed the primary gold- bearing sulphide mineralization was refractory in nature and bio- oxidation tests were initiated. A close line spaced airborne magnetic and radiometric survey was contracted over a significant part of the land position (Figure 6-1). Interpretation of this radiometric data was completed by Independent Engineers in 2005 (Figure 6-2 and Figure 6-3).

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Figure 6-1:      1997 airborne magnetic and radiometric survey

Figure 6-2:      Maud Creek area pits and prospects (Independent Engineers)

Note: interpreted from radiometric data

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Figure 6-3:      Maud Creek regional geology and structural interpretation by Independent Engineers 2005

In 1997, Kilkenny Gold NL acquired Kalmet and undertook RC and diamond drilling. They carried out significant drilling and increased the global resource to 995,000 ounces (Indicated and Inferred Mineral Resources).

Further metallurgical test work was completed including pilot scale flotation and bio-oxidation program. In 1998, Signet Engineering completed a full feasibility study for the extraction and processing of oxide, transition and primary mineralization from Maud Creek. A comprehensive draft Environmental Impact Study was also produced.

In 1998 Kilkenny Gold commissioned SRK to complete a structural assessment and interpretation of aeromagnetic results. The report, maps and interpretation are quite detailed.

A major flood in the Katherine area in 1998 saw the loss of a significant amount of technical data in the form of reports and diagrams.

AngloGold acquired rights to mine the oxide zone of the Main Zone deposit at Maud Creek and treat the ore at the Union Reefs plant. Mining operations were conducted during 2000. A total of 173,581 tonnes grading 3.32 g/t Au for 18,527 ounces were obtained. Ore was trucked from Maud Creek to the Union Reefs mill.

Hill 50 Gold NL acquired the Maud Creek project from Phoenix Mining Ltd in March 2001 and conducted an extensive review of previous exploration, which identified five gold targets within the property. A program of rock chip sampling was conducted at the Runways prospect. Additional RC and diamond drilling was completed at Gold Creek and surrounding prospects.

In late 2001, Harmony Gold Company Ltd launched a takeover bid for Hill 50 and by mid 2002 had successfully completed the acquisition of the company and all its assets including the Maud Creek deposit. A photo geological interpretation of the property was completed by Snodin (2002), who recommend areas for further follow-up work.

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In December, 2004, and prior to being acquired by GBS Australia, Terra Gold Mining Pty Ltd (TGM) purchased an option to acquire the Maud Creek project from Hill 50, which by that time had been acquired by Harmony Gold. In January 2005, TGM drilled a single combined percussion- diamond hole into the mineralized Main Zone to supply a limited quantity of sample for metallurgical test work purposes. Following a preliminary due diligence examination, in May 2005 TGM exercised its option to purchase the Maud Creek project.

In 2005, four holes totalling 711 meters consisting of 406 meters of RC pre collar and 305 meters of HQ3 diamond drill core were completed. These holes were designed primarily to provide additional samples for metallurgical test work.

In August 2005, GBS Gold Australia Pty Ltd, a wholly owned subsidiary of GBS Gold, announced its intention to acquire all of the issued share capital of Terra Gold Mining, including its interest in the Maud Creek Gold Project. The acquisition was completed in January 2006 and the 100% interest was transferred to GBS Gold.

GBS completed resource calculations as well as mining, geotechnical and hydrogeological studies of the Maud Creek deposit. They also completed an extensive EIS report on the deposit area.

An agreement to acquire a number of properties, including the Maud Creek property, was entered into on June 19, 2009 from GBS Gold International Inc. (GBS Gold) (in liquidation). GBS Gold operated the Tom’s Gully and Brock’s Creek underground gold mines, mined several open pit gold deposits and operated two gold processing plants, one at Tom’s Gully, the other at Union Reefs, near Pine Creek, Northern Territory, until September, 2008, when administrators were appointed.

On November 6, 2009, the mining tenements including the Maud Creek Property were registered in the name of Crocodile Gold, which became Newmarket Gold in July 2015.

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7

Geological Setting and Mineralization

   
7.1

Regional Geology

The Maud Creek Project lies within the Archean to Paleoproterozoic Pine Creek Orogen (PCO) which is located in the north of the Northern Territory and extends from Katherine in the south to Darwin in the north (Figure 7-1). The PCO is exposed over 47,500km2 and consists of a deformed and metamorphosed sedimentary basin with a thickness of over 4 km and overlies a Neoarchean (ca 2670-2500 Ma) granitic and gneissic basement (Ahmad & Hollis, 2013). The PCO hosts over a thousand mineral occurrences and is recognised as one of the most prospective mineral provinces within Australia (Ahmad & Hollis, 2013). Known resources include uranium, gold, and platinum group metals (PGMs), as well as substantial base metals, silver, iron and tin-tantalum mineralization.

Figure 7-1:      Pine Creek Orogen (northern orange zone) within the Northern Territory

(Source: Ahmad & Hollis, 2013)

The basement terrain of the PCO consists of a series of late Archean granite-gneiss basement domes which have subsequently been overlain by fluvial to marine sedimentary sequences of the Paleoproterozoic. These sequences have been divided into the Woodcutters and Cosmo Supergroups, which are separated by a major unconformity representing a time break of 160 Ma (Ahmad and McCready, 2001). Several highly reactive rock units are included within Cosmo Supergroup including carbonaceous shale, iron stones, evaporite, carbonate and mafic to felsic volcanic units of the South Alligator and Finniss River Groups. A northwest trending fabric is evident throughout these sequences resulting from greenschist facies metamorphism and multiphase deformation. A period of widespread felsic volcanism in aureoles between 500 m and 2 km wide overprint the earlier regional metamorphism following deposition of the Cosmos Supergroup (Snowden Report, 2008). A period of extension deformation following intrusions of these granitoids resulted in an extensive array of northeast and northwest trending dolerite dykes intruding the metasedimetary sequences.

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Gold mineralization within the PCO is defined as orogenic in nature and is recognised to have common geological, geochemical, mineralogical and thermochemical characteristics (Ahmad & Hollis, 2013). Gold mineralization is commonly strongly structurally controlled within the region with gold exploiting structures such as anticlines, strike slip shear zones and duplex thrusts as well as located in proximity to the Cullen Granite Batholith (Snowden Report, 2008). Mineralization is commonly recognised within the upper Woodcutters Supergroup and Cosmos Supergroup, specifically within the South Alligator Group and lower parts of the Finnis River Group. Of particular stratigraphic importance for mineralization are the Wildman Siltstone of the mount Partridge Group and the Koolpin Formation, Gerowie Tuff, and Mount Bonnie Formation of the South Alligator Group and the Burrell Creek Formation of the Finnis River Group as well as the Tollis Formation (Figure 22) (Snowden Report, 2008). Goldfields hosted within these units include Pine Creek, Mount Todd, Howley, Golden Dyke, Maud Creek and Brocks Creek gold fields (Ahmad & Hollis, 2013). Descriptions of these prospective host units are briefly summarised as follows.

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Figure 7-2:      Summary Stratigraphic Chart of the Pine Creek Orogen Source: Ahmad & Hollis, 2013

The Wildman Siltstone is the upper most unit of the Mount Partridge Group of the Woodcutters Supergroup and consists of a succession of laminated banded silty pyritic carbonaceous phyllite with minor sandstone and tuff beds, with an overall thickness of approximately 1,000 m. This unit is unconformably overlain by the Koolpin Formation of the South Alligator Group (Ahmad & Hollis, 2013).

The South Alligator Group is the oldest member of the Cosmos Supergroup and is divided into three units (Koolpin Formation, Gerowie Tuff, and Mount Bonnie Formation). Compositionally this group consists of a succession of iron rich sedimentary rocks, tuff, carbonate rocks, shale, greywacke and siltstone (Snowden Report, 2008). The Koolpin Formation is the lowermost unit of the South Alligator Group. It consists of sulphidic and carbonaceous argillite, ferruginous chert, ironstone, silicified dolomites and phyllitic mudstones which were deposited in a low energy environment (Ahmad & Hollis, 2013). The Koolpin Formation varies in thickness from less than 300 m to in excess of 1000 m. The Gerowie Tuff is up to 750 m thick and is comprised of mudstone, siliceous shale, siltstone and tuff with subordinate amounts of laminated cherts and carbonaceous siltstones. Minor quartz nodules and iron rich sedimentary sequences are additionally recognised (Ahmad & Hollis, 2013).

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Numerous semi-conformable sills of pre-orogenic Zamu Dolerite intrude the Koolpin Formation and the Gerowie Tuff and vary in thickness from several meters to a few hundred meters. The Mount Bonnie Formation is the uppermost unit of the South Alligator group and consists of greywacke, carbonaceous siltstone, chert, tuff and ironstone and with a variable thickness between 150 and 400 m thick (Ahmad & Hollis, 2013).

The Burrell Creek Formation is the lowermost sequence within the Finnis River Group and is comprised of a thick (<3000 m) sequence of turbiditic sediments including greywackes, siltstones and mudstones (Snowden Report).

The Tollis Formation is the youngest member of the Finnis River Group and is host to several gold deposits including Maud Creek, Mount Todd, and Quigleys deposits (north, south, extended) (Ahmad & Hollis, 2013). This unit has limited aerial extent and consists of a succession of interbedded mudstone, slate, metagreywacke and minor felsic volcaniclastic shale that was conformably overlies the Burrell Creek Formation. This unit has previously been attributed to the El Sherana Group, however more recent interpretations of have placed this unit within the Finnis River Group (Ahmad & Hollis, 2013).

7.2

Property Geology

The Maud Creek gold field lies approximately 20km to the east of Katherine and lies within the south-eastern part of the Pine Creek Geosyncline (Figure 7-3). The Maud Creek goldfield hosts the historic Maud Creek Mine and the Maud Creek deposit (historically known as the Gold Creek deposit) (AngloGold Report, 2000). Proterozoic rock units in the Maud Creek area comprise the Tollis Formation, Maud Dolerite, Dorothy Volcanics (formerly Dorothy Volcanic Member), Edith River Volcanics, and Kombolgie Formation.

The Maud Creek deposit is hosted within the Tollis Formation which outcrops in the centre of the northwest of the Maud Creek area. This unit is typified by thin to thick beds of alternating greywacke and mudstone, minor conglomerate, altered mafic to intermediate volcanic rocks and banded ironstone. The Dorothy Volcanics consist of basaltic lava, pyroclastic rocks, tuffaceous sediments and sills and locally lie in faulted contact with the Tollis Formation (Ahmad & Hollis, 2013). To the east of the Maud Creek Deposit, the Muad Dolerite intrudes the Tollis Formation and outcrops as irregular bodies of up to 200 m in width (Snowden Report, 2008).

In the northern portion of the Maud Creek area, felsic volcanics of the Edith River Group are unconformably overlain by the fluvial sediments of the Kombolgie Formation. In the south and west the Tollis Formation is masked by Cambrian Antrim Plateau Volcanics and Tindals Limestone.

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Figure 7-3:      Location Map of the Maud Creek Deposit

7.2.1

Property Mineralization

The Maud Creek deposit is hosted within the Tollis Formation of the Finnis River Group. Mineralization is associated with a north-south trending Gold Creek Fault Zone (GCFZ) that forms the contact between mafic tuffs of the Dorothy Volcanics to the east and sedimentary rocks of the Tollis Formation to the west. The GCFZ and primary Maud Creek mineralization dips steeply to the east (65-75 ). The GCFZ is characterized by intense deformed and brecciated to cataclastic zone up to 10 to 15 m width (AngloGold Report, 2000). The GCFZ and Maud Creek mineralization is associated with stockworks and massive quartz veining, silica flooding and brecciation as well as intense graphitic and chloritic alteration (Ahmad & Hollis, 2013). Additional alteration recognised includes silica, carbonate, fuchsite and haematite. The contact zone orebody geometry has been defined as lenticular in shape with a steep plunge (70-80) to the south-east. This principal mineralized zone extends approximately 250 m north-south, and ranges in width from several meters to up to 50m width. The deposit remains open at depth (Snowden Report, 2008). Mineralization is recognised to extend beyond the contact vein lodes, with dispersion up to 25 m into the hangingwall tuff and 5 m into the footwall sediments. Outside of the primary vein orebody minor hangingwall microbreccia zones are recognised predominantly occurring proximal to minor faulting parallel to the GCFZ (AngloGold Report, 2000).

Away from the main contact fault zone, gold is recognised within a subvertical shear zone which lies proximal to the contact of the Maud Dolerite (Snowden Report, 2008). Mineralization is less continuous in this zone, with the absence of any major vein lode systems evident.

Gold occurs within the deposit as both free gold and as refractory gold in pyrite and arsenopyrite (Snowden Report, 2008). Sulphides can constitute up to 5% of the orebody with pyrite and arsenopyrite and gersdorffite recognised. These sulphides form as disseminations as well as massive intervals containing up to 50% pyrite (Ahmad & Hollis, 2013). Quartz makes up the remaining gangue mineral of the orebody assemblage.

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Ahmad & Hollis, 2013, Pine Creek Orogen: Ahmed M and Munsen TJ (compilers). Geology and mineral resources of the Northern Territory, Northern Territory Geological Survey, Special Publication 5.

7.3

Deposit Mineralization

The Maud Creek deposit is hosted within the Tollis Formation of the Finnis River Group. Mineralization is associated with a north-south trending Gold Creek Fault Zone (GCFZ) that forms the contact between mafic tuffs of the Dorothy Volcanics to the east and sedimentary rocks of the Tollis Formation to the west. The GCFZ and primary Maud Creek mineralization dips steeply to the east (65-75°). The GCFZ is characterized by intense deformed and brecciated to cataclastic zone up to 10 to 15 m width (AngloGold Report, 2000). The GCFZ and Maud Creek mineralization is associated with stockworks and massive quartz veining, silica flooding and brecciation as well as intense graphitic and chloritic alteration (Ahmad & Hollis, 2013). Additional alteration recognised includes silica, carbonate, fuchsite and haematite. The contact zone orebody geometry has been defined as lenticular in shape with a steep plunge (70-80°) to the south-east. This principal mineralized zone extends approximately 250 m north-south, and ranges in width from several meters to up to 50 m width. The deposit remains open at depth (Snowden Report, 2008).

Mineralization is recognised to extend beyond the contact vein lodes, with dispersion up to 25 m into the hangingwall tuff and 5 m into the footwall sediments. Outside of the primary vein orebody minor hangingwall microbreccia zones are recognised predominantly occurring proximal to minor faulting parallel to the GCFZ (AngloGold Report, 2000).

Away from the main contact fault zone, gold is recognised within a subvertical shear zone which lies proximal to the contact of the Maud Dolerite (Snowden Report, 2008). Mineralization is less continuous in this zone, with the absence of any major vein lode systems evident.

Gold occurs within the deposit as both free gold and as refractory gold in pyrite and arsenopyrite (Snowden Report, 2008). Sulphides can constitute up to 5% of the orebody with pyrite and arsenopyrite and gersdorffite recognised. These sulphides form as disseminations as well as massive intervals containing up to 50% pyrite (Ahmad & Hollis, 2013). Quartz makes up the remaining gangue mineral of the orebody assemblage.

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8

Deposit Types

The contribution from gold deposits in Proterozoic sedimentary basins to total gold production has increased markedly over the past two decades, both globally and within Proterozoic basins in Australia. Consequently, many Proterozoic basins are now considered high priority exploration targets.

8.1

Deposit Models

A variety of genetic models, ranging from magmatic through hydrothermal to syngenetic, have been postulated in the past for the formation of gold deposits in the Pine Creek Geosyncline (Figure 8-1). Gold and base metal mineralization in the Pine Creek Geosyncline is commonly associated with granite intrusions and have often been classified as high temperature contact aureole deposits. A secondary host rock control has also been suggested due to the association of gold mineralization with carbonaceous metasedimentary rocks.

However, much of the gold mineralization occurred after the main intrusive event, the intrusion of the Cullen Batholith, and the relationship of gold mineralization and carbonaceous rocks is not the most important control on mineralization. More recently, authors have argued that gold mineralization is structurally controlled; occurring in brittle-ductile structures at the greenschist-amphibole facies boundary and hence has an epigenetic origin (Partington & McNaughton, 1997).

In places, e.g. the Cosmo-Howley area, duplex thrust folds with buckle folding or basin and dome structures appear to be more significantly mineralized. The presence of shear systems linking anticlines higher in the sequence also appears to have provided the ideal fluid focusing mechanisms to localize gold-bearing fluids.

Accepting that gold deposits of the Northern Territory have a structurally controlled mesothermal setting, then on the basis of host rock and mineral association they can be divided into seven types:

  Gold-quartz veins, lodes, sheeted veins, stockworks, saddle reefs (Pine Creek Orogen);
     
  Gold-ironstone bodies (Tennant Inlier);
     
  Gold in iron rich sediments (Pine Creek Orogen, Tanami);
     
  Polymetallic deposits (Iron Blow, Mt Bonnie);
     
  Gold-PGE deposits (South Alligator River area);
     
  Uranium-gold deposits (Pine Creek Orogen, Murphy Inlier);
     
  Placer deposits; and
     
  Over half of the gold occurrences are gold-quartz vein deposits.

Native gold is the main ore mineral and is commonly present as micron sized grains; coarse nuggets are rare.

Gold is commonly associated with pyrite, arsenopyrite and pyrrhotite and in places with minor base metal sulphides. Quartz, chlorite, sericite and carbonates are the common gangue minerals in the gold quartz deposits.

All gold deposits in the Northern Territory show some structural control at the regional and deposit scales, with most deposits within the Pine Creek Orogen trending northwest-southeast. Base metal veins in the Pine Creek Orogen strike significantly differently than the gold veins, suggesting different discrete mineralizing events. They are interpreted to be syngenetic.

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Most deposits show a preference for competency contrast situations in dilatant or low pressure zones, such as anticlinal crests, recurrent shear zones and necking zones. Gold mineralization is invariably late, occurring after orogenic events.

Common factors for most gold deposits include:

 

Gold deposits are nearly all in low grade, sub greenschist to lower greenschist facies regionally metamorphosed sediments (commonly greywacke-siltstone-shale);

   
 

Anticlinal hinges and shear zones are generally the most favourable loci;

   

 

Subsequent to regional metamorphism and deformation, the metasediments were intruded by I- Type granite and the gold mineralization are within the contact metamorphic aureole;

   

 

Fluid inclusion data suggest the involvement of moderate to high salinity fluids in temperature range from 200 300°C; and

   
 

Stable isotope data suggest a magmatic/metamorphic origin of these fluids.

Five main types of mineralization have previously been recognized within the Pine Creek Orogen. These include:

Sheeted and stockwork quartz vein systems located along major anticlinal hinges in the Mount Bonnie and Burrell Creek Formations and to a lesser extent, the Gerowie Tuff. Mineralization is hosted by carbonaceous or sulphidic host rocks (Woolwonga) or along zones of competency contrast between greywacke and shale (Enterprise, Union Reefs, Goodall, Alligator, Faded Lily, Howley, Big Howley, Yam Creek and Fountain Head) or dolerite (Bridge Creek). Axial planar quartz veins have been identified in some deposits (Enterprise and Woolwonga). Stratabound quartz reefs occur in most of these deposits, and may develop into saddle reefs along fold hinge zones (Enterprise, Union Reefs and Fountain Head);

     

Sediment-hosted stratiform gold mineralization and quartz-sulphide- vein-hosted stratabound gold mineralization in cherty ironstone and carbonaceous mudstones of the Koolpin Formation (Tom’s Gully, Cosmo Howley, Golden Dyke and Rising Tide) or the Gerowie Tuff (Brocks Creek);

   
 

Stratiform, massive to banded, sulphide- silicate-carbonate mineralization in the Mount Bonnie Formation (Mt Bonnie and Iron Blow);

   

Sediment-hosted stratiform and stratabound gold mineralization in cherty, dolomitic and sulphidic shales of the Mount Bonnie Formation, with sheeted quartz-sulphide veins (Rustler’s Roost); and

   

Sheeted or stockwork quartz-feldspar-sulphide veins hosted by Zamu/Maud Creek Dolerite sills (Maud Creek, Howley, Howley South, Bridge Creek and Kazi). Most gold mineralization in the Pine Creek Orogen occurs within the South Alligator Group, especially above the Middle Koolpin Formation, and in the lower parts of the Burrell Creek Formation. At Maud Creek gold mineralization is hosted by the Tollis Formation that represents the uppermost unit of the El Sherana Group and unconformably overlies the Burrell Creek Formation. Most of the fold- associated deposits were probably formed during intrusion of granitoids such as the synorogenic Cullen Batholith and the Burnside- Granite.

The most important regional scale exploration vectors to the orogenic style of gold mineralization are:

The position of the biotite isograd in the contact metamorphic aureole of the Cullen granitoids. The biotite isograd needs to be mapped out carefully in areas of exploration interest and exploration focused on the biotite- albite epidote contact-metamorphic zone.


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NNW-NW oriented anticlinal axes appear to be the most productive. However, exploration cannot be totally restricted to anticlines in this orientation, as other anticlines or even synclines may be mineralized.

   
 

Strongly interbedded and contrasting rock types (e.g., greywacke-siltstone) particularly in the upper parts of the stratigraphy in the Mount Bonnie and Burrell Creek Formations in particular.

   

Carbonaceous or iron- rich lithologies in proximity to indications of gold mineralization. Such lithologies and any veins within them need to be mapped out carefully to help locate potential trap sites for economic gold mineralization.

Figure 8-1:      Structural – stratigraphic model for Newmarket Gold deposits

Pine Creek Orogen (Sener, 2004)

8.2

Structural Models

Assuming that the majority of gold deposits within the Pine Creek Orogen are structurally controlled and mesothermal/orogenic (cf. Groves et al. 1998) in origin, it is likely that the known gold deposits are associated with regional shear zones and fault systems that were formed during orogenesis. By analyzing maps displaying total magnetic intensity (TMI) data, a number of continuous, NNW- trending first-order faults can be defined within the sedimentary-dominated rock sequences of the Burnside tenement area (Figure 8-2).

The majority of known gold deposits within the tenement area are spatially associated with the first-order, NN-trending shear-zones. It is therefore likely that these first order shear zones acted as conduits for epigenetic gold bearing fluids during/after orogenesis and they control the distribution of gold mineralization known in the tenement area. Additional factors such as the presence of the South Alligator Group, proximal antiformal hinges (e.g., Cosmo-Howley) or converging secondary shear zones (e.g. Crosscourse) would also play an important role in localizing gold mineralization.

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The major shear zones are separated by rock sequences that regularly preserve NNW-trending, doubly-plunging- anti-formal hinges with no clear evidence for strike-slip deformation along these NNW trending structures. South of the Burnside granite area, a series of NE-trending shear zones and faults have also been defined (Figure 8-2). Based on preserved asymmetries of rock sequences either side of these NE-trending faults, dextral-dominated strike-slip deformation possibly occurred along these relatively later structures.

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Figure 8-2:      Pine Creek Regional structural interpretation

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9

Exploration


9.1

VTEM Airborne Survey

A total of 590 line kilometers of VTEM survey were flown in the Maud Creek area (Figure 9-1) in 2011 covering an - area of approximately 300 km2. Line spacing was usually 200 meters in the northern area (NW SE direction) and at 400 meter line spacing (NE-SW direction) in the southern part of the area (Table 9-1). Southern Geoscience completed an initial interpretation and report of the survey data and Newmarket Gold geologists incorporating geology and available geochemistry reinterpreted this report (Figure 9-2).

Sixteen strong conductive targets and two moderately conductive targets were identified in the Maud Creek block. It should be noted that the strength and quality of conductors is significantly reduced from those defined at Burnside and Moline. The flat lying limestones and basalts that mask the underlying Proterozoic rocks likely play a significant role in the anomaly definition.

Figure 9-1:      Maud Creek Property with VTEM flight lines

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Figure 9-2:      Strong VTEM Conductors on a 35 hertz conductor base map

The 25 m line spacing airborne magnetometer/radiometric survey flown by Kalmet in 1997 clearly defines the extent of the younger Atrium basalts (Figure 9-3) . Their signature produces a distinct noisy mottled effect on various manipulations of the magnetic data. The Maud Dolerites appear to have distinct magnetic anomalies along the margins of the body but the centre of the body appears to be magnetically quiet. It is quite possibly a differentiated intrusive.

Late, generally northeast trending dolerite dykes are readily apparent on the aeromagnetic image. Disruptions can be seen in these dykes indicating that later faults have created minor offsets.

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Figure 9-3:      Merged 1997 and 2012 Aeromagnetic data with strong VTEM conductors (Card, 2012)

Table 9-1:      VTEM conductor prioritization – Maud Creek survey area

Conductor
#
SG
Priority
Priority Length (m) Surface
Work
Host
Fm
Comments
MCLT005 1 1 400-600 No lmstn SW dipping, near a lmstn o/c, quartz veining noted in the area, previously Maxwell modeled, don’t have the model. Low ground between o/c. Deep seated. NW strike, no distinct magnetic signature . Weak Pb and Au in soil anomalies. Geochemistry likely masked by lmstn. May be on a NNW trending lineament
MCLT003 2 2 200 No lmstn Embayment in lmstn? Flanking linear
magnetic anomaly to the north. Dyke?
Right on the edge of
High res mag survey area. Soil color
anomaly on Google? Possibly on a
north south lineament. Isolated
conductor
MCLT007 3 2 200 No lmstn Just north of a stream, soil color
anomaly, weak magnetic anomaly
indicated on high resolution magnetic.
Distinct direct magnetic anomaly.
Work this first. Deep seated. Need to
model
MCLT008 3 3 200 No lmstn Small magnetic anomaly, near o/c

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Conductor # SG
Priority
Priority Length
(m)
Surface
Work
Host
Fm
Comments
MCLT009 3 200 No lmstn Small magnetic anomaly, larger anomaly to the NW, covered by high resolution magnetic. Near o/c to the north. Deep seated >200 m?
MCLT010 3 3 200 No lmstn Small magnetic response, on the
exploration tenement. o/c to the north
soil color anomaly to the north
MCLT011 3 3 200 No   Small magnetic response, on the
exploration tenement. o/c area
MCLT012 3 3 200 No lmstn Small coincident magnetic anomaly,
on exploration tenement. o/c in area
MCLT014 3 3 200 No lmstn Small coincident magnetic anomaly,
on exploration tenement. North of o/c
MCLT006 3 3 200 No lmstn Close to a NE trending lineament
located to the SE. SE dip. Pb in soil
anomaly
MCLT016 3 3 600- 800 No lmstn SG says probably culture. On road
but migrates to the south. South dip.
Fence line but no different from other
fences. Walked the road. Lots of road
metal contamination from mine area.
Limestone cover. o/c to south.
Magnetic anomaly coincident on
eastern line. Noisy data.
MCLT015 4 3 200 No lmstn Very similar to 007, 008, 009, 010,
011, 012, 014. Determine the cause
of 007 first. Very small target . Non-
magnetic, o/c area. Edge of soil
sample survey
MCLT002 4 3 800-1000 No lmstn Possibly thick conductive overburden.
On a north-south lineament. No
distinct magnetic signature. NE strike.
On the NW trending base metal trend
but striking the wrong direction
MCLT001 4 2 800-1000 No lmstn Possibly conductive overburden. No
distinct magnetic signature. On
exploration tenement. Changes strike
direction from ENE to NS. On base
metal anomaly but strikes the wrong
way. Low ground. Possible o/c area .
Possible north-south lineament with
Chessmen and Red Queen to the
north
MCLT004 4 1 200 No lmstn At the south end of a linear magnetic
feature that hosts Maud Creek to the
north. Also on a NW trending
magnetic feature that may have an
association with the base metal in soil
anomaly. Watch out for buffalo.
Increased vegetation in the area.
Slight soil color anomaly


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Conductor
#
SG
Priority
Priority Length (m) Surface
Work
Host
Fm
Comments
MCLT013 4 2 200 No ? Very small and narrow, very isolated
response. Weak magnetic anomaly.
Possible east-west lineament.
Outside soil survey area.
MCLT017 4 1 600-800 Yes Maud
dolerite
At old Maud Creek workings. Change
in strike direction to conductor. ENE
to NNE. Associated with magnetic
anomaly.
Northern response outside property
boundary. Highly conductive regolith.
Alluvials in area
MCMT002 NP 1 200 No lmstn SE of a large magnetic anomaly
interpreted to be in a regional NE
trending structure . Appears to be in a
nature park, karst caves.
Interesting target but deep seated and
too many social issues. Ignore

9.2

Stream sediment survey

In the mid 1990’s a fairly extensive stream sediment survey (approximately 173 samples) was carried out in the Maud Creek area. Samples were taken along various drainage systems at sample spacing of 500 meters or closer. It is believed that a BLEG analysis was carried out with elements such as silver, copper, lead, zinc, bismuth, molybdenum and antimony also determined (Table 9-2).

In 2012 it was decided to stream sediment sample the 600 km2 area of EL 28902 in order to quickly determine its mineral potential. Arnhem Exploration Pty Ltd were contracted to carry out the survey.

A total of 164 stream sediment samples were collected along streams at approximately 1 km spacing. Figure 9-4 displays the sample distribution over the- tenement area. Note that there is some duplication/overlap in the areas sampled with the 1996 97 survey. Samples were collected from several sites at any particular collection point in order to alleviate point anomaly sources. Each site was photographed. Samples were sieved to -75 microns in the field. A total of 9 duplicate samples were taken for QAQC purposes.

A visual examination of the QAQC sample duplicates indicates that the gold values exhibit good repeatability while the other elements returned acceptable levels of repeatability. It should be noted that one sample exhibited marginal repeatability in lead and copper.

A correlation matrix of the 2012 survey data population indicates that there are weak correlations between gold and arsenic (0.38), bismuth (0.36), nickel (0.42) and tin (0.37) . There are significantly higher correlations with silver and copper (0.87), lead (0.88), antimony (0.92), arsenic (0.61) and barium (0.66) . It is suspected that several mineral deposit types are present in the area and that the correlation relationships are reflecting those differing styles.

A brief statistical look at the results for selected elements from the 1996-97 survey and the 2012 survey in Table 9-2 reveals the following:

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Table 9-2:      Comparative statistics between 1997 and 2012 soil survey results

Year Mean Ag Grade
(ppm)
Mean As Grade
(ppm)
Mean Gold grade
(ppb)
Mean Cu Grade
(ppm)
Mean Pb Grade
(ppm)
1997 0.021 4.1 0.699 33.1 55.2
2012 0.026 2.0 0.695 19.6 26.4

There is a good agreement between gold and silver values for both surveys. There is a wide divergence for copper, lead and arsenic. The base metal soil anomaly located south of the Maud Creek deposit skews the 1997 survey results. If both surveys were to be merged then the latter 3 elements would have to be normalized before they could be plotted and displayed effectively.

Underlying geology plays a distinct role when it comes to interpreting the stream sediment data. Largely Cenozoic and Mesozoic materials that overlay Proterozoic lithologies and/or Cambrian age sediments or volcanics underlie the significant land area to the east of the Maud Creek deposit. The masking effect of these younger sediments has likely diluted any anomalous effects from underlying Proterozoic rocks so subtle anomalies need to be field checked. It is conceivable that there are windows through the younger sediments that expose the Proterozoic as they do at the Copper Breccia occurrence that are located at the east end of the original Maud Creek tenements.

The gold results generated a number of anomalies. Most obvious is the cluster of anomalous samples in the Maud Creek deposit area and that area, which is underlain by Maud Dolerite. The 1997 survey did not produce an extensive gold anomaly in this area. It is suspected that sampling the finer sediments in 2012 produced more consistent results.

The 1997 survey produced a strong and extensive Au anomaly in the Red Queen/Chessmen area (northwest part of the property). This area was not re sampled in 2012 but the area to the southwest was and the gold anomaly appears to extend into this area.

In the area that the geology map indicates (Figure 9-4) there is extensive cover a number of gold anomalies occur. These all have anomalous multi-element associations and all need to be ground checked.

Anomaly 1 is a cluster of two samples that are anomalous in gold, iron, chromium, arsenic, lead, molybdenum, tin and uranium. A preliminary interpretation of regional aeromagnetic data indicates that this anomaly may be situated within the southern limits of the Maud Dolerite unit (as defined by magnetics). SRK’s structural interpretation of the region carried out for Kilkenny Gold in 1998 indicates several large ENE trending structures that cut through this target area.

Anomaly 2 is one sample that is highly anomalous in gold (9.7 ppb) as well as molybdenum, bismuth and arsenic. The area should be resampled with either 3 or 4 stream sediments or a small soil grid established to cover the upstream area. The anomaly area does not exhibit any distinct magnetic signature

Anomaly 3 is a cluster of 4 samples that is anomalous in gold, copper, bismuth, arsenic, lead, chromium, uranium, tin, silver and barium. It is interpreted to be in the southern extension of Maud Dolerite. SRK’s structural interpretation of the region carried out for Kilkenny Gold in 1998 indicates several large ENE trending structures that cut through this target area. Anomaly 4 is located at the very northeast part of the survey area and is anomalous in gold, silver, iron, bismuth, arsenic, lead, chromium, uranium, tin and silver. The area is likely underlain by Tollis Fm. The regional magnetic data indicates it is proximal to a northeast trending dolerite dyke. Northwest trending structures can be interpreted. The gold in stream sediment anomaly appears to occur at the intersection of two major structures according to the SRK interpretation.

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Anomalies 5, 6, 7, 8, 9 and 10 are one and two point anomalies. They all need to be ground checked. The extensive base metal anomaly defined from the Kalmet 1997 stream sediment and soil surveys is clearly defined on the normalized copper and lead stream sediment map. - It would appear that is may extend over a distance of 30 35 km. It obviously weakens to the east and west but this may be a function of increased thicknesses of younger cover. The base metal anomalous area is also defined with elevated values in barium, cadmium, chromium, iron, potassium, magnesium, manganese, nickel, tin and zinc.

The base metal anomaly in stream sediments and soils doesn’t appear to be directly related to any individual VTEM anomaly and there does not appear to be a correlation to any particular magnetic response. The area of the soil anomaly is largely underlain by young Cambrian aged volcanics, although one area is underlain by Maud Dolerite. If the geochemical response comes from the Proterozoic rocks then there is some sort of mechanism allowing it to percolate through the younger volcanics. Looking at the airborne magnetic and radiometric data it would appear that the base metal anomaly occurs at or very close to the younger volcanic – limestone contact. The geochemical anomaly can be traced for >10 km in a generally east west direction.

Figure 9-4:      Maud Creek Area – 2012 stream sediment survey – selected gold anomalous areas

9.3

Soil sampling surveys

In 1997 Kalmet carried out an extensive soil survey over the Maud Creek area on lines 400 meters apart and samples taken at 25 meter intervals but composited to 50 meter intervals (Figure 9-5). Samples were analyzed for gold, silver, copper, lead, zinc and antimony. A number of anomalous areas were defined but the coarse line and sample spacing didn't allow for accurate directional interpretations.

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Early in 2012 fourteen target areas were identified and given a letter designation. These did not include the VTEM anomaly reprioritization. The objective was to determine through rock chips, soil sampling and mapping the location of possible future exploration drillholes.

Prioritized Targets:

Anomaly A UTM 221500E 8404400N

Two gold in soil anomalies. The eastern one is about 1200 meters long with coincident antimony while the western one is 400 meters long with coincident antimony over 800 meters. Both have elevated arsenic values. It is interpreted that these are not overbank stream sediment related. No significant base metal anomalies.

   
 

No significant VTEM anomaly associated with the geochemical anomalies.

   

 

 

The magnetic data indicates a possible NE trending dyke (late dolerite).

   

 

 

Possible K anomaly. Distinct uranium/thorium anomaly.

   

 

 

Possible north-south structure interpreted from Total Count and SRK structural data.

   

 

 

Chessmen and Red Queen occurrences to the north, possibly on the same north- south structure.

   
 

Likely underlain by Antrim volcanics, basalt

   

 

 

Nothing distinct on Google image.

   

 

 

May be the source area for the distinct and widespread gold in stream sediment anomaly in the area. -

Anomaly B UTM 221500E 8406400N

 

Gold in soil anomaly over 1200 m running north south with coincident antimony and arsenic over a 400 m length. There are streams to the north and south so there is a possibility that there may be contamination from overbank stream sediments.

   

 

 

SRK structural geology indicates Tollis limestones and a north south structure, which may connect up with Target A.

   

 

 

Very weak VTEM channel 25 response. Possible NNW trending feature.

   

 

 

Magnetics indicate an interpreted NE trending dolerite dyke. Possible north south folds.

   

 

 

Small but distinct uranium/thorium anomaly.

   

 

Nothing distinct on Google image.

   

 

 

Need a compilation- of past work at Red Queen and Chessmen prospects.

Anomaly C UTM 221000E 8407400N

  The south part of the anomalous area has a stream close by so there may be overbank stream sediment contamination. One anomalous gold in soil sample to the west. There are elevated arsenic and antimony results to the north. Possible overbank contamination.
     
  No distinct VTEM response.
     
  Possibly the NW end of a NNW trending structure.
     
  Magnetics indicates a NE trending late dolerite dyke.
     
  There is a north south trending structure to the east.
     
  Distinct Total Count (TC) anomaly with some potassium contributing.
     
  Nothing distinct on Google image. Obvious stream.

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Anomaly D - UTM 222200E 8400000N

  Two anomalous Au in soil results. Soils here were not analyzed for arsenic, antimony or base metals.
   
  VTEM indicates a possible north-south structure on channel 25.
     
  Likely underlain by Tollis limestone.
     
  No distinct magnetic features.
     
  Radiometrics indicates that the target is on the south edge of a major NW trending TC anomaly.
     
  SRK map indicates possible volcanics at the edge of limestones.
     
  Nothing distinct on Google image.

Anomaly E UTM 224000E 8401000N

 

Gold in soil anomaly over 400 800 meters running north south. Elevated arsenic and antimony. On the edge of base metal anomaly which is part of the major NW trending response.

     
  Kangaroo Flats prospect to the north. Don’t know anything about this prospect.
     
  Distinct magnetic anomaly, unknown cause.
     
  No distinct radiometric anomalies.
     
  SRK map indicates east-west structures.
     
  Underlain by Antrim volcanics.
     
  Nothing -distinct on Google image.

Anomaly F UTM 224300E 8402400N

  Gold anomaly in soils 400 800 m north south. No significant arsenic, antimony or base metals. Possibly an overbank stream sediment anomaly.
     
  Kangaroo Flats prospect is 500 m to the west.
     
  No VTEM response.
     
  Radiometrics: possible NW structure and possible north-south structure. No distinct anomalies.
     
  Underlain by Antrim volcanics
     
  Low priority- target.

Anomaly G UTM 225000E 8401000N

 

Two point gold in soil anomaly. No arsenic, antimony or base metals. sediment.

   
 

Near Curlies area to the SE.

   

 

 

Underlain by Antrim volcanics.

   

 

 

Distinct potassium anomaly.

   

 

 

SRK map indicates an east-west structure.

   

 

 

No distinct magnetic features.


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Anomaly H - UTM 225400E 8401600N

  Maud Creek deposit area anomaly.-2000
     
  Gold in soil anomaly extends 1600 m north-south.
     
  Coincident arsenic and antimony with weak molybdenum and silver. No elevated base metals. Maybe contaminated from overbank stream sediments.
   
  VTEM weak channel 25 response trending north-south.
     
  Sediment-tuff contact.
     
  Antrim volcanics to the immediate west and south.
     
  Radiometrics indicates a distinct north-south low as well as NE and NW lineaments.
     
  Magnetic low feature. West of Maud Dolerite.

Anomaly I UTM 226000E 8402000N

  Gold in soil anomaly 600 m east west x 500 m north south. Coincident arsenic and antimony anomalies. Quite possibly overbank sediments.
     
  Chlorite Hills to the NE.
     
  Surprise area between targets H and I.
     
  Underlain by Maud Dolerite.
     
  North of NE trending late dolerite dyke.
     
  Weak potassium anomaly – probably Maud Dolerite signature.
     
  Close to-Maud Creek and old workings.

Anomaly J UTM 226700E 8401600N

  Gold in soil anomaly 400 800 meters north south trending. Coincident arsenic anomaly.
     
  Antimony coincident over 1200-1600 m.
     
  Jibaroo prospect.
     
  Maud Creek to the north. Possible overbank stream sediment contamination.
     
  No VTEM response.
     
  K/Th anomaly possibly caused by dolerite. NE structure indicated.
     
  North-south structure to the east.
     
  Good target area. Maud look-a-like target – geophysically.
     
  East-west structure.

Anomaly K - UTM 227600E 8401600N

  Gold in soil low order anomaly with coincident anomalous antimony and strong arsenic. Fairly widespread so may be overbank contamination. Near a stream.
     
  North – south structure to the east.
     
  No distinct radiometric anomalies.
     
  No distinct VTEM anomalies.
     
  Magnetic low, magnetic destruction?
     
  Underlain by Maud Dolerite.

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Anomaly L - UTM 228200E 8401600N

  Runways prospect.
     
  No streams in the immediate area.
     
  Gold in soil anomaly over 400-800 meters with coincident arsenic and antimony.
     
  Base metals are very low order.
     
  Magnetic anomaly caused by Maud Dolerite.
     
  Potassium/thorium anomaly interpreted to be caused by Maud Dolerite.
     
  Bracketed by north-south structures.
     
  Good structural target. East margin of Maud Dolerite.

Anomaly M UTM 226400E 8399600N

Small isolated gold in soil anomaly with coincident arsenic and weak antimony. Right on several streams so quite possibly overbank contamination.
     
  Drovers prospect,-south anomaly at south end of prospect.
     
  Distinct north structure, good target area.
     
  Intersecting NE and NW trending structures.
     
  Potassium/thorium anomaly caused by Maud Dolerite.
     
  Linear magnetic low trending NNE.

Anomaly N - UTM 227500E 8399600N

  Single point anomaly with coincident antimony and arsenic anomaly to the west.
     
  No proximal streams.
     
  No distinct base metal anomalies.
     
  Potassium/thorium interpreted north-south structure.
     
  Maud Dolerite overlain by Antrim volcanics.
     
  Possible NE trending magnetic anomaly interpreted to be caused by a dolerite dyke.
     
  No distinct VTEM anomaly.

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Figure 9-5:      Location of soil anomalies requiring further work

A limited field program carried out later in 2012, designed to investigate the 14 target areas, defined the following:

Anomalies A and B

Anomalies A and B were originally contoured with a north-south bias, however, field examination indicated that the anomalous gold-in-soil geochemical values should be- contoured in a north- northeast direction (Figure 9-6). When this was done two narrow gold-in soil geochemical trends- with coincident arsenic and antimony anomalous values were defined, coincident with two 020 025° trending felsic dykes and/or silicification/brecciation.

Additional detailed soil sampling was recommended and ultimately carried out in 2012.

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Figure 9-6:     Red Queen – Chessman Area – Soil anomaly designation

The Chessman and Red Queen occurrences both occur on the northernmost trend (B) (Figure 9-6). The Chessman occurrence, located at UTM 8406600N; 221460E, is located south of a 58.6 ppb gold soil anomaly, and the Red Queen occurrence, located at UTM 8406230N; 221330E, south of a 11.3 ppb gold soil anomaly (both anomalies defined by past operators work). This siliceous zone on which both the Chessman and Red Queen occurrences are located continues southward for at least- an additional two kilometers, siliceous brecciated material located at anomalous gold in-soil sample sites which returned 33.9 ppb and adjacent to soil sample site which returned 11.6 ppb.

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Anomaly H

Anomaly H was only briefly visited as it defines the known mineralization at the Main Maud Creek- deposit (Figure 9-8). The strongest gold-in soil analytical results occur exactly on what is now the Maud Creek open pit mined in 2000. The northern and southern extensions of the gold-in-soil anomaly defined the mineralized trend of the Maud Creek deposit.

Chlorite Hills (Northern part of Anomaly I)

The strong gold in-soil anomalous results at the north end of Anomaly I are probably defining the gold mineralization associated with the Chlorite Hills veining (Figure 9-7). Two fences of drillhole collars were located along a 25 x 25 meter grid spacing. Not all of these drillhole collars are in the Newmarket Gold database, however, those that are were drilled at a 270° azimuth, parallel to the observed quartz veining. Additional drilling completed by Kilkenny Gold was oriented in multiple directions, 320°, 140°, 180°, 90° and 270° azimuths. Anomalous gold, ranging from 0.54 g/t over 4 meters to 2.25 g/t over 4 meters, was intersected. The assembly of the drilling database for the Maud Creek Property area is still in progress.-

There were no surface indications of a north south structure as indicated by Hill 50 NL in an internal memo by Bob Watchorn, however, the drilling completed by Kilkenny with 090° and 270° azimuths did intersect- anomalous gold (up to 2.25 g/t over 4 meters), possibly indicating the existence of north south mineralized structures in the area.

A traverse was conducted between the Maud Creek open pit to the Chlorite Hills pits to try to determine if the mineralized quartz veins at Chlorite Hills extended to the Maud Creek deposit, however, there were no outcrop exposures until the Maud Dolerite. The contact between the Tollis Formation and the Maud Dolerite occurs at a north south oriented creek approximately 400 meters east of the Maud Creek deposit. A second traverse was completed from the Chlorite Hills 040° trending pits ending exactly at the one drillhole completed at the Surprise area, possibly indicating a 040° structure. (RC hole MCT-013 completed by Newmarket Gold 2011, 270° Azimuth, -60° dip. No significant gold analytical results were returned. The entire hole is in dolerite). There was no evidence noted to support previous interpreted north-south structures in this area.

O’Shea’s to Anomaly J

The O’Shea’s occurrence is located at UTM 8401920N, 0226920E. An outcrop of very fine grained siliceous rock disseminated with very fine sulphides with brecciated angular clasts to 10 cm occurs at the occurrence within coarse grained diorite. The silicified zone trends 30° with several pits excavated along its trend. Quartz vein mapping by Kilkenny Gold define NE trending veins in this area with additional minor NW trending veins.-

The main workings strike north south with drilling completed east-west. A fair amount of historical drilling has been completed at the occurrence. RC drilling in 1998 targeted the NE trending shearing and associated workings. The mineralization intersected a 2 to 4 meter wide quartz/sulphide shear within the dolerite dipping 60° to 70° to the southeast. -

A traverse at 30° was completed directly to Anomaly J (Jilleroo occurrence) gold in-soil geochemical target arriving onto a series of shallow pits (Figure 9-7). These pits were investigating siliceous hematitic, highly altered rock and quartz veining and/or brecciation in hematitic diorite. Minor malachite was noted along fractures. Shearing and quartz veining were striking 135°. Two sub- parallel 135° trending shears were observed.

The Jilleroo occurrence may be occurring at the intersection of the continuation southward of the 30° O’Shea’s shear and the 135° structures.

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Figure 9-7:      Chlorite Hills and O’Shea’s soil anomaly designations - Anomaly G, south end of H and south end of I to O’Shea’s single point gold in-soil

Drawing a straight line lining up the eastern most anomalous gold-in-soil analytical result at Anomaly G (11 ppb) to the south end of Anomaly H (Maud Creek) (16 ppb) to Anomaly I (Surprise) (16.6 ppb) defines a 60° trending gold in soil anomaly (Figure 9-9). Extending this trend to the northeast the trend will pass immediately north of O’Shea’s and arrive at the Carpentania Pass target. The northeastern portion of this trend parallels a siliceous felsic dyke mapped in the area. This 60° trend has been noted numerous times throughout the property. The gold-in-soil analytical results may define a 60° crosscutting structure in the area or perhaps a fault that cuts the Maud Creek deposit at the southern end.

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There has been extensive drilling completed (both RC and RAB) over the northern and southern extensions of the Maud Creek deposit by past operators. This drilling covered Anomaly G and Anomaly H. Very little drilling has been completed over the area east of The Maud Creek Deposit.

Pussy Shear Zone

A brief visit was conducted at the southern end of the Pussy shear zone area (Figure- 9-8). Extremely hematitic gossanous material with a stockwork of cm wide quartz veinlets and sub crops of agglomerate, similar to that which occurs at the Maud Creek deposit were noted. - A postulated north south shear zone was not confirmed. Additional prospecting should be completed along this trend to confirm Hill 50 NL’s Internal Memorandum defining a north south structure. Several narrow 135° trending shears within diorite were noted. Several hand drawn sketches (completed by Kilkenny Gold) indicate en-echelon NW-SE trending structures within a north-south corridor. Grab samples collected by Kilkenny returned values between <0.01 to 0.18 g/t Au.

Figure 9-8:      Pussy Shear zone soil anomalies

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Kitten (Drivers) Anomaly

A brief visit was conducted at the Kittens (Drovers) Anomaly M area (Figure 9-9). A corridor of extensive quartz and quartz breccia- float trending 60° to 70° was located. Within this corridor several discontinuous siliceous sub crop ridges, massive to brecciated trending 80° were observed. In 1998 Kilkenny Gold drilled a fence of RC holes across the southern portion of this area (17 holes totalling 842 meters). Drilling did not intersect any significant gold values. Two different intrusives were intersected by the drilling (best gold value was 0.8 g/t over one meter interval along one of the intrusive contacts).

Figure 9-9:       Chlorite Hills and Kittens (Droves) soil anomaly areas

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Anomaly L

Gold in-soil Anomaly L occurs at the Runways target. The area is flat non-descriptive terrain with numerous float boulders of very fine grained dolerite within an area of reddish soil with patchy manganese coatings. There were no obvious indications as to the cause or source of the elevated gold-in-soil analytical results. Past operators completed two east-west lines of drillholes across this area, which extended across Anomaly K. A total of 44 RC holes (2,534 meters) were drilled. The drilling returned no significant gold results (best value 0.14 g/t Au over 4 meters). Buried intrusive and tuff were intersected. The NW trending magnetic anomaly in the area appears to correspond with these intrusives.

Anomaly E

Anomaly E (defined by two 7.6 ppb gold north-south correlated gold-in-soil analytical results) located on strike with VTEM anomaly MCLT_002 was prospected as part of the property examination. The area is flat, overlain with numerous boulders and rubble of quartzite, fine grained sandstone and chert. The overburden cover, as seen- in creek beds is at least one meter thick. There was no obvious cause for the elevated gold in-soil analytical results.

VTEM Anomaly MCLT_002

The area of VTEM conductor MCLT_002 is extremely flat. Minor outcrops of quartzite overlain by a thin cover of limestone were found. There was no obvious cause for the VTEM anomaly.

Subsequently, it was decided to soil sample two areas in detail, Chlorite Hills-O’Shea’s and the Red Queen/Chessman area. Samples were taken along lines 100 meters apart, oriented to cross cut regional structures and geology. Samples were taken 25 meters apart. The ionic leach method of analysis was selected so samples were taken at shallow depths of 10-20 cm and were sieved in the field to -75 microns.- ALS Chemex’s Au + ME-MS41 0.0001 - 0.1 Au by Aqua Regia with ICP-MS finish multi element package was the analytical method used.

Chlorite Hills Area

In the Chlorite Hills/O’Shea’s area a total of 596 soil samples were collected by Arnhem Exploration with assistance from Newmarket Gold field assistants. This includes 20 sample duplicates taken to monitor QAQC of the commercial lab. These were subsequently shipped to ALS Chemex facility in Darwin NT. - Instructions were to have the samples analyzed using Chemex’s ionic leach ME-MS23 multi element package. A visual inspection of the 20 QAQC duplicate samples indicates that the repeatability for the sample population is good with no obvious errors.

A correlation matrix of the entire population indicates strong correlation (0.77) between gold and copper and a weak to moderate correlation between gold and silver and arsenic (0.35, 0.32) . Table 9-3 displays a comparison of the mean for a few elements from the 2 sample areas at Maud Creek and the one VTEM target area east of Bon’s Rush in the Burnside area.

Table 9-3:      Comparative statistics for Maud Creek and Bons Rush Area – soil sample results

Area Mean Ag Grade
(ppb)
Mean As Grade
(ppb)
Mean Gold grade
(ppb)
Mean Cu Grade
(ppb)
Chessmen 2.67 6.5 1.13 1 270
Chlorite Hills 8.9 4.97 5.35 2 249
VTEM Anomaly Burnside Area 3.41 13.43 0.25 975

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The gold values in the Chlorite Hills area are significantly higher than those from the other two areas. The same applies for silver and copper (Figure 9-10). One explanation may be that the Chlorite Hills/O’Shea’s area has seen historic mining with some significant ground disturbance and this contamination may be partially responsible for the elevated values in multiple elements.

Nevertheless, the gold in soil values exhibit an interpreted east west trend in gold that is not obviously repeated in other elements. The O’Shea’s and Chlorite Hills areas stand out as being quite anomalous in gold and somewhat in copper. O’Shea’s is anomalous in silver and antimony. The west side of the grid displays anomalies in silver, calcium, iron, magnesium, nickel and anomalously low in zirconium, uranium and thorium. It is suspected that this area represents a contact with Maud Dolerite, which is interpreted from magnetic data to be a differentiated intrusive unit.

The eastern margin of the grid displays anomalism in calcium, magnesium and nickel. It is suspected that this exhibits a contact with a phase of the Maud Dolerite unit.

An anomalous area at the southeast end of the grid centred on 226900E 8401400N displays elevated values in a number of elements including antimony, thorium, titanium, uranium, zirconium, lithium, iron, arsenic, copper and weakly in Au. The area needs to be ground checked.

It is suspected that multi-element anomalism in the region of 226300E 8402200E may be due to overbank sediments associated with Maud Creek, which is situated immediately to the northeast. This needs to be ground checked.

Figure 9-10:      Au (ppb) for ionic leach soil results at Chlorite Hills, Maud Creek

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Chessman Area

In the Chessman/Red Queen area Arnhem Exploration collected a total of 1,942 samples with assistance from Newmarket Gold field assistants (Figure 9-11). This includes 61 sample duplicates taken to monitor QAQC of the commercial lab.

A visual inspection of the 61 QAQC duplicate samples indicates that the repeatability for the sample population is good for most elements with acceptable variability between duplicate pairs. For gold values four of the duplicate pairs exhibit a variability that is outside acceptable limits. The other elements return acceptable values for the most part.

A correlation matrix of the entire population indicates a fairly strong match of 0.67 between gold and silver. The gold in soil values presents a unique distribution that is not matched with any other element. In Figure 9-11 a NNE trending anomaly extends for 2.5 kilometers along the west side of the grid that would appear to extend beyond the grid to the NNW. A preliminary interpretation would indicate that the gold anomaly sub-parallels the west contact of a lithological unit that appears to be associated with a NNE trending syncline (graben?). In all likelihood its emplacement is structurally controlled. There is no distinct magnetic correlation with the gold anomaly although it would appear to cut through a strongly magnetic NE trending dolerite dyke located at the NW end of the grid. Another Au in soil anomaly at the north end of the grid needs to be investigated on the ground. A highly folded magnetic anomaly located at the south end of the grid can be traced for many kilometers to the southeast. Malachite staining was noted in the area. The magnetic anomaly is coincident with contorted anomalies in copper, silver, magnesium, barium, calcium and lead and possibly a weak gold response. It is markedly negative in molybdenum, barium, tin, antimony, uranium and iron. Conceivably, this anomalous situation correlates with other significant base metal anomalies defined at the southeast end of the Maud Creek property. Further field investigation is required.

A variety of other elements are obviously defining differing lithologies that define the synclinal structure. An example is iron that defines the east and west margins of the syncline. It, along with a number of other elements also defined some folded structures within the core of the syncline. Field investigation is required to link these anomalous situations to specific rock types. The centre of the syncline has its own distinct geochemical signature and the ionic leach data along with magnetic and radiometric information can be used with great effect to help with geological and structural interpretations.

Calcium and magnesium data can be used to interpret crosscutting (NW-SE) trending dykes. The VTEM data has defined a NW trending feature that correlates directly with a fence. A stronger- NW trending conductive feature at the north end of the soil grid is likely associated with a dyke like feature.

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Figure 9-11:      Ionic leach Au (ppb) in soil results, Chessman Area, Maud Creek

9.4

Rock Chip Sampling

During the reconnaissance prospecting of various soil anomalies a series of 58 rock samples were collected in the Red Queen/ Chessman and Chlorite Hills areas. Figure 9-12 displays the gold results.

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Figure 9-12:     Maud Creek area gold in 2012 rock chip sample results

Anomalous gold values also display elevated values in silver, arsenic, bismuth, copper, iridium, lead, antimony, tin, tungsten and zinc. There is a distinct phosphorus depletion associated with the anomalous gold values. In the Red Queen/ Chessman area anomalous gold in rock chip values are Iocated at the north-west margin of the syncline where the soil results are anomalous. At the Chlorite Hills/O’Shea’s area anomalous gold in rock chip samples seem to be associated with the eastern margin of the Maud Dolerite.

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10

Drilling

Over 94,000 meters of RAB, RC and Diamond drilling has been completed at the Maud Creek Project and surrounding areas (Table 10-1 and Table 10-2). While the drilling prior to 2011 was not completed by Newmarket Gold, a significant amount of historical data is available for review and reporting. Based on the quality of the data available, SRK see no indication that the drilling information cannot be used for a Mineral Resource estimation.

Table 10-1:     Drill statistics for the Maud Creek deposit

  Diamond RC
Company Period # of holes Meters # of holes Meters
CGAO 2011 6 3 180.33 14 702
Terra Gold 2007 1 211.50 0 0
Kalmet/Hill 50/Terra Gold 1995-97, 2001-2002, 2005 70 19 800.12 0 0
Kalmet 1994 0 0 356 42 390.8
Kalmet 1993 0 0 36 2,212
Placer 1990-91 20 2 992.37 0 0
Placer 1989-91 0 0 36 3,503.8
Total 97 26 184.32 442 48 808.60

Table 10-2:      Historical drilling be previous tenement holders

Prefix # of holes  Type  Company Period Notes
DRC 44 RC KGNL    
GRC 17 RC KGNL    
KR 282 RAB KALMET    
MC01 - MC25 25 DR MCML Dec 1993  
MCE001 - 021 21 RC KGNL    
MCP037 - 301 265 RC KALMET Sep 1994 - Jun 1996  
MCP302 - 416 115 RC      
MCP417 - 488 40 RC HILL 50 Aug 2001 - Aug 2005  
MCW 54 RC KGNL Apr 1998 - Nov 1998  
MD001 - 033 33 DD KALMET 1995  
MD034 - 052 21 DD KGNL 1997 Includes 2 wedge holes
by HILL 50 (2001).
MD053 - 062 14 DD HILL 50 Aug 2001 - Aug 2002  
MD063 - MD065 2 DD TGML Jul 2005 - Aug 2005  
MRB 216 RAB KALMET Aug 1997 - Sep 1997  
MRC 36 RC KALMET Jan 1993  
RAB 588 RAB MCML Oct 1993  
SRRC 14 RC KGNL May 1998  
SWM 9 RC KALMET Jun 1998 - Sep 1998  
TMCD 1 DD TGML    
WD 20 DD PLACER Jun 1990 - Jun 1991 Includes diamond tails
to WP holes
WP 40 RC PLACER Jun 1989 - Jun 1990  

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10.1

2011 Drilling program

Newmarket Gold, then trading as Crocodile Gold conducted infill drilling at Maud Creek in 2011: Diamond – Some holes were collared using PQ (122.6 mm) to allow for better recovery and to prevent collar collapse, this was then reduced to HQ once ground conditions allowed (Table 10-3). Diamond drilling was used at Maud Creek by Newmarket Gold (Figure 10-1) to ensure accurate logging of structures, lithology, alteration and mineralization, as well as capture of geotechnical data. Several diamond drillholes were also used for metallurgical sampling.

Table 10-3:      Parameters used for infill diamond drilling at Maud Creek

Deposit Diamond RC First Drilled Last Drilled
Holes Meters Size Holes Meters  Size
Maud Creek 6 3 179 HQ-PQ 14 700 5” 19 Sept 2011 16 Nov 2011

Reverse Circulation – RC drilling was used in areas where diamond drilling was not required or appropriate, for example, to test the potential mineralization to the south of the Maud Creek pit where limited drilling was identified.

Figure 10-1:      Plan view of the RC and DDH Drilling completed in 2011

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The objective of the drilling program was to test the down plunge component of the Maud Creek deposit at depth. Four holes were used to test for deposit extension to the south, while two were used for verification of the resource model (Figure 10-2 and Table 10-4).

Figure 10-2:      Long section looking west showing grade shells and 2011 diamond drillholes

Table 10-4:      Maud Creek 2011 drilling data

Hole_ID XCOLLAR YCOLLAR ZCOLLAR DEPTH TYPE AZIMUTH DIP
MC001 19550 8799.996 1130 339 DD 275 >-60.0
MC002 19500 8849.996 1130 372.6 DD 271 -59.7
MC003 19650 8799.999 1130 414 DD 279 -60.0
MC004 19600 8849.999 1130 483.6 DD 275 -60.0
MC005 19820 8974.997 1130 756.6 DD 270.9 -60.8
MC006 19770 9039.992 1130 813.5 DD 270 -60.1
MCRC001 19228.856 8539.006 1130.685 50 RC 275 -60.0
MCRC002 19281.995 8537.7 1130.949 50 RC 275 -60.0
MCRC003 19332.654 8542.016 1131.567 50 RC 275 -60.0
MCRC004 19384.094 8523.776 1132.461 50 RC 275 -60.0
MCRC005 19232.577 8725.36 1129.345 50 RC 275 -60.0
MCRC006 19282.385 8746.099 1130.508 50 RC 275 -60.0
MCRC007 19330.733 8732.725 1131.416 50 RC 275 -60.0
MCRC008 19381.4 8733.986 1131.779 50 RC 275 -60.0
MCRC009 19235.469 8937.985 1128.153 50 RC 275 -60.0
MCRC010 19286.715 8933.594 1128.934 50 RC 275 -60.0
MCRC011 19335.354 8930.858 1128.718 50 RC 275 -60.0

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Hole_ID XCOLLAR YCOLLAR ZCOLLAR DEPTH TYPE AZIMUTH DIP
MCRC012 19384.542 8936.011 1128.761 50 RC 275 -60.0
MCRC013 19789.85 9342.812 1124.795 52 RC 275 -60.0
MCRC014 19210.585 10341.349 1124.459 50 RC 275 -60.0

10.1.1

Surveying

All holes were either surveyed by single shot and/or gyro survey. All drillhole collars were picked up by surveyors, and some historical drill collars were also resurveyed.

10.1.2

Core Recovery

Core recovery was not recorded for the Maud Creek 2011 drilling.

10.2

Sampling prior to 2011

Prior to the then Crocodile Gold’s acquisition, four companies had previously run exploration programs at Maud Creek, utilising both Diamond (23,004 m) and RC (48,107 m) drilling techniques (Table 10-2). A summary of the associated procedures was carried out by Snowden in their 2006 report Addendum to the Technical Report entitled Independent Technical Review of the Burnside, Union Reefs, Pine Creek and Maud Creek Gold Projects, Northern Territory, Australia - Resource Update, Maud Creek Gold Project. This report summarised any available historical data from Maud Creek produced by Placer Dome, Kalmet, Hill 50 and Terra Gold. GBS Gold requested the review by Snowden in 2006, but never conducted any drilling programs at Maud Creek during their tenure.

10.2.1

Surveying

As part of the review, GBS contracted a professional surveyor to locate all drillhole collars.

GBS also utilised the following downhole survey instruments to verify the orientation of the drillholes:

  Sperry Sun system; and
     
  Single shot system.

Measurements were typically taken every 25 m.

Figure 10-3 and Figure 10-4 highlight the changes in azimuth and dip with drillhole depth as captured by GBS. The drillhole azimuths typically wander +/- 15°, with azimuth showing a relationship with depth.

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Figure 10-3:      Changes in Azimuth with Depth for drilling prior to 2011

Figure 10-4:      Changes in Dip with Depth for drilling prior to 2011.

Several corrections in the acQuire database were made by Newmarket Gold for typographic errors of the single shot surveys. Also, surveys not present in the database were added from source data (MC005 and MC006). Gyro surveys of MC002, MC005 and MC006 were also reprocessed, due to an incorrect application of the grid rotation.

10.2.2

Core recovery

Core recovery data is available for nine diamond holes, all of which were drilled by Hill 50 between 1997 and 2001 (Table 10-5). Overall the recorded recoveries are considered to be acceptable.

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Table 10-5:      Core recovery of drillholes prior to 2011

Hole ID Average recovery %
MD045W1 100
MD045W2 97
MD053 100
MD053W1 98
MD055 98
MD055W1 97
MD055W2 96
MD055W3 98
MD056 99

Verification work completed by Newmarket Gold identified several holes which had core recovery recorded (MD017 – 19) quantitatively as ‘Good’. There are some recovery records in the Database which conflict with measurements in the source data (MD014, MD024, MD029 and MD30), and geotechnical logs.

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11

Sample Preparation, Analysis, and Security


11.1

Sampling Techniques

Samples used to inform the Maud Creek block model estimate are sourced from both diamond drill core and reverse circulation chip samples collected over the last 20 years.

11.1.1

Reverse Circulation Sampling for the 2011 drilling program:

Figure 11-1 outlines the sampling procedure for the 2011 drilling campaign. Sample intervals (1.0 m) are washed and sieved by a geologist and then inspected to determine its geological attributes. Geology is entered directly onto standard logging sheets in either hard copy or digital form via a portable computer using standardised geological codes. Each washed sample is then stored in a chip tray and stored in a shed at the core farm for future reference.

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Figure 11-1:      Reverse Circulation Sampling flow chart

11.1.2

Diamond Sampling for the 2011 drilling program:

Much of the drill core produced from the Maud Creek area is composed of barren sediments and tuff. Therefore, not all diamond drill core is required to be sampled.

Core orientations were marked on the bottom of the core using a Camteq Orishot tool at the end of each rod. Drill core was then orientated by technicians based on these orientation marks. The Geologists then log each hole for weathering, lithology, structures, alteration, mineralization and geotechnical information. Zones of core loss are identified and marked by inserting marker blocks recording the exact length of the core loss.

At the completion of logging geologists mark the core for sampling and photograph each tray dry and wet. Samples intervals are chosen based on lithological or mineralization contacts. Sample boundaries are often made at pre-existing breaks; otherwise the half core is cut perpendicular to the core axis using an Almonte automated diamond core saw.

Minimum sample size is 0.3 m and maximum size is 1.5 m. The core was cut so as to divide the mineralization in half whilst preserving the orientation line. Some drillholes were sampled for their entire length and some were sampled from 20 – 50 m in the hangingwall through to the end of the hole.

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11.1.3

Sampling prior to 2011

Details of the sample collection, preparation and quality control techniques employed by each of the previous operators of Maud Creek are not fully documented. Procedures documented in annual reports written by Kalmet Resources (1996) note that:

 

Sampling techniques varied for each drilling program;

   

 

 

A review of the assay database, analytical quality control and sampling techniques was undertaken in July 1996 by Geocraft Pty Ltd, an independent consultant;

   
 

Two meter composite RC samples were collected by riffle splitting; and

   

 

 

5-6 kg samples were dispatched to Alice Springs, where they were riffle split to a nominal 3kg prior to pulverizing of the entire sample.


11.2 Data Sampling and Distribution

The Maud Creek model has been shown during validation to be subject to varying drillhole density and sample locations, which has affected lode geometry. Within the upper/central parts of lodes the drilling is regular and of sufficient density, but subject to decreasing densities and irregular spacing at depth.

11.3

Testing Laboratories


11.3.1

2011 Drilling program

Assaying of the drill core and reverse circulation samples was completed by either NAL at Pine Creek, the NTEL or the ALS labs in Darwin. All laboratories used are independent of Newmarket Gold and are well known to SRK Consulting as competent assayers. Once the assaying laboratory’s personnel receive the drill or chip samples they undertake sample preparation and chemical analysis. Results are returned to Newmarket Gold staff, which validate and input the data into relevant databases.

All analytical work including sample preparation, analytical procedures, QA/QC measures and associated security and chain of custody procedures have been completed in accordance with the established protocols routinely used by Newmarket Gold.

SRK Consulting considers that these procedures and protocols are of acceptable quality and are broadly consistent with international best practice standards. Lab visits have been conducted by Newmarket Gold staff to meet with the management of the laboratories and to inspect the facilities.

11.3.2

Sampling prior to 2011

Various reports written by previous tenement holders document a range of laboratory testing or investigations. The 2006 Snowden report summarised these assessments (Table 11-1):

Table 11-1:      Summary of QAQC reports written for Maud Creek

Company Report Year Description
Placer Report NT25/91 1991 202 Check-analysis by ALS Townsville of samples previously analyzed by Classic Laboratories
Kalmet MG03/002 June 1994 Comparison of gold results in 3 sets of ‘twinned’ RC and Diamond Drillholes
Kalmet KMT196 Table 4 1995 Resampling of 18 intervals in 11 holes that reported > 20 g/t Au
Kalmet KMT196 Section 5.2.3 1996 Rectification of deliberate sample corruption in Laboratory (1297 samples)

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11.4

Sample Preparation


11.4.1

2011 Drilling program

The following sample preparation activities (Figure 11-2) were undertaken by Newmarket Gold staff for the 2011 drilling program:

 

Standards, blanks, barren quartz flush and duplicates placed in pre numbered calico bags;

   

 

 

Sample is placed into calico bags;

   

 

 

Calico bags loaded into green plastic bags with the sequence of samples in the bag labelled on the outside;

   
 

The green plastic bags were then placed into dispatch cages to be picked up by courier and taken to the Laboratory; and

   
 

At the completion of each hole the core trays are stored in racks for future retrieval.

Figure 11-2:      Diamond Drilling Sampling flow chart

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The primary commercial laboratory used for the Maud Creek drilling campaign was Northern Territory Environmental Laboratories (NTEL), (now Genalysis), with Australian Laboratory Services (ALS) in Darwin acting as an umpire lab. Samples sent to ALS were prepared in Darwin and then sent to either the ALS laboratory facilities in Perth or Townsville for analysis.

The following sample preparation activities are undertaken by laboratory staff (Figure 11-3):

 

Samples are received and checked against the submission sheet;

   

 

 

Average sample weight for the submission is taken;

   

 

 

Each sample is then dried at 105°C until fully dry;

   

 

 

Entire sample is initially crushed in a jaw crusher to approximately 2 mm;

   

 

 

Each sample is then rotary split with 300 g taken for milling and assay and remainder set aside as a coarse reject and returned to Newmarket Gold;

   
 

The 300 g sample is then milled to pass through a roll crusher to 2 mm;

   

 

 

Samples are riffle split into two sub-samples – one is milled while the other is retained as a coarse reject and returned to Newmarket Gold;

   
 

The retained sub-sample is milled to 85% passing 75 µm with 1 in 20 samples wet screened to check for compliance;

   
 

Each milled pulp samples is further split to provide 25 g for fire assay (FA25) with an AAS finish and <1 g used for multi-element, if required. At ALS 30 g of the pulp is weighed off for fire assay with an AAS finish (AA26); and

   
 

Any remained pulp sample is kept for future analysis and returned to Newmarket Gold.


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Figure 11-3:      Laboratory sampling flow chart

11.4.2

Sampling prior to 2011

Multiple owners have generated the Maud Creek data set in several phases over the last 20 years. The following references were noted by Snowden in 2006:

  Kalmet Resources engaged Lantana Exploration Pty Ltd in 1995 to re-enter all the analytical data obtained up to that time. This work included all drillhole data up to and including hole number MCP061;
   
 

Kalmet 1996; engaged geological consultant Geocraft Pty Ltd to verify the database and analytical procedures;

   
 

Kilkenny 1998; the MRT 1998 resource study states that data files were merged into an Access database and validated using MRT internal systems and no significant errors were detected;

   
 

Harmony 2003; a competent person has classified the mineral resource estimates for the Maud Creek Project in accordance with the JORC Code and these estimates have been released by Harmony to various Stock Exchanges. This chain of Competent Person Statements as required under the JORC Code has been relied upon by Terra Gold to attest to the validity of the drilling data; and


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Terra Gold has undertaken a validation of the database using Micromine validation routines and reported that no significant errors were detected.


11.5

Sample Analysis


11.5.1

2011 Drilling program

Maud Creek drill core and RC samples were assayed for gold and multi element analysis. Gold grades are determined by fire assay/ atomic absorption spectroscopy (AAS) and multi element (silver, arsenic, bismuth, calcium, cobalt, chromium, copper, iron, potassium, manganese, molybdenum, sodium, nickel, lead, antimony, tin, titanium, zinc, zirconium) by ICP Atomic Emission Spectrometry.

The following procedure is undertaken:

  50 g of pulp is fused with 180 g of flux (silver);
     
  Slag is removed from the lead button and cupellation is used to produce a gold/ silver prill;
     
  0.6 mL of 50% nitric acid is added to a test tube containing prill, and the test tube is placed in a boiling water bath (100°C) until fumes cease and silver appears to be completely dissolved;
   
  1.4 mL of hydrochloric acid (HCl) is added;
     
  On complete dissolution of gold, 8 mL of water is added once the solution is cooled; and
     
  Once the solids have settled, the gold content is determined by fire assay/atomic absorption AAS.

The following procedure is undertaken for multi element analysis:

  A 0.25 g sample is pre- digested for 10-15 minutes in a mixture of nitric and perchloric acids;
     
  Hydrofluoric acid is added and the mixture is evaporated to dense fumes of perchloric;
     
  Residue is leached in a mixture of nitric and hydrochloric acids;
     
  Solution is then cooled and diluted to a final volume of 12.5 mls; and
     
  Elemental concentrations are measured simultaneously by ICP Atomic Emission Spectrometry.

11.5.2

Sampling prior to 2011


 

Kalmet utilised ALS in Alice Springs for gold and arsenic assays;

   

 

 

All samples were analyzed by ALS in Alice Springs for gold by fire assay method PM209 and for arsenic by AAS method G003 or G102;

   
 

Selected holes as MD21 to MD31 were assayed for copper, lead, zinc, silver, nickel, antimony, bismuth, chromium by AAS method G102. Samples from MD21 were also analyzed for Hg by AAS method G008;

   
 

Check samples for gold were analyzed by Analabs in Townsville;

   

 

 

Duplicate check samples were assayed for gold by either ALS in Alice Springs or Assay Corp in Pine Creek; and

   
 

SG measurements were completed on 22 oxide and transition ore and wall rocks, selected from MD15 to MD19. Measurements were completed by Assay Corp in Pine Creek.


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11.6

Laboratory Reviews

The two laboratories used by Newmarket Gold for Maud Creek offer different preparation techniques with the 25 g fire assay by NTEL and a 30 g fire assay by ALS. The following summarises findings with respect to assay work from the two independent laboratories:

There are some errors in the datasets. Typographical errors, wrong standards, recorded/sent to the lab, obvious swaps in the databases. These errors are collaborative from both the laboratories and the database operator;
   
  ALS report lab standards while NTEL do not;
     
  Outright errors should not be appearing in the database (standards and replicates);
     
  Proper control charting methods should be applied to fire assay batches that indicate standards outside proper control limits;
   
  The lack of blanks inserted prior to sample submission needs to be addressed; and
     
  It is of the opinion of SRK that the sampling preparation, analysis and security procedures are all adequate for use in these mineral resource and reserve estimates.

NTEL is an independent laboratory based in Darwin. The relationship between NTEL and Newmarket Gold is on a client/supplier arrangement with a contract in place for service. ALS laboratories are certified using the ISO9001:2008 accreditation (Quality Management Systems – Requirements). They also hold the NATA Technical accreditation under ISO17025:2005.

11.7

Assay Quality Assurance and Quality Control


11.7.1

Standard Reference Material

2011 Newmarket Gold drilling program:

Certified standards are submitted to the laboratory on a regular basis. A standard is inserted into every batch every 117 samples during the RC program and every 26 samples during the diamond drilling program (or less).

There were initially 3 standards across all ranges used during the RC drilling program and 5 standards across all ranges used during the diamond drilling program (Table 11-2).

Each standard for each drill type is charted chronologically to check for compliance and any progressive trends, which may be apparent. An example of the chart used to chronologically check the standards is presented inFigure 11-4.

A total of 48 standards were used against the 1243 samples taken for the diamond drilling program with 6 standards inserted for the 699 samples taken for the RC program. No laboratory standards were inserted.

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Table 11-2:      Standard ST202/5355 Compliance table for Maud Creek Newmarket Gold 2011 drilling program.

Standard ST202/5355
Recommended value 2.37
Mean Result 2.35
AUD% difference versus -1.0
RV  
Standard Deviation 0.07
Number of assays 17
Number > -2SD 0
Number > +2SD 0
% +/- 2SD 100

Figure 11-4:      Standard ST202/5355 Compliance chart for Maud Creek Newmarket Gold 2011 drilling program.

Sampling prior to 2011

The Maud Creek digital database notes that four certified standards were inserted during the Maud Creek drilling programme run by Placer between 1990 and 1991 (Table 11-3).

Table 11-3:      Maud Creek Certified Laboratory Standards

Standard Au (g/t)
OxE20 0.548
OxG22 1.035
OxH19 1.344
OxH29 1.298

Figure 11-5 to Figure 11-8 are plots sourced from pre 2011 drill data, and show difference (%) and absolute difference (%) for each standard assay on the left-hand vertical axis, with gold grade (g/t) on the right-hand vertical axis. The standard value is plotted in purple, and the CRM lab value is plotted in blue. Values of greater than 20% difference between certified value and the determined value appear to be related to submission error or incorrect standard.

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Figure 11-5:      Standard OxH19

Figure 11-6:      Standard OxE20

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Figure 11-7:      Standard OxE21

Figure 11-8:      Standard OxG22

11.7.2

Blank Material

2011 Newmarket Gold drilling program:

Blank materials included in the sample stream were derived from several sources; barren core, barren coarse rejects, crushed Bunbury Basalt (from Gannet Holding Pty Ltd, referred to in this report as blank). Blank results above 0.02 g/t Au are queried and any issues resolved. Results are chronologically charted to visually check compliance (Figure 11-9). No blanks were inserted into the Maud Creek RC drilling program. For the diamond drilling program, a total of 72 blanks were inserted with 98.6% at or below 0.02 g/t Au.

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Figure 11-9:      Compliance chart for blanks used in the 2011 Maud Creek drilling program.

Sampling prior to 2011

There was no evidence in any reporting indicating the insertion of blank material with samples submitted to laboratories.

11.7.3

Duplicate Assay Statistics

2011 Newmarket Gold drilling program:

Relative precisions have been used to analyze the precision of duplicate samples. The relative precision is a measure of dissimilarity, that is, if both distributions are exactly the same, this value will equal zero increases as the distributions become more dissimilar.

In this report, relative precision has been calculated using all data pairs for the ranges of below detection (<0.01 g/t) to 0.20 g/t Au, 0.21 to 0.5 g/t Au, 0.51 to 0.7 g/t Au, 0.71 to 1.00 g/t Au, 1.01 to 1.40 g/t Au, 1.41 to 5.00 g/t Au and >5.00 g/t Au. This is to isolate the large conditional variance of errors associated with assay determinations near both lower and upper analytical detection limits and to selectively analyze results within these set ranges.

An example of the analysis tables for the 2011 Maud Creek drill program is given in Table 11-4 to Table 11-6 and Figure 11-10.

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Table 11-4:      Duplicate analysis table for the 2011 drilling program

Repeat Maud Creek DD NTEL FA25 Total Program
Mean original results: 0.06
Mean repeat results: 0.07
Number of assays: 39
Standard Deviation: 0.07
Sum of Differences: -0.39
Sum of Diff * Diff: 0.18
Mean Difference: -0.01
% Results within +/- 2SD 102
Results within 30% precision level 86
Average absolute % Difference: 24
% Assays original <or = repeat 83

Table 11-5:      Duplicate correlation table for the 2011 drilling program:

Range (g/t) Original vs Repeat
Combined 0.952
<0.20 0.777
0.21 – 0.50 -
0.51 – 0.70 -
0.71 – 1.00 1.000
1.01 – 1.40 -
1.41 – 5.00 -
>5.01 -

Table 11-6:      Duplication R Table for the 2011 drilling program

Range (g/t) # of
assays
% of
total #
Mean
original
Mean
repeat
% diff
between
means
(bias)
Average %
diff
between
assays
(bias)
Absolute
average %
diff between
assays
(total error)
Standard
Deviation
<0.20 38 97 0.03 0.04 -21.8 -9.00 24 0.07
0.21 – 0.50 0 0 - - - - - -
0.51 – 0.70 0 0 - - - - - -
0.71 – 1.00 1 3 1.08 1.19 -10.2 -10.19 10 0.00
1.01 – 1.40 0 0 - - - - - -
1.41 – 5.00 0 0 - - - - - -
>5.01 0 0 - - - - - -
Total: 39 100 0.06 0.07 -16.5 -1.3 24 0.07

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Figure 11-10:      Duplicate correlation plot; range <0.2 g/t for 2011 drilling program

Eighty Six per cent of diamond duplicates for Maud Creek fall within the 30% precision level. All but one of the 39 original samples is above 0.2 g/t Au with 24 original samples being below the detection limit. Of the 29 RC duplicates taken, 24 of the original samples are below the detection limit with 87% falling within the 30% precision level.

Sampling prior to 2011

The database for pre 2011 drilling contains laboratory repeat data (Au1 and Au2). Figure 11-11 is a relative difference plot of the two data sets. It would generally be expected that as the average grade of each pair of data increases the relative difference between the paired data would decrease. The plot does not indicate this, suggesting some issues with laboratory precision. This may reflect the presence of coarse or nuggety gold. It is understood that no screened fire assay analysis were undertaken to help assess whether coarse gold is an issue at Maud Creek. Furthermore it must be noted that these are (presumably) repeats initiated by the laboratory, and not blind submissions of field duplicates, which may be normally expected to show poorer precision than laboratory repeats.

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Figure 11-11:      Relative difference plot original vs lab repeats for pre 2011 Maud Creek drill data

11.7.4

Internal laboratory Repeats

2011 Newmarket Gold drilling program:

Internal laboratory repeats were taken for both RC and diamond drilling at the primary laboratory (NTEL). Relative precisions have been used to analyze the precision of repeat samples. The relative precision is a measure of dissimilarity, that is, if both distributions are exactly the same, this value will equal zero increases as the distributions become more dissimilar.

In this report, relative precision has been calculated using all data pairs for the ranges of below detection (<0.01 g/t) to 0.20 g/t Au,0.21 to 0.5 g/t Au, 0.51 to 0.7 g/t Au, 0.71 to 1.00 g/t Au, 1.01 to 1.40 g/t Au, 1.41 to 5.00 g/t Au and >5.00 g/t Au. This is to isolate the large conditional variance of errors associated with assay determinations near both lower and upper analytical detection limits and to selectively analyse results within these set ranges.

An example of the analysis tables for the 2011 Maud Creek drill program is given in Table 11-7 to Table 11-9 andFigure 11-12.

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Table 11-7:      Repeat analysis table for the 2011 drilling program:

Repeat Maud Creek DD NTEL FA25 Total Program
Mean original results: 0.36
Mean repeat results: 0.35
Number of assays: 356
Standard Deviation: 0.16
Sum of Differences: 2.75
Sum of Diff * Diff: 9.55
Mean Difference: 0.01
% Results within +/- 2SD 98
Results within 30% precision level 96
Average absolute % Difference: 1
% Assays original <or = repeat 97

Table 11-8:      Repeat correlation table for the 2011 drilling program:

Range (g/t) Original vs Repeat
Combined 0.991
<0.20 0.999
0.21 – 0.50 0.993
0.51 – 0.70 0.940
0.71 – 1.00 1.000
1.01 – 1.40 0.998
1.41 – 5.00 0.976
>5.01 0.998

Table 11-9:      Repeat R Table for the 2011 drilling program

Range
(g/t)
# of
assays
% of
total #
Mean
original
Mean
repeat
% diff
between
means
(bias)
Average
% diff
between
assays
(bias)
Absolute
average %
diff between
assays
(total error)
Standard
Deviation
<0.20 271 76 0.02 0.02 -0.5 -0.1 0 0.00
0.21 – 0.50 34 10 0.34 0.34 0.3 0.5 1 0.01
0.51 – 0.70 7 2 0.59 0.60 -2.2 -1.9 2 0.04
0.71 – 1.00 10 3 0.83 0.83 0.0 0.0 0 0.00
1.01 – 1.40 8 2 1.14 1.14 -0.2 -0.2 0 0.01
1.41 – 5.00 23 6 2.74 2.65 2.3 2.3 6 0.25
>5.01 3 1 9.13 8.87 13.5 13.5 21 2.01
Total: 356 100 0.36 0.35 2.1 0.1 1 0.16

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Figure 11-12:      Repeat correlation plot; range <20 g/t for 2011 drilling program

Ninety six per cent of diamond repeats for Maud Creek fall within the 10% precision level. One hundred and ninety seven of the original samples are below the detection limit. Eighty five of the 356 original samples are above 0.2 g/t Au. Of the 7 RC repeats taken, none of the original samples are below the detection limit with 26% falling within the 10% precision level.

   
11.7.5

Inter- laboratory Repeats

   

2011 Newmarket Gold drilling program:

   

An example of the analysis tables for the 2011 Maud Creek drill program is given in Table 11-10 to Table 11-12 and Figure 11-13.

Table 11-10:      Inter-laboratory repeat analysis table NTEL: ALS for the 2011 drilling program:

Repeat Maud Creek DD NTEL:ALS Total Program
Mean original results: 1.52
Mean repeat results: 1.42
Number of assays: 77
Standard Deviation: 0.35
Sum of Differences: 7.75
Sum of Diff * Diff: 9.18
Mean Difference: 0.10
% Results within +/- 2SD 95
Results within 30% precision level 73
Average absolute % Difference: 7
% Assays original <or = repeat 44

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Table 11-11:      Inter-laboratory repeat correlation table NTEL: ALS for the 2011 drilling program:

Range (g/t) Original vs Repeat
Combined 0.990
<0.20 0.953
0.21 – 0.50 0.964
0.51 – 0.70 0.109
0.71 – 1.00 0.743
1.01 – 1.40 0.885
1.41 – 5.00 0.907
>5.01 0.998

Table 11-12:      NTEL: ALS Inter-laboratory repeat R Table for the 2011 drilling program

Range (g/t) # of
assays
% of
total #
Mean
original
Mean
repeat
% diff
between
means
(bias)
Average %
diff
between
assays
(bias)
Absolute
average %
diff
between
assays
(total
error)
Standard
Deviation
<0.20 14 17 0.11 0.10 5.8 6.24 11 0.02
0.21 – 0.50 21 26 0.40 0.39 2.8 2.87 5 0.03
0.51 – 0.70 6 7 0.63 0.66 -4 .2 -4.26 7 0.05
0.71 – 1.00 7 9 0.84 0.84 0.0 0.06 4 0.05
1.01 – 1.40 6 7 1.13 1.11 1.3 1.11 4 0.06
1.41 – 5.00 24 29 2.64 2.36 10.5 7.44 10 0.61
>5.01 4 5 8.34 8.09 2.9 3.71 4 0.45
Total: 82 100 1.52 1.52 0.0 1.1 7 0.34

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Figure 11-13:      NTEL: ALS Inter-laboratory repeats for all ranges NTEL FA25: ALS AA25 for the 2011 drilling program

Seventy three percent of all diamond pulp samples fall within the 10% precision level for inter laboratory repeats. The limited population of inter laboratory repeats for both the diamond and RC programs limits this data. Seventeen percent of all RC pulp samples fall within the 10% precision level for inter laboratory repeats however only 2 samples are above 0.2 g/t with samples approaching the lower detection limit significantly affecting the precision results.

For the 2011 Newmarket Gold soil sampling program:

Soil sampling programs were undertaken at Maud Creek. 2488 soil samples were taken at Maud Creek with 82 (3.2%) of them being duplicate samples. – Samples were sent to ALS in Perth and analyzed using their Ionic Leach MEMS-23 method. Eighty four percent of duplicate samples taken from Maud Creek were within the 30% precision level.

11.7.6

Sampling prior to 2011

Some umpire analysis was undertaken between ALS (the primary laboratory) and Assay Corp laboratories. It is not known whether any common CRM was submitted to both laboratories to assist in calibrating the results. Figure 11-14 is a log Q-Q plot of the ALS data against the Assay Corp data and suggests that ALS is slightly under-reporting relative to Assay Corp.

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Figure 11-14:      ALS: Assay Corp Inter-laboratory check analysis for pre 2011 Maud Creek drilling

Some samples were re-split, re-assayed and compared against the original data. Figure 11-15 is a log scatter plot of the re-split data which suggests a slight bias towards the original assay, although the overall correlation is acceptable. However, as the bulk of the data grades are less than 1 g/t Au, the results have little relevance to the Maud Creek estimate.

Figure 11-15:      ALS Re-split check assays for pre 2011 Maud Creek drilling

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11.8

Sample Transport and Security


11.8.1

2011 Newmarket Gold drilling program

A Newmarket Gold staff member is stationed on the RC drill rig while samples are being drilled and collected. At the end of shift samples were generally transported to the sample collection area where they are stored in crates as they await transportation the lab. Samples are shipped in regular intervals so they are not in crates for a length of time. These samples are located at the Brocks Creek exploration office, which can be secured if no staff member is on site.

In terms of diamond drilling, the core is collected daily from the rig and transported to the exploration office near the old Brocks Creek underground mine. Prior to the samples being transported, a photo was taken on site. This was to ensure there was a record of the material drilled before it left site, it also served the purpose of having a geological record of the drilling in case the core was damaged during transportation. The drill core is then stored in the core shed for logging and sampling. The core shed is located in a compound with security fencing. This location is locked up when no Newmarket Gold staff member is on site. Samples are cut at this location and loaded into lab crates once in calico sample bags as they await collection. These samples are then transported directly to the lab for analysis.

Once assaying is complete the results are returned in digital format to the data entry personnel employed by Newmarket Gold. These files are then loaded directly into a Datashed database. Validation via a visual comparison of standard and blanks against received values. Any questionable results are then raised with the laboratory and resolved. Submissions outside given QAQC guidelines are rejected and not loaded until resolved by the laboratory. The Datashed database is located at the Exploration office and the software is a SQL database with built-in security limiting access to people outside the Company network.

11.8.2

Sampling prior to 2011

After taking custody of the drill core, Geologists’ conducted an industry compliant program of geological logging, photography, density measurements, and core sampling. Core was logged in detail onto paper and then entered into the project database. A site visit was completed in January 2006 by Snowden and the drill core was found to be well handled and maintained.

11.9

Conclusions


11.9.1

2011 Newmarket Gold drilling program

The results from the QAQC analysis of drilling has indicated a good level of confidence in assay grades for use in the resource model. The following recommendations for improvements in the current procedures are:

 

An immediate follow up with the laboratory when controls fail;

   

 

 

Increase in the regularity of blank material within the sample stream with 1:50 for RC drilling and 1:20 for diamond drilling;

   
 

An increase in the regularity of standards inserted to the desired 1:25 rate;

   

 

 

Inter laboratory repeats to meet or exceed a rate of 1:20 to original samples;

   

 

 

Assay results to be thoroughly assessed for errors prior to loading;

   

 

 

Conducting an analysis on barren core that is re–used to serve as blanks for future batches; and

   

 

 

Regular tracking of QAQC compliance.


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11.9.2

Sampling prior to 2011

Recommendations were made by Snowden in 2006 to improve the sampling procedures:

  The current QA/ QC programs should be continued for all future sample programs at Maud Creek;
   
  Continuation of the compilation and documentation of historical work undertaken at Maud Creek;
     
  Systematic analysis and reporting of the QA/QC data acquired during sampling; and
     
  Regular auditing of the database and sampling procedures in order to maintain the integrity of the database.

Since Newmarket Gold has taken ownership of Maud Creek, the recommendations made by Snowden in 2006 have been incorporated into sampling procedures.

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12

Data Verification

A thorough examination of all available information was conducted by SRK and Lee Beer of Newmarket Gold. A summary of the issues identified can be found below:

  Several elements from historical assaying had not been imported into the database, notably sulphur and arsenic;
   
  Differences in end of hole depths between the Datashed and acquire databases;
     
  Grid conversion issues between regional MGA and calculated local grid coordinates;
     
  Lack of clarity when distinguishing between original and re-drilled holes (same collar coordinates but differing end of hole depths);
   
  Issues with gryoscopic surveys due to the magnetic correction being erroneously applied;
     
  Missing assays from the Datashed database;
     
  Missing assays from original source (pdf document);
     
  Different assay values between Datashed and acQuire databases;
     
  ‘Self-referencing’ field duplicates;
     
  Conflicting core recovery values between Datashed and acQuire databases;
     
  Overlapping intervals in geology logging;
     
  Interval gaps in geology logging;
     
  Unrecognized logging codes;
     
  Logged intervals beyond end of hole depth;
     
  Different types of values logged for the same variable; for example, RQD logged at both a percentage and a meter value; and
   
  Missing geology logs from either digital database; only present in scanned pdf document.

All issues have either been rectified by Newmarket Gold/SRK or were deemed immaterial for the current resource estimate and will be entered/corrected in the Datashed database when time permits. SRK also conducted a site visit to verify the logging codes used in the 2011 Newmarket Gold drilling, and any available historic drilling. SRK believes the level of geological logging utilised throughout the Maud Creek drilling programs is sufficiently consistent and representative to use for a Mineral Resource Estimate.

All available QAQC reports were analyzed by SRK, to ensure the sample preparation and analysis conducted for each drill program was consistent with industry standards (Table 12-1). QAQC reports prior to 1998 did not include any analysis of the CRM’s or blanks, but did include an investigation of field duplicates. Check laboratories have been utilised throughout the Maud Creek drilling programs, generally between ALS, Assay Corp and NAL. Generally QAQC reports were created by external consultants, either GEOCraft or Snowden, and concluded that the sample preparation and analysis techniques used had been appropriate and consistent with industry standards.

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Table 12-1:      Summary of QAQC Reports completed for Maud Creek

Company Year Program Laboratory Check
Lab
CRM
Analysis
Checked
Blank
Analysis
Checked
Duplicate
Analysis
Checked
QAQC
report
CGAO 2011 MC (6) NAL ALS Y Y Y SRK
2015/2006
Snowden
MCRC (14) NAL ALS Y Y Y SRK
2015/2006
Snowden
Terra Gold 2007 TMCD (1) SGS Y Y Y GEOCraft
2005
Terra Gold 2005 MD (2) ALS ALS Y Y Y GEOCraft
2005
Hill 50 2001
-
2002
MCP (40) NAL Y Y Y Snowden
2005/
GEOCraft
2005
MD (16) NAL Y Y Y Snowden
2005/
GEOCraft
2005
Anglo Gold 1999 - 2000 MRC Assay Corp Amdel Y Y AngloGold
internal
Standard
monitoring
Kilkenny Gold 1997
-
1998




















DRC (44) Assay Corp N N Y Snowden
2005/
GEOCraft
2005
GRC (17) Assay Corp N N Y Snowden
2005/
GEOCraft
2005
MCE (21) Assay Corp N N Y Snowden
2005/
GEOCraft
2005
MCW (51) Assay Corp N N Y Snowden
2005/
GEOCraft
2005
MD (19) Assay Corp N N Y Snowden
2005/
GEOCraft
2005
SRR (14) Assay Corp N N Y Snowden
2005/
GEOCraft
2005
Kalmet 1993

1996

KR (282) ALS Assay Corp N N Y GEOCraft
1996/2005
MCP (265) ALS ANALA BS N N Y GEOCraft
1996/2005

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Company Year Program Laboratory Check
Lab
CRM
Analysis
Checked
Blank
Analysis
Checked
Duplicate
Analysis
Checked
QAQC
report
MD (33) ALS ALS N N Y GEOCraft
1996/2005
MRB (216) ALS N N Y GEOCraft
1996/2005
MRC (36) ALS ALS N N Y GEOCraft
1996/2005
SWM (9) ALS N N Y GEOCraft
1996/2005
Placer 1989 – 1991 WD (20) Classic Laboratories ALS N N Y GEOCraft
1996/2005
WP (40) Classic Laboratories ALS N N Y GEOCraft
1996/2005

12.1

Site Visit

   

Kirsty Sheerin of SRK visited site between the 3 and 7 August 2015 in order to examine and check the logged core that was still available at site (Table 12-2). Observations are detailed below:

   
 

At the immediate footwall to the mineralized zone shale units were regularly observed. Sometime these were logged as such in the primary lithology field, and sometime they were captured in the alteration/structure or comments fields of the database. The shale units are generally very sheared. Where the shale units correlate with mineralization it is possible the low hardness of the shale compared to the sandstone and tuff created a zone of weakness where shearing and subsequent mineralization has occurred. They are generally devoid of veining, so any mineralization is associated with the shale itself, but whether it pre or post-dates the shearing is not known.

   

From a competency point of view, there appear to be two reasons the footwall is less competent then the tuff. The tuff has generally the same composition/provenance as the sandstone footwall (felsic volcanic), but it is obviously more brecciated in texture. This has allowed the silica and chlorite alteration to penetrate the tuff more. This in turn has increased its hardness. In comparison, the footwall is composed of layers of sandstone, siltstone and shales. These individual rock types are more compacted than the tuff and therefore have been less altered, except between the rock types where the differences in grain size is more pronounced. This has allowed regular shearing along the lithology boundaries within the footwall to occur (particularly where the shale is located) and therefore a less component footwall. This variation between sandstone, siltstone and shale doesn’t appear to have been logged consistently, but it would be of significant use from a mining perspective, particularly within the first 10-15 m of the footwall contact.

   

The isolated mineralization observed outside the main corridor of mineralization was also investigated. While there were only a few instances of this in the available core, generally any mineralization was associated with isolated veining or a shear/fault zone.

   

Overall the logging contained in the database correlated well with the main lithological contact boundaries observed in the core. There were slight discrepancies in some areas (generally a lack of detail), but none which SRK deemed to be inappropriate for use in a resource model.


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Table 12-2:     Holes and intervals checked

BHID From To Meters Logged
MC002 284.5 316.47 31.97
MC006 563.4 587.29 23.89
MC004 430.31 463.08 32.77
MD009 96.5 113.8 17.30
MD035 435.72 477.55 41.83
MD038 408 450.3 42.30
MD042 337.25 370.8 33.55
MD027 180.1 218.77 38.67
MC005 578.33 621.41 43.08
MD057 349.9 382.5 32.60
MD051 435.65 473.65 38.00
MD045W1 490.5 520 29.50
MD039 360.8 385.03 24.23
  Total 429.69

The logging correlated well with what was perceived to be the main lithological contact boundaries. There were slight discrepancies in some areas (generally a lack of detail).

At the immediate footwall to the mineralized zone regular shale units were observed. Occasionally they were logged as such, and sometimes it was captured in the alteration/structure. Generally the shales were quite sheared so this is understandable. Where they correlate with mineralization it is possible the low hardness of the shale compared to the sandstone and tuff created a zone of weakness where shearing and subsequent mineralization has occurred. They are generally devoid of veining, so any mineralization is associated with the shale itself, but whether it pre or post-dates the shearing is not clear.

From a competency perspective, there appear to be two reasons the footwall is less competent then the tuff. The tuff has similar composition/provenance as the sandstone footwall (felsic volcanic), but it is obviously more brecciated in texture. This has allowed the silica and chlorite alteration to penetrate the tuff more pervasively. This has subsequently increased its hardness. In comparison, the footwall is composed of layers of sandstone, siltstone and shales. These individual rock types are more compacted than the tuff and therefore have been less altered, except between the lithologies where the difference in grain size is more pronounced. This has allowed regular shearing along the lithology boundaries within the footwall to occur (particularly where the shale is located) and therefore a less component footwall. This variation between sandstone, siltstone and shale doesn’t appear to have been logged consistently, but it would be of significant use from a mining perspective, particularly within the first 10 – 15 m of the footwall contact.

The isolated mineralization observed outside the wireframed 0.1 g/t halo was also investigated. Only a few instances of this were available at the core shed to inspect, however, the few examples seen suggest that generally any mineralization is associated with isolated veining or a shear/fault zone.

The information gathered from the validation logging was then cross checked thoroughly with the database, and SRK believe the geology model created was appropriate for use as the basis of the resource estimate. Any variations observed with the logging will be used to better understand the genesis of the deposit and help design a future infill drill program.

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12.2

2011 Newmarket Gold soil sampling program

Newmarket Gold utilize specialized industry computer software to manage its drillhole and assay database and employ dedicated personnel to manage the database and apply appropriate QAQC procedures to maintain the integrity of the data. Data is assessed for errors against standards and blanks prior to loading into Maxwell GeoServices Datashed™ database software. Data is then spatially assessed in commercially available mining software package Micromine™ for any other questionable results.

Previously, consultants have completed various database checks, which have not identified any reportable errors, which would have raised any concerns about the integrity of the data. During the preparation of this report, which has included search and lookup of assay results, generation of plans and sections and estimation of mineral resources, the Qualified Persons did not encounter any difficulties with the database; SRK believes the historical data/database has been verified to a sufficient level to permit its use and confidence in its reliability.

Wherever possible Newmarket Gold has also conducted on ground checks of data, this includes the– re surveying of historic drill collars and previously mined open pits. The checking of the open pits has involved the use of a surveyor with a depth sounder to test the bottom of the pit against previous pit pickups. This was done to ensure an accurate depletion of the Mineral Resource.

During the past 2 years Newmarket Gold has spent a large amount of time and money reviewing all historic data in both hard and soft copy forms. This has given the Company a much better understanding of the original data that is available for cross checking and review.

12.3

Sampling prior to 2011

Access software was implemented to manage the Maud Creek database in 2006. The software includes a strict, controlled and structured set of fields and columns to manage the data flow, and checks to alert the database manager of any data importation issues.

The geological interpretation, core logging facility and core storage areas were inspected by Snowden in 2006. In all instances the lithologies, mineralization, alteration and sample intervals were found to agree with the drill logs.

Snowden reviewed the database and confirmed that the data extracted for resource estimation matched the primary database records. Overall the review in 2006 concluded that the data has been verified to a sufficient level to permit its use in a CIM compliant resource estimate.

In 2006 Snowden reviewed all previous drilling data and concluded that the lack of documented and relevant QAQC data and protocols was material to the previous estimate, and until addressed would impact upon the ability to classify the resource estimate with greater confidence than Inferred. Based on this advice a resampling programme was implemented by GBS, whereby remaining core was resampled and assayed together with the submission of independent certified reference materials (CRMs).

In July 2005 179 previously cut and analyzed core intervals were resampled and submitted to SGS laboratory in Perth for analysis along with a number of CRMs. The results of the programme are detailed in the Technical Report Maud Creek Project Drill-Hole Data Validation for Resource Assessment by Andrew Milne of GeoCraft Pty Ltd dated August 2005. Standard fire assay analysis was undertaken on the 179 samples, and then 57 samples were re-assayed using screen fire techniques to compare against the corresponding fire assay analysis.

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The following concerns regarding the programme were identified:

  Sampling and analysis of different parts of the core,
     
  Different proportions (half vs a quarter) of the core in some cases; and
     
  Different Laboratories performing the initial (ALS or Assay Corp) and Resample (SGS) analysis.

Figure 12-1 and Figure 12-2 are precision plots comparing the original data vs the resample and the fire assay data vs the screen fir e data, respectively. Figure 12-1 shows about 30% of the data plotting above the 20% precision line, which given the concerns raised above is an acceptable level. Figure 12-2 suggests that coarse gold is not affecting the fire assay results and that they can be considered acceptable for use.

Figure 12-1:      Precision plot original vs resample

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Figure 12-2:      Precision plot fire assay vs screen fire assay

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13

Mineral Processing and Metallurgical Testing


13.1

Metallurgical Testing

An extensive program of metallurgical testing was carried out from 1994 through to 2006. Much of the focus and testing was on downstream oxidation processes on refractory mineralization, such as bioxidation (Biox) and the GEOCOAT® Process. This summary of metallurgical testing considers only those parts relevant to the current flowsheet selection, i.e.:

  Crushing and grinding.
     
  Gravity recovery.
     
  Flotation.
     
  Tailings and concentrate dewatering.

Direct cyanidation leaching of mineralization and concentrates was tested on the fresh (sulphide) mineralization with poor results and is omitted from this summary.

The major body of work was undertaken for Kalmet Resources N.L. by Ammtec in 1996 – 1998. Nine separate reports focussed on flotation testing, while only one dealt directly with SAG milling.

SAG milling was not tested in detail because the previous owners envisaged a relatively low throughput processing facility (300 ktpa) and a conventional crushing plant suitable for an oxide material gold heap leach as well as the fresh material mill feed. For the current development strategy, more SAG milling test work is required.

13.1.1

Comminution

Three reports cover measurements of physical parameters for crushing and grinding:

  Ammtec report A5161 Metallurgical test work on variability samples VL 5-8 and VF 1- 8 from the Maud Creek Project for Kalmet Resources (December 1996) .
   
  Ammtec report A6076 Metallurgical testing of variability samples VF9 – VF15 from the Maud Creek Gold Project for Kilkenny Gold NL (February 1998).
   
  Ammtec report A6443, SAG milling test work associated with the Maud Creek Gold Project for Kilkenny Gold NL (October 1998).

Crushing

Reports A5161 and A6443 include measurements of crushing work index (CWi) and unconfined compressive strength (UCS).

Report A5161 describes samples VL 5 – 8, selected from hole MD20 intervals to represent a profile of increasing depth and sulphur level (less oxidised). Ten specimens were selected from these intervals for crushing work index tests and 5 specimens were selected for UCS tests. Results are as follows:

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Table 13-1:      Crushing Test results from Report A5161

Sample Drillhole Depth (m) Ore type CWi (kWh/t) Lithology
CWi-1 MD21 28.5 Oxide/transition 16.2 Oxidised tuff, quarts & carbonate
CWi-2 MD21 32.7 Transition 8.5 Breccia tuff, quartz carbonate, graphite
CWi-3 MD21 41.5 Transition 11.1 Massive quartz, graphitic, minor pyrite & arsenopyrite
CWi-4 MD20 67.3 Primary 8.2 Graphitic quartz carbonate stockwork, arsenopyrite
CWi-5 MD20 82.0 Primary 10.9 Massive quartz, pyrite, arsenopyrite,
CWi-6 MD14 86.1 Primary 13.5 Tuff, weak stockwork, pyrite
CWi-7 MD14 92.0 Primary 10.0 Tuff
CWi-8 MD14 105.1 Primary 8.5 Massive quartz vein, minor graphite
CWi-9 MD14 109.2 Primary 8.9 Quartz breccia, graphite, pyrite, arsenopyrite
CWi-10 MD14 112.5 Primary 5.9 Graphite quartz stockwork, pyrite, arsenopyrite
      Average 10.2  
      Maximum 16.2  
      Minimum 5.9  
      75th percentile 11.05  

The UCS results ranged from 52 – 285 MPa with an average of 163 MPa.

Report A6443 describes the samples tested as from ‘5 trays of recently drilled HQ core’. Intervals are specified but the exact drillhole was not identified. The core includes oxide, main lode and hanging wall intervals. Four sulphide core specimens were tested for UCS; three from the main lode and one from the hanging wall. Sixteen sulphide core specimens were tested for crushing work index (CWi). Results are as follows, they exclude the oxide sample test results:

Table 13-2:      Crushing Test Results from Report A6443 (Oxides Excluded)

  CWi (kWh/t) UCS (MPa)
Average 10.7 132
Maximum 23.0 182
Minimum 4.5 106
75th percentile 13.2 139

These results indicate moderate average power requirements, but with wide variation. The UCS results would be classified as strong (60 – 200 MPa), with the maximum result approaching the very strong level (>200 MPa) but is not at levels that cause crushing difficulties with appropriate equipment selection. An additional sample VL5-8 had additional UCS tests undertaken on it. The deepest sample demonstrated very high competency (285 MPa) and a 75th percentile of 208 MPa. The combined UCS 75th percentile used for the process design criteria (PDC) is 182 MPa.

The average CWi is only moderate in strength but the maximum level is would be classified as strong. In both cases (UCS and CWi), this suggests more tests would be required to get a reliable average. In the absence of further data, a conservative value is chosen for design; however, they support SAG milling as a viable comminution flowsheet option.

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Grinding

Ammtec report A5161 describes composite samples VF4/6 being subjected to measurements of abrasion index and bond rod and ball mill work indices (BRMWi and BBMWi). The composite was a 50/50 blend of primary ore sample VF4 and primary footwall sample VF6:

  VF 4: Drillhole MD 20, (61.0 - 74.0 m, 80.0 - 83.7 m).
     
  VF 6: Drillhole MD 14, (98.9 - 116.5 m).

These samples had quite high gold grades of 10.83 g/t and 12.31 g/t respectively.

Ammtec report A6076 describes the BBMWi testing of composites VF10 and VF14. The sample origin is provided as:

  VF 10: Drillholes MD 40 (323.3 – 329.8 m), MD 41 (314.55 – 319.0 m), MD 44 (325.0 – 337.0 m).
   
  VF 14: Drillholes MD 35 (459.4 – 464.8 m), MD 37 (463.0 – 469.1 m), MD 38 (429.0 – 436.0 m).

Results from the two reports are summarised as follows:

Table 13-3:      Grinding Test Results from Reports A5161 and A6076

Average depth
(m)
BRMWi
(kWh/t)
BBMWi
(kWh/t)
Earlier work   16.65 17.73
VF4/6 78 17.8 18.1
VF10 283   18.6
VF14 392   19.64
Average   19.0 18.8

Grinding power consumption was also recorded in the flotation pilot plant runs:

  1996 pilot run: 12.24 kWh/t.
     
  1997 pilot run: 11.95 kWh/t.

The equipment and feed size were not in accordance with the standard Bond work index methods, so the results should be treated with caution, however they are indicative. In summary, all the comminution test programs indicate a moderately hard, highly abrasive ore requiring high grinding energy and highly variable crushing energy.

Summary data from John MacIntyre’s comminution evaluation report are as follows:

Table 13-4:      Summary of Comminution Results from J MacIntyre Metallurgical Evaluation (Sept ‘98)

    Oxide Primary
Bond crushing work index kWh/t 5.0 10.2
Bond Rod mill work index kWh/t 16.65 19.0
Bond Ball mill work index kWh/t 17.73 18.8
Abrasion index   0.489 0.678
Unconfined compressive strength MPa 49 163

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The following observations are reproduced from the MacIntyre report:

  1.

The average Bond crushing work index values for both the oxide and primary zones are low. Although the database is limited the primary zone crushing work index does not appear to increase with depth to a vertical depth of 100 meters.

     
  2.

Both the oxide and primary zone rod and ball mill work index values are high. No rod mill work index value has been measured on samples obtained from a depth greater than 78 meters. The primary zone ball mill work index values do not appear to be depth sensitive to depth of 392 m.

     
  3.

The oxide abrasion index is above average and the primary zone abrasion index is high.

     
  4.

The unconfined compressive strength (UCS) values are strongly dependent on depth, increasing from 37 MPa at 5 meters depth to a very high value of 285 MPa at 90 meters depth. No UCS values have been measured on samples obtained from a depth greater than 90 meters.

Subsequent to the MacIntyre report, Ammtec report A6443 described abrasion index (Ai), milling work index and JK drop weight tests on a comminution composite sample made up of HQ core from the main lode and hanging wall (oxides excluded). As with the crushing work, the intervals are specified and come from a single, unidentified drillhole. Results were very similar to the earlier reports as shown below.

Table 13-5:      Comminution Results from Report A6443

Ai BRMWi
(kWh/t)
BBMWi
(kWh/t)
Maud Creek
‘Comminution composite’
0.6487 18.3 19.94

The same composite sample was used for JK Drop Weight tests, giving the following results:

  A: 72.8;
     
  b: 0.68;
     
  Axb: 49.5; and
     
  ta: 0.41.

These parameters define the ore-specific breakage function which can be input to the JK simulation software to predict SAG mill performance and sizings. The tests show this sample to be moderately hard.

Further metallurgical testing is recommended to establish the SAG mill performance, through additional DWi and possibly Advanced Media Competency Tests (AMCT) as well as additional comminution tests on samples at depth. A single composite from a single drillhole does not give sufficient confidence in performance of this critical operation. Future tests should be based on geo-metallurgical domains in order to match grinding requirements with the mine plan and to ensure sample representivity for the new mine plan.

Relatively conservative comminution parameters have been incorporated into the design criteria. No significant risks are considered in this aspect of the test work but additional test work is required to provide further confidence and optimisation of any future design.

13.1.2

Gravity Gold Recovery

Gravity Recoverable Gold (GRG) test work was included in the three flotation pilot plant runs as well as several of the batch test work programs.

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Several preliminary programs were conducted by Amdel and Metcon in 1994. Gravity recovery was found to be significant but highly variable.

Ammtec report A4997 Optimisation flotation testing of Maud Creek primary gold ore for Kalmet Resources (May 1996) describes flotation and gravity testing of a composite prepared from drillhole WD 16, 122 – 144 m. Batch tests were done in a Knelson (gravity) Concentrator after crushing the sample to 100 % passing 1 mm, giving 11.36 % Au recovery.

Ammtec report A5161 Metallurgical test work on variability samples VL 5-8 and VF 1-8 from the Maud Creek Project for Kalmet Resources (December 1996) included gravity pre-treatment by Knelson Concentrator on samples designated VF 1 – 8. Gravity gold recovery varied from 2.05 % to 74.05 % on the VF5 sample, noted as being high in carbonates. Results are provided in summary Table 13-6.

The high carbonate sample VF5 was further investigated in Ammtec report A6260 Flotation optimisation work associated with the Maud Creek gold project for Kilkenny gold NL (June 1998). As previously demonstrated, very high gravity recovery was observed. A 65.7 % recovery of feed gold was recorded from a Knelson concentrator treating P80 500 micron ore.

The first pilot plant operation is described in Ammtec report A4952 Pilot scale flotation testing of Maud Creek primary gold ore for Kalmet Resources (March 1996). The bulk composite used had a relatively high gold head grade of 9.88 g/t (average). Gravity gold recovery equivalent to 14.5 % of the feed was reported. There was also free gold found in the flash flotation concentrate, but this was deemed too fine to be gravity recoverable.

The second pilot plant also included gravity recovery, as described in Ammtec Report A5367 (Part A), Pilot scale flotation testing of Maud Creek primary ore for Kalmet Resources NL (March 1997). Gravity recovery equivalent to 14.5% of the feed ore was reported.

Ammtec report A6076 Metallurgical testing of variability samples VF9 – VF15 from the Maud Creek gold project for Kilkenny Gold NL (February 1998) included a series of gravity recovery tests. Unlike previous tests, results were generally poor. However, the gravity tails from these tests all responded well to flotation and overall recoveries remained high. A summary table is provided in Table 13-7. The J MacIntyre report (Sept 98) makes the following observations in respect to gravity gold recovery:

  1.

The amount of gravity gold recovered generally increases with gold head grade. A minimum amount of gravity gold may be recovered for gold head grades generally less than 5.6 g/t.

     
  2.

The first and second pilot plants recovered …15.8% average….of the total gold.

     
  3.

The first and second pilot plants recovered 1.430 g/t and 0.975 g/t (1.203 g/t average) of their head grades of 7.837 g/t and 7.373 g/t (7.605-g/t average). That is 18.2% and 13.2% (15.8% average) of the total gold that was recovered as amalgam gold.

     
  4.

The amalgam gold- gold head grade relationship predicts that approximately 1.55 g/t or 20.4% of the gold is recovered as amalgam gold for the average pilot plant head grade of 7.605 g/t. This relationship has been adjusted downwards by 0.35 g/t such that it reflects the actual average amount of gold recovered by both the pilot plants. The adjusted relationship therefore predicts that 1.049 g/t or 14.1% of the total gold to be recovered as amalgam gold for a 7.40 g/t head grade.

     
  5.

The amount of amalgam gold recovered appears to decrease with depth, especially when samples grading more than 10 g/t are excluded from the database.

     
  6.

The amount of amalgam gold recovered appears to be independent of whether the sample is oxide, transition or primary ore.


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  7.

The amount of amalgam gold recovered appears to be independent of either the sulphur or arsenic head grade.

Some later test work also assessed gravity recovery. IMO project No. 1930 Scoping test work on Maud Creek gold ore samples for Harmony gold operations Ltd (June 2003) included gravity separation of a low (1.75 g/t) and high grade (6.14 g/t) composite. The composites were prepared from intervals of main zone (MZ) and eastern shear (ESZ) from a several different drillholes. Gravity recovery to low grade concentrates was reported as 44% from the low grade and 53% from the high grade. These results are probably not realistic due to the high mass pull reporting to the gravity concentrate but support the inclusion of a gravity recovery circuit and potential for reasonable gravity recovery on grades more likely to be fed to the processing plant.

The most recent pilot program is covered in Ammtec report A9911 Pilot flotation on Maud Creek deposit for Terra Gold Mining Ltd (February 2006). Composites samples were prepared from drillholes MD63 and MD65. Gravity recovery of 18.9 % was recorded from a head grade of 1.94 g/t.

In summary, the Maud Creek contains significant but varying amounts of gravity recoverable gold. The relationship presented by John MacIntyre is as follows:

g/t amalgam gold = 0. 760 x (gold head grade) - 4.58

Based on the relationship above, the amount of gravity recoverable gold would be minimal (negative) for an average head grade of 4.38 g/t. This is not supported by much of the test work, there are significant variations in recoveries in the various test work that has been completed, and the value given in the John MacIntyre report is considered too conservative.

In summary, it is recommended that further gravity recovery test work be performed on ore that more closely reflects the current average head grade. A gravity circuit would typically be included for GRG above 10% and in this case is further supported given the cost of transporting a flotation concentrate for third party processing.

Until further assessment is made on the deposit, a gravity recovery of 20% has been used for design purposes based on variability and pilot plant results. If gravity recovery proves to be lower than this, test work demonstrates gold is subsequently recovered in flotation.

13.1.3

Flotation

Eleven separate test work reports describe the flotation programs undertaken for Maud Creek. All were carried out by Ammtec Laboratories. The reports can be separated into several phases of work, being; preliminary, variability and optimisation test work followed by pilot programs. A significant amount of flotation test work has been undertaken and is considered to be sufficient for a PFS. Piloting would normally be considered to be at a feasibility level of assessment.

Preliminary programs

Ammtec Report A4909 Preliminary assay and flotation testing of a Maud Creek mineralization composite for Kalmet Resources (January 1996).
     
Ammtec Report A4930 Preliminary metallurgical test work on Maud Creek primary gold ore composite for Kalmet Resources N.L. (January 1996).

The first of these (A4909) was a simple flotation test on sample from hole (MD 09) with head grade 7.16 g/t Au. Flotation recovery was very high at 95.55%, to a concentrate containing 128.0 g/t Au. Cyanide leach extraction on the tail gave only 55.1% recovery.

The second program (A4930) was a single set of rougher flotation batch tests to confirm the suitability of the sample for the first pilot program (A4952). The head grade was quite high (11.1, 8.66 g/t Au). In this sample 94.7% of gold was recovered to a concentrate containing 53.9 g/t.

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The grind size for both of these test programs was 80% passing 75 microns.

Optimisation programs

The next set of tests are considered optimisation batch test work programs.

 

Ammtec report A4997 Optimisation flotation testing of Maud Creek primary gold ore for Kalmet Resources (May 1996).

   
 

Ammtec Report A5376 Part B Maud Creek carbonate depression flotation test work for Kalmet Resources (February 1997).

   
 

Ammtec A6260 Flotation optimisation work associated with the Maud Creek gold project for Kilkenny gold NL (June 1998).

   
 

Ammtec report A9617 Flotation test work on Maud Creek sample TMCD4002 for Terragold Ltd (May 2005).

The first program aimed to optimise the grind size, circuit configuration, and reagent scheme. The sample was taken from drillhole WD 16, 122 – 144 m, with an average head grade of 3.73 g/t. The optimum reagent scheme was reported to be:

  125 g/t SIBX (collector).
     
  40 g/t AP3477 (collector).
     
  50 g/t CuSO4 (activator) .

However, the performance was not sensitive to either the type or dose of reagent over the ranges tested.

Flash flotation caused only a marginal increase in gold recovery. Cleaner flotation was not included in the program. At the time it was not considered to be necessary.

Grind sensitivity size tests showed a gradual drop in gold recovery with increasing grind size to flotation. From this data, a P80 of 75 microns was selected for subsequent test work. However, this is based on a single drillhole and may not be repeatable.

Report A5376 describes an unsuccessful attempt to depress flotation of carbonates using proprietary reagents. Carbonates are mainly an issue if Biox processing is used downstream; where acid can be a major cost.

Report A6260 focussed on the effect of downstream Biox processing on flotation. Specifically, using acidic water for flotation and using flotation tails for neutralisation of Biox liquor. There was also some extra flotation testing on the high carbonate sample VF5. High overall recovery was found to be possible from closed cycle rougher/scavenger tests, but with some sacrifice in concentrate grade. Report A9617 used intervals from drillhole TMCD04002, 192 – 208 m. Batch flotation tests were done on a composite and also on three separate drillhole intervals with varying gold grades. The composite gave very good flotation performance; with a 95.3 % gold recovery to a 190 g/t concentrate.

Recovery by grind size showed no improvement from 106 microns down to 75 micron. This is in contrast with the previous results of A4997. Grind size optimisation between 75 and 150 microns. The process design criteria has selected a size of 75 microns for recovery and concentrate quality purposes and to reflect pilot plant parameters. There may be justification in relaxing this grind size target marginally in future assessments.

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Figure 13-1:      Recovery by Grind Size (Extracted from Ammtec Reports A4997 & A9617)

Variability programs

  Ammtec report A5161 Metallurgical test work on variability samples VL 5-8 and VF 1- 8 from the Maud Creek project for Kalmet Resources (December 1996).
   
  Ammtec report A6076 Metallurgical testing of variability samples VF9 – VF15 from the Maud Creek gold project for Kilkenny Gold NL (February 1998).

The first of these tested samples taken from varying depth through the ore profile, testing first gravity then flotation on the gravity tails. Results are shown below in Table 13-6.

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Table 13-6:      Flotation Results from Ammtec Report A5161

Sample
composite
Ore zone Drillhole Average
depth
(m)
Gold
grade
g/t
Gravity
recovery
%
Flotation
recovery
%
Overall
recovery
%
VF1 Oxide MD 21 41 15.35 20.91 70.73 91.64
VF2 Oxide/transition MD 21 33 7.63 63.75 29.94 93.69
VF3 Trans/primary MD 21 33 12.72 41.01 55.64 96.65
VF4 Primary MD 20 61 10.83 17.24 79.68 96.92
VF5 Primary hanging wall MD 14 76 7.76 74.05 24.56 98.61
VF6 Primary foot wall MD 14 93 12.31 46.07 52.06 98.13
VF7 Primary MD 3 121 16.09 70.73 28.43 99.16
VF8 Primary MD 27 175 3.36 2.05 93.23 95.28
Average 41.98 54.28 96.26
Maximum 74.05 93.23 99.16
Minimum 2.05 24.56 91.64

Overall recoveries were consistently high although the gravity/flotation proportion varied widely.

The second variability program, A6076, tested samples VF9-VF15. Flotation recovery on ore was consistently over 95%, with one exception, VF12 at 85.65% . The gravity recovery was significantly lower with only one sample demonstrating any notable GRG. These samples were significantly deeper than the previous optimisation tests.

Table 13-7:      Flotation Results from Ammtec Report A6076

Sample
composite
Drillhole
ID
Average
depth
(m)
Gold feed
grade
g/t
Gravity
recovery
Au %
Float
recovery
Au %*
Overall
recovery
Au %
VF9 MD 038 298.7 5.94 0.47 95.44 95.46
VF10 MD 040
MD 041
MD 044
326.6
316.8
331.0
5.96 11.99 95.47 96.01
VF11 MD 036
MD 045
290.6
438.5
14.97 0.24 94.67 94.68
VF12 MD 035 383.7 5.89 0.53 85.65 85.73
VF13 MD 034 438.0 4.77 0.38 97.15 97.16
VF14 MD 035
MD 037
MD 038
462.1
466.1
432.5
3.90 1.94 94.85 94.95
VF15 MD 036
MD 045
393.6

474.8
5.14 1.87 96.67 96.73
Average 2.49 94.27 94.39
Maximum 11.99 97.15 97.16
Minimum 0.24 85.65 85.73

*Flotation recovery % from gravity tail

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There appears to be a grade versus recovery relationship as shown below in Figure 13-2, however, the correlation co-efficient is poor. Even removing the main outlier does not significantly improve it. At the expected feed grade of approximately 4.38 g/t as per the design criteria, the overall recovery of 95% is considered to be conservative.

Figure 13-2:      Recovery vs. grade (Ammtec tests A5161 and A6067)

Pilot plant programs

Three flotation/gravity pilot plants have been run. In each case, the main purpose was to generate concentrate for Biox testing (normally it would be difficult to justify this level of testing for just a conventional concentrator only). The results also provide good process design data.

  Ammtec report A4952 Pilot scale flotation testing of Maud Creek primary gold ore for Kalmet Resources (March 1996).
   
  Ammtec Report A5367 (Part A) Pilot scale flotation testing of Maud Creek primary ore for Kalmet Resources NL (March 1997).
   
  Ammtec A9911 Pilot flotation on Maud Creek deposit for Terra Gold Mining Ltd (February 2006).

The first pilot program used a 5 tonne composite sample. The head grade was quite high, with assay measurements of 11.1 and 8.66 g/t Au. Sulphur grades were also quite high at 2.59% .

The pilot test work circuit included closed circuit grinding with flash flotation and mill discharge passed over a corduroy cloth to collect ‘gravity’ gold. Cyclone overflow passed to rougher, middling and scavenger flotation cells. The reagent scheme consisted of:

  50 g/t of copper sulphate added to the mill;
     
  150 g/t of collector SEX stage added as the collector; and
     
  10 g/t of frother MIBC stage added.

The flotation feed P80 was 118 microns, while most concentrate was reground to 17 microns. Gravity recovery was reported as 18.2%, 92.1% from flotation of the gravity tail and 93.5% overall. The flotation concentrate grade contained 51.4 g/t Au and 19.5% sulphur. The overall concentrate gold grade was 63.82 g/t Au.

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Flotation gold recovery was 3.5% lower than expected from bench scale tests. This was attributed to the very fine nature of the RC drill chips used.

Flash flotation recovered 41% of the gold, 44% of the sulphur and 38 % of the arsenic into 4.0% of the mass.

The second pilot run in 1997 used a bulk sample of lower grade material taken from RC chips, with a lower average Gold grade of 6.69 g/t and 1.85% sulphur. Flotation feed P80 ranged from 87 to 100 microns. Flotation recovery varied from 87.3% to 94.8%, with higher recovery corresponding to lower concentrate grade. This supports the future development of the feed and product grade versus recovery relationship once a more extensive data set is available.

Combined results from the four survey points are shown in the chart below.

Figure 13-3:      1997 Pilot Plant Survey Results

The lower two recovery points included cleaner flotation while the higher recovery points did not. The cleaner stage was added to increase sulphur grade to over 18 - 20%. It also raised the gold grade from 47.3 g/t to 69.4 g/t.

The following points regarding the second pilot plant are extracted from the J MacIntyre evaluation report:

 

The pilot plant consisted of a gravity concentration stage, a flash flotation stage and a secondary flotation stage. Total flotation residence time was 46 minutes;

   
 

Mill feed was crushed to a P100 of 4.0 mm and also had a very fine P80 crushed product size of 1.2 mm;

   
 

A bulk secondary float was employed for the first two days of the pilot plant. A cleaning stage was used on the middling and scavenger concentrates for the last two days; and

   

The plus 180-micron fraction of the flash flotation concentrate (approximately 8.7 -kg wet) was the only product reground to a target P100 of 180 microns.


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The following points are concluded.

  1.

An excellent reconciliation exists between the calculated head grade and the drillhole head grade. The calculated head grade of 7.37 g/t agrees within 1% of the drillhole grade of 7.41 g/t.

     
  2.

A total gravity plus flotation recovery of 91.9% was realised. The first two days recovery of 94.3% was achieved without any cleaning stages – this produced a saleable concentrate grade. This reduced to 89.6% for the last two days when a cleaning stage was employed on the middling and scavenger concentrates.

     
  3.

13.2% of the total gold was amalgamated gravity concentrate gold.

     
  4.

93.0% of the gravity tail gold was recovered by flotation in the first two days. This is 3.2% less than the 96.2% gravity tail recovery predicted from Section 2.9.3's grind-recovery relationship. The very fine nature of the RC drill chip would have also contributed to the much lower than predicted recovery.

     
  5.

Flash flotation recovered 45% of the gold (41% for the first pilot plant), 50% (44%) of the sulphur and 41% (38%) of the arsenic into 3.8% (4.0%) of the mass.

     
  6.

The sulphur recovery for the first two days of 99.0 % is similar to the predicted sulphur recovery of 98.5%. Sulphur recovery reduces to 95.6% when a cleaning stage was employed for the last two days.

     
  7.

The arsenic recovery of 90.1 % for the first two days is similar the 89.4% predicted from Section 2.9.3 's grind-recovery relationships. Arsenic recovery reduces to 85.7% when a cleaning stage was employed for the last two days.

     
  8.

The second pilot plant sample contains 205 g/t of copper, of which 84% is recovered into the combined flotation concentrate at a mean grade 1,801 g/t.

     
  9.

Days 3 and 4's cleaned concentrate contains 4.4% of the carbonate being 54% of the amount of carbonate contained in Day 1 and 2's un-cleaned concentrate of 8.2%.The cleaned concentrate mass of 7.0% is also 57% of the value of the uncleaned concentrate mass of 12.3%. That is both the un- cleaned and cleaned concentrate grades are similar at 6.0 % and 5.4 % respectively.

     
  10.

The amount of calcium contained in the concentrate is similar to the carbonate content. That is 8.0% calcium verses 8.2% carbonate for the un-cleaned concentrate and 3.6 % calcium verses 4.4% carbonate for the cleaned concentrate.

     
  11.

The second pilot plant sample contains a negligible amount of mercury. The highest value recorded in the concentrate was 0.038 g/t.

     
  12.

The combined flotation concentrate's mean P80 of 66 microns is much coarser than the 17 microns for the first pilot plant, but is consistent with the Biox requirements.

The comments regarding carbonates and Biox requirements can be ignored if concentrate is to be exported.

The most recent test program was conducted in 2006 as described by Ammtec report A9911 Pilot flotation on Maud Creek deposit for Terra Gold Mining Ltd (February 2006). Tests were conducted on historical drill core from holes MD063 and MD065, drilled in the late 1990s. The condition of the core may have deteriorated to some extent (oxidised) post drilling. The samples had not been refrigerated.

A series of bench scale tests showed high flotation rougher recovery (~93 - 94%) and also effective cleaner flotation. The pilot flotation run, however, showed lower recovery:

  Gravity gold: 18.9%
     
  Flotation concentrate: 64.6%
     
  Overall recovery: 85.5%

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The grade of the final concentrate was reported as 22.2% sulphur and 40.7 g/t Au. It is not clear why the pilot run gave much poorer results than the batch tests, but the results may have been compromised by the age of the sample or the pilot operation/stability.

13.1.4

Tailings and concentrate dewatering

Thickening tests were carried out by Supaflo (now Outotec) and BPR in support of the Biox process design. The more conservative results from the range below are adopted for the design:

Flotation tailings

  Flocculant dose: 25 – 35 g/t (M-358).
     
  Capacity: 0.7 – 0.8 t/m2/h.
     
  Underflow density: 68 – 70%.

Flotation concentrate

  Flocculant dose: 25- 35 g/t (M-E10).
     
  Capacity: 0.6 – 0.7 t/m2/h.
     
  Underflow density: 56 – 58%.

No data is available on concentrate filtration.

13.2

Future Test work

The Maud Creek mineralization has already been subject to extensive metallurgical testing, adequate to support a pre-feasibility study. However, there remain some gaps that should be covered for definitive feasibility and could reduce the design risk of the PFS if it was decided to undertake additional test work earlier.

Many tests do not specify the lithological domain of the samples, and some do not specify the sample origin at all. In some cases it could be better to design around the uncertainty rather than undertake further testing. For example, the crusher could be sized for the worst-case hardness rather than attempting to determine an accurate average value.

SAG milling parameters are based on a single composite of material from five drill core, but the exact composite makeup is of unclear origin. It would be worthwhile collecting fresh drill core samples by ore domain to confirm the JK (or equivalent) SAG mill parameters.

Much of the comminution test work was undertaken at shallower depths. It may be worth considering additional comminution test work on deeper samples if new sample becomes available. At this level of study materials handling test work is not considered to be essential.

The optimum grind size was determined for conventional crushing followed by ball milling. This should be re-assessed using primary crush – SAG milling using current conditions. There may be a considerable saving in capital and operating cost by choosing a coarser size. This should ideally be conducted on ore-domain composites and supported by economic trade-off analysis.

The flotation and gravity circuits can be designed from the available results. However, confirmation of gold recovery by ore domain would be useful for production forecasting and economic analysis. Furthermore, some flexibility should be built in to the layout to allow for circuit changes, such as retro-fitting flash flotation and/or cleaner flotation. Operating experience and changes in the market conditions/payment terms may justify these measures to increase product grade. Further development of a feed and product grade versus recovery relationship will improve confidence in predicting the overall recovery.

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Limited flash testing was done over the life of the project however when tested, such as in the final piloting program, it has been shown to be relatively effective. The flowsheet has been shown to be simple and robust with good product grades without it and to keep the flowsheet simple and reduce costs; it has been excluded at this point.

Filtration of concentrate samples should be included in future programs. These can be easily added to the flotation confirmation tests and will be valuable for sizing and selection of the filters. It is not considered a risk for design. Conservative assumptions have been made at this time.

Potential customers will most likely want to test samples of concentrate before agreeing to sales terms. Future tests should be used to generate concentrates for marketing purposes.

This study focuses mainly on metallurgy and processing of the sulphide ore. For the oxide and transition material (limited tonnage of transitional included in the LoM), controlled potential sulphidation (CPS) should be considered to enable flotation. CPS is an established technology currently used on gold and combined gold/copper mineralization at several mines in Australia and overseas. After comminution, ore is treated with sodium hydrosulphide (NaHS) which reacts with the oxide minerals, rendering the surface hydrophobic and thus amenable to flotation. This could considerably increase the ore tonnage treatable through the flotation plant. Reagent consumption rates and recoveries would need to be established by testing.

It is noted that gravity and flotation recovery test work on transitional mineralization is limited and the metallurgical behaviour is not particularly well understood. It only makes up approximately 230 kt of the overall LoM feed and therefore extensive testing cannot be justified, including sulphidation, but it will make up a large part of the first year’s tonnage so having a reasonable understanding of its performance, particularly the recovery is important for cash flow. This needs further consideration at the next level of study. Discounted recovery of 85% has been used for modelling purposes but there is likely to be a high degree of variability in recovery depending on the level of oxidation.

13.2.1

Delineation of Mineralization Oxidation Extent

The mineralization can be classified three ways; fresh, transitional and oxide. Fresh material is the primary focus of the study. Small amounts of transitional material could be mined with the fresh material, however most will report to waste. Oxide material will not be generating any revenue for pit optimisation purposes and will be treated as waste. The total oxide tonnage is relatively low as the bulk of it has been previously mined; it does not justify a standalone CIL/CIP gold plant. It remains a project opportunity if it can be processed economically at an existing conventional gold cyanide leach mill such as Union Reef. The intent of this study is to treat the Maud Creek as a standalone project and therefore oxide material has not been incorporated into the base case. It is considered potential project upside.

Based on the classifications in Table 13-8, Zones 3 to 5 are to be treated as ore and will drive the open pit and underground mines.

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Table 13-8:      Classification of Zones in Maud Creek Deposit assumptions

Weathering
Zone
Description Sulphur Total
(%)
Metallurgical recovery (%)
1 Oxide 0.1 0%
2 Oxide/transition 0.43 0%
3 Transition 0.6 85%
4 Transition/fresh 1.26 95%
5 Fresh 1.24 95%

The figure below shows the orebody and is divided into the five ore types, with light blue indicating Zone 1 oxide material, light green indicating Zone 2 oxide/transitional material, green indicating Zone 3 transitional material, yellow indicating Zone 4 transitional/Fresh material, and red indicating Zone 5 Fresh material which makes up the bulk of the overall Project tonnes. Figure 13-4 shows approximately 50 m of Zones 1 to 4 ore covering the Zone 5 Fresh material.

Figure 13-4:      Cross Section of Maud Creek Deposit

The oxides and oxides/transitional material in Zones 1 and 2 will be stockpiled separately as waste but present a future processing opportunity. Zones 3 and 4 are processed but have a lower transitional material metallurgical recovery assigned to them.

13.3

Geometallurgy

In order to assess the impact of the Maud Creek mineralogy on reagent consumption and other geometallurgical considerations during processing, a collation of all existing metallurgical test work was conducted (Table 13 -9). Any reports which detailed the drillhole ID and from/to depth of the samples tested were incorporated into a copy of the Maud Creek drillhole assay table. To ensure there was no confusion with the existing multi-element data, any additional elemental data from metallurgical testing was given the prefix ‘Met_’. All metallurgically tested intervals, including elemental and processing test work, were assigned an identifying ‘Met Sample ID’; a combination of the year of the test work and the ‘To’ depth of the sample. Where composited samples had been collected, the same value was applied for all intervals, as designated by previous assay sampling.

A total of 7,696 values were recorded from the metallurgical reports to create a geometallurgical table. These were then added to the existing 107,677 multi-element values in the assay table. The variables in the geometallurgical database identified as of interest to processing by Simulus include silver, arsenic, bismuth, carbon, carbonate carbon, CO3, organic carbon, total carbon, sulphide sulphur, sulphate sulphur, total sulphur and antimony. Along with collar and survey information this geometallurgical data was then imported into Leapfrog and interpolations creating using the existing structural trends created for the geology model. These interpolations were then imported into Datamine and in conjunction with the drillhole database and lithology logging, amended in cross section to ensure geological considerations were taken into account.

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At this point, due to limited data the carbon (carbonate carbon CO3-2, organic carbon and total carbon) were combined into one carbon wireframe interpretation, and the sulphur sulphide sulphur, sulphate sulphur and total sulphur combined into one sulphur wireframe interpretation. This exercise was conducted to determine whether there was enough data to create geostatistically robust preliminary geometallurgical domains. Unfortunately, the lack of data, presence of composite values (same value across a large interval) and preference for metallurgical testing to the south and west of the main ore body meant this was not possible.

Arsenic and silver were the only two variables with sufficient samples spread across the ore body, due to multi-element testing during previous drill programs. Arsenic was estimated in the block model using its own variography and kriging parameters, but limited to the ore domains previously created based on the gold samples. Due to the low numbers of silver, bismuth, carbon, sulphur and antimony data these variables were was estimated into the same gold ore domains, rather than into their own domains. Also, due to insufficient data to generate variograms, parameters from the gold variography and kriging neighbourhood were used instead.

This process allows the model to indicate that, for example, sulphur testing has been conducted in a certain area, but not a quantification of the amount of sulphur present. Figure 13-5 to Figure 13-10 show the distribution of the testing of these variables compared to the main vein, minimum vein and 0.75 g/t Au halo wireframes which were used as ore domains for estimation. The estimation of these variables will allow a more targeted approach to the next phase of metallurgical testing, and give an indication of how the geometallurgical variables correlate to the existing lithology and alteration spatially.

For Figure 13-5 to Figure 13-10 below the pink wireframe is the main vein, red wireframe the minimum vein and the orange wireframe the 0.75 g/t halo wireframe. All views are long section facing south-east.

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Figure 13-5:      Long Section View Facing South-East Showing Silver Data Compared to Ore Domains 1

Figure 13-6:      Long Section View Facing South-East Showing Arsenic Data Compared to Ore Domains 1

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Figure 13-7:      Long Section View Facing South-East Showing Bismuth Data Compared to Ore Domains 1

Figure 13-8:      Long Section View Facing South-East Showing Carbon, Carbonate Carbon, CO3, Organic Carbon and Total Carbon Data Compared to Ore Domains 1

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Figure 13-9:      Long Section View Facing South-East Showing Sulphide Sulphur, Sulphate Sulphur and Total Sulphur Data Compared to Ore Domains 1

Figure 13-10:      Long Section View Facing South-East Showing Antimony Data Compared to Ore Domains 1

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Table 13-9:      Historical Metallurgical Reports Used to Construct Geometallurgical Database

Year Report Name Company Number of
Samples
Tested
Composites Element Testing Metallurgical
Testing
2005 Maud Creek
Flotation Test
work on Maud
Creek Sample
TMCD04002
Terra Gold 16 16 Au
As

S
Fe
Cyanide
recovery
2003 Maud Creek
Scoping Test work
on Maud Creek
Gold Ore Samples
Harmony Gold 262 14 Au
Ag
As
Cu
Pb
Zn
Hg
S
Bi
Fe
Organic C
Total C
Cyanide
recovery
Preg robbing
Gravity
1998 Metallurgical
Evaluation of the
Maud Creek
Project

Kilkenny Gold 125 125 As
S
Fe
Organic C
CO3
Cyanide
recovery
Inferred pyrite
1998


Metallurgical
Testing of
Variability Sample
VF9
Kilkenny Gold 89 89 Cyanide
recovery
1998 Kilkenny Gold
Resources Leach
testing – hidden in
report
#68_116459
Maud Creek, NT,
Assays, Lab files,
1996-2005
Kilkenny Gold 21 21 Cyanide
recovery
Au head and
tails
1996 Interim Working
Report on the
Metallurgical
Evaluation of the
Maud Creek Gold
Project
Kilkenny Gold 44 23 Au
Ag
As
Cu
Fe
Organic C
Carbonate
Total C
Sulphate S
Sulphide S
Total S
Bottle roll test
Product size
Cyanide
recovery
Cyanide
soluble head
and residue
grade
Reagent data
Lime
consumption
Leach
kinetics
1996 Maud Creek-
Metallurgy
Reports
Kamlet Resources 12 12 Au
Ag
As
Cu

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Year Report Name Company Number of
Samples
Tested
Composites Element Testing Metallurgical
Testing
Fe
Organic C
Total C
Sulphate S
Sulphide S
Total S
1995 Maud Creek-
Metallurgy
Reports
Kamlet Resources 10 10 As
S
Fe
1994 Maud Creek
Metallurgy
AMDEL
2994_OCR
Kamlet Resources 99 99 Au
Ag
As

Cu
Sb
Sulphide S
Total S
Gravity
Cyanide
recovery
1994 Stage 1
Metallurgical
Testing Maude
Creek
Kamlet Resources 25 25 Au
Ag
As
Cu
Sb
Pt
Pd
R
Rh
Os
Ir
Organic C
Carbonate C
Gravity
Inferred pyrite
Cyanide
recovery









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14

Mineral Resource Estimate


14.1

Introduction

The elements to be estimated are Gold (Au), Arsenic (As), Carbon (C), Copper (Cu), Sulphur (S) and Antimony (Sb). The estimation of the elements has been based on assays sourced from drilling data and metallurgical tests where available. The data available as at March 2016 consisted of reverse circulation (RC), diamond core (DD) and rotary air blast (RAB). RAB samples were excluded from the estimation at the exception of 2 samples.

14.2

Lithology and Structural Model

The Maud Creek lithological model was constructed on the local grid coordinates covering dimensions 1,185,000 m (east) and 1,600,000 m (north). The model incorporates several datasets including Diamond and RAB drilling, AngloGold pit mapping and historic SRK aeromagnetic interpretations (SRK, 1998). All datasets were imported into Leapfrog for subsequent 3D modelling. A topography surface was constructed from collar points and used to constrain the top of the lithological model.

The Maud Creek deposit is hosted within the Proterozoic El Sherana Group units and the mineralization hosted at the faulted contact between the Dorothy Volcanic Member and sediments of the Tollis Formation. The Dorothy Volcanic Member strikes approximately north-south and consists of volcanic tuff with minor interbedded zones of sediments. The Tollis Formation strikes north-south and consists of sandstone and metasediments. The deposit is bound to the east by the Maud Creek Dolerite which intrudes the Tuff sequence. A small Andesite body is also observed to the north of the Maud Creek Open Pit. It is located at the faulted contact between the Sandstone and Tuff, forming a discrete body (Figure 14-1, cover units not included). These key units are also overlain by a thin layer of sedimentary cover and Cambrian Volcanics.

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Figure 14-1:      Maud Creek Deposit lithology model showing primary units, cover not shown

The key unit formations described above were determined based on the logging code ‘Lith Code 1’ from the dataset provided by Newmarket Gold. For the purpose of 3D modelling the combination of several lithologies was sometimes required to form a key lithology group (Table 14-1).

Table 14-1:      Maud Creek Lithology Model Groupings

Lithology Group Codes
Cover ALUV, CLA, CLAY, CLY, GO, LOM, MUD, SOIL, SPLT, SND, CALC, CLCR, LMST, BLT
Dolerite DLT, DOL, INTD
Tuff IGM,MAFT,TUF
Sediments (Sandstone) MSED, GYWK, QTZ, Slst, MDST, SDST, Ssl
Andesite ANT
Vein BX, VEBX, VEIN, SHLE

Within the model area eight faults have been identified (Figure 14-2) based on historic AngloGold pit mapping as well as aeromagnetic interpretations conducted by SRK Consulting (SRK, 1998). Orientations of these structures were extracted from the mapped pit data and interpreted based on available datasets. Generally, the faults exhibit reverse movement, with limited offsets in the range of meters. Figure 14-3 shows the interpreted fault architecture, indicating apparent reverse movement along faults. To the south of the Maud Creek Deposit a major east-west structure with sinistral strike-slip movement has been interpreted based on the drilling data and aeromagnetic interpretations. Additional faults are likely present in the modelled area, however only faults which have significant structural control on the deposit have been constructed.

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Figure 14-2:      Interpreted fault architecture of the Maud Creek Deposit

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Figure 14-3:      Cross section view looking north (northing 9008mN)

14.3

Vein Model

Three veins have been interpreted based on the main lithology logging (Lith Code 1); a continuous primary vein, discrete upper vein and discrete lower vein (Figure 14-4). The veins generally follow the main faulted contact between the Sandstone and Tuff units and have an apparent plunge to the south-east, which follows the contact fault (Figure 14-5). The primary vein typically hosts the highest gold grades and lies within the Sandstone/Tuff contact, however deviations from this contact are evident. The upper and lower veins generally hosts lower gold grades and exhibit limited continuity located above and below the faulted contact. Additional vein material was evident in the logging however these veins typically have distinctly limited continuity and have not been included within the final model. The presence of contact veining decreases to the north of the Anglo Pit, based on available drilling.

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Figure 14-4:      Cross-section view looking north of the three modelled veins (primary, upper and lower), contact fault (dark blue) and two of the eight faults modelled (light blue)

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Figure 14-5:      Vein architecture of the Maud Creek Deposit illustrating three primary veins and indicating a south-eastward plunge

The primary vein is strongly associated with the Sediment/Tuff contact but does not strictly follow the contact. Therefore, it has been modelled as an ‘overprinting’ volume on the sediment, tuff, dolerite and andesite lithology wireframes. Lithological codes BX, VEBX, VEIN or SHLE are present in most holes that intersect the contact. Where these codes are absent the vein has, in most cases, been modelled as pinched out. Where SHLE is present at, or proximal to, the contact, the intercept will carry grades similar to that of the BX, VEBX and VEIN intercepts (approximately 4 g/t Au). There are also intervals of SHLE located distal to the contact, and these do not carry grade.

For estimation purposes the upper and lower veins were combined, resulting in two domains; main vein (primary) and minor vein (upper and lower).

14.4

Grade Halo Models

Outside of and adjacent to these lithologically defined veins are many intercepts which carry similar grades to those within the vein itself. These extend up to 25 meters into the hangingwall and to a lesser extent into the footwall. In addition, a greater than 0.1 g/t Au halo can be observed up to 50 meters into the hanging wall and occasionally in the footwall. Excluding the veins mentioned above, the interval statistics did not indicate any grade distinction with lithology type.

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Mineralization consists of two distinct zones (east and west) controlled by two north-south striking structures (Maud Creek Contact Fault and North-South Fault 1) (Figure 14-6). The western zone of the mineralization is primarily controlled by the structural contact between the footwall Sandstone and hangingwall Tuff. This zone illustrates the strongest concentration and highest grade of mineralization within the Maud Creek deposit. The fault contact strikes approximately north-south and is interpreted to have undergone reverse movement. The fault structure is filled with quartz stockwork veins; the primary host of high gold mineralization, with additional gold hosted within the surrounding wall rock. The eastern zone of mineralization is controlled by a north-south striking structure that has been inferred based on aeromagnetic interpretations (SRK, 1998) (Figure 14-6). This structure lies proximal to the contact between the Maud Creek Dolerite and Tuff units. Within this zone gold generally forms steeply dipping discrete lenses with limited continuity noted along strike.

Figure 14-6:      Cross section illustrating 0.75 g/t grade shell (yellow), primary contact vein (red) and fault architecture (blue)

To capture the complexity of the interactions and multiple orientations of the many faults within the deposit, grade shells generated in Leapfrog were used to model the wall rock mineralization in two stages. Although not obvious in the statistics, observation of the grade downhole suggested a sharp break at around 0.75 g/t Au. Therefore, two nested grade shells were modelled at 0.75 g/t Au and 0.1 g/t Au to be used as estimation domains. These were generated using grade from all intercepts, including those within the vein model.

In some areas of widely spaced drilling (50 m – 100 m down dip) the grade shell models could not be made continuous, even though the vein had been interpreted as continuous. This is a limitation of the Leapfrog software and chosen methodology. The alternative was to manually wireframe this domain but this was considered more time consuming and less likely to capture the multiple orientations observed. There are also locations where the grade observed at the contact was too low or thin to sustain a grade shell (Figure 14-7).

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Figure 14-7:      Cross section looking north (9250 mN) exhibiting low grade hole MCP088 through the fault contact.

14.4.1

Assumptions on non- continuity of grade adjacent to vein

The vein material (VEIN, VBX, BX or SHLE) is observed in the majority of holes that go through the contact between the Tuff and the Sediments and the model assumes Vein continuity between holes where the vein lithologies are recorded. The mineralization above 0.75 g/t Au in the footwall (Sediments) and hanging wall (Tuff) directly adjacent to the vein does not show similar continuity. Mineralization > 0.75 g/t Au may be present or absent in either the footwall or hanging wall from one hole to the next. This is observed throughout the deposit. An example is shown in Figure 14-8. MCP125 contains almost no grade in the footwall but 5 m of moderate grades in the hanging wall. The next hole down dip, MD024, contains 7 m of moderate grades in the footwall and 4 m of moderate grade in the hanging wall. The next hole down dip, MCP469 contains 2 m of grade in the footwall and 4 m of grade in the hanging wall. An assumption of continuity from hole to hole for footwall and hanging wall material, particularly in the footwall, cannot be made and this is reflected in the limited connectivity of the 0.75 g/t and 0.1 g/t Leapfrog grade halo domains.

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It is also worth noting that the highest grades do not always occur in the vein and that some vein material is very low grade (Figure 14-9).

Figure 14-8:      Section 9150N (20 m grid)

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Figure 14-9:      Section 9175 (20 m grid)

14.5

Domaining

The mineral resource was estimated into six domains as described in the previous sections:

  0P1WEST;
     
  0P75WEST;
     
  0P1EAST;
     
  0P75EAST;
     
  MAIN (major vein); and
     
  MINOR (minor veins).

14.6

Compositing

GEOVIA GEMS software was used to desurvey and composite the drilling. Drillholes were flagged by the different domains and then composited to 1 m intervals within the units with a minimum length of 0.01 m (Figure 14-10). This resulted in 16.5% of the composites with a length less than 1 m, however the influence of small intervals on the gold grade is limited (Table 14-2). A total of 22,449 composite samples were created.

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Figure 14-10:      Composite length histogram

Table 14-2:     Influence of the composite length on gold mean grade

Domain Au mean grade (g/t) Length weighted Au mean grade (g/t) Differences
0P1WEST 0.54 0.53 1.7%
0P75WEST 4.61 4.75 -3.0%
0P1EAST 0.64 0.64 -0.6%
0P75EAST 3.71 3.86 -4.0%
MAIN 6.50 6.54 -0.5%
MINOR 5.72 6.01 -5.0%

14.7

Metallurgical Samples

The secondary elements arsenic, carbon and sulphur in the database included combined assays gathered from metallurgical studies and drill assays. The numbers of assays from metallurgical studies were minimal. Due to the compositing length the values tend to be repeated several times along a same hole creating a bias in domains with limited samples. Arsenic is an important element and the influence of the arsenic assays from metallurgical studies was analyzed. Table 14-3 indicates a fairly strong similarity between the Arsenic grades from drilling assays and metallurgical samples. Only domains with lower sample numbers indicate strong dissimilarity. Quantile-Quantile plot were also used to assess any arsenic distribution differences. The distributions were negligibly influenced by the additional metallurgical samples as shown in the MAIN domain inFigure 14-11.

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Table 14-3:      Comparison of arsenic assay from drilling only and including metallurgy samples

Domain Samples Number AS mean grade (g/t)
Drilling Metallurgical Difference Drilling Metallurgical Difference
0P1WEST 5271 5365 1.8% 571 600 4.9%
0P75WEST 2149 2440 11.9% 2493 2630 5.2%
0P1EAST 1182 1188 0.5% 791 820 3.5%
0P75EAST 42 102 58.8% 3524 8311 57.6%
MAIN 641 719 10.8% 3352 3442 2.6%
MINOR 98 147 33.3% 2867 3526 18.7%

Figure 14-11:      MAIN – Quantile-Quantile plot comparing arsenic from drilling only and including metallurgy samples (ASSMET)

In MINOR and 0P75EAST, the assays from the metallurgy study represent 33% and 58.8%, respectively, of the total samples. Most of these samples are located on the same hole and have been duplicated due to the compositing, as shown for the MINOR domain inFigure 14-12. Despite the differences observed in Arsenic grades the 0p75East domain and the MINOR domain all metallurgical samples were included in the drilling dataset and used for the estimation as the spatial location locations of the met samples tend to be different to the regular sample grades.

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Figure 14-12:      MINOR – Arsenic - Left: Plan view of data, Right: Cumulative histogram, Red: Arsenic from drilling only, Green Arsenic from the metallurgical samples

14.8

Block Model Definition

The block model size for estimation was set to 5 m x 10 m x 10 m (X, Y, Z) as this was a reasonable selective mining unit size for both open pit and underground studies, was proportional with the steep dipping north striking orientation and was a reasonable compromise to fit both the close and wiser spaced drilling.

The block model for the estimation is a proportional block model based on the domain boundaries. A sub-block model of 0.625 m x 1.25 m x 1.25 m (X, Y, Z) created in Geovia Surpac was used to visualise the blocks report the resources. The vein models are quite narrow in places and this level of sub blocking was the largest that reproduced the actual wireframe volumes within acceptable limits.

Details of the model are shown in Table 14-4.

Table 14-4:      Block model properties

  X Y Z
Origin (m) (Lower SW corner) 19 000 8 645 400
Cell size (m) 5 10 10
Number of cells 140 140 80
Sub cell size 0.625 1.25 1.25

14.9

Grade Interpolations

The estimation strategy differed depending on the elements and domains. The elements that were estimated are Au, As, Ag, C, Cu, Sb and S. No strong correlations between the elements have been noted.

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The Ordinary Kriging (OK) method of interpolation was selected to estimate the gold and arsenic grade within all six domains as these had the most complete coverage. The other elements were separately estimated using an Inverse Distance of order 2 (ID2) within the western domains and MAIN domain and a Nearest Neighbourhood (NN) method of interpolation for the eastern domains and MINOR domain. Ag, C, Cu, Sb, and S do not have complete sampling coverage and as a consequence the block estimates also do not have complete coverage of the Au estimated blocks.

The methodology used for the estimation consisted of

  Performing cell declustering tests;
     
  Studying the influence of outliers;
     
  Validating the choice of using samples from all drillhole type;
     
  Studying the spatial variability for gold and arsenic (variography);
     
  Defining an estimation strategy; and
     
  Validating the estimation.

14.10

Declustering

Drilling includes various types of holes with unequal spacing, the highest drilling density corresponding to the open pit area and near surface. A cell declustering test was performed using a grid specific range specific to each domain and element. The declustering results are shown inTable 14-5, Table 14-6 and Table 14-7. A positive difference means that the declustered grade is lower than the undeclustered grade.

Table 14-5:     Declustering results in the vein domains

Elements MAIN MINOR
Cell Size
(m)
Declustered
Mean (g/t)
Differences Cell Size
(m)
Declustered
Mean (g/t)
Differences
AU 40x40x40 5.67 13% 90x90x90 3.26 43%
AS 90x90x90 2 993.1 13% 90x90x90 2 707.9 23%
AG 60x60x60 3.42 -0.8% 50x50x50 2.84 12.5%
CU 60x60x60 176.3 2.9% 50x50x50 88.9 29.9%
C 60x60x60 1.62 1.4% 50x50x50 1.04 -2.6%
SB 60x60x60 19.4 14.4% 50x50x50 29.95 18.1%
S 60x60x60 1.43 -3.1% 50x50x50 1.8 -5.1%

Table 14-6:     Declustering results in the western domains

Elements 0P1WEST 0P75WEST
Cell Size
(m)
Declustered
Mean (g/t)
Differences Cell Size
(m)
Declustered
Mean (g/t)
Differences
AU 40x40x40 0.47 13% 40x40x40 3.11 32%
AS 50x50x50 817.8 -36% 30x30x30 2 564.2 3%
AG 60x60x60 1.79 13.5% 60x60x60 2.99 28.2%
CU 60x60x60 91.2 4.2% 60x60x60 112.4 5.5%
C 60x60x60 1.75 13.3% 60x60x60 1.59 0.4%
SB 60x60x60 20.0 14.8% 60x60x60 26.2 12.6%
S 60x60x60 1.07 -21.1% 60x60x60 1.32 1.5%

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Table 14-7:      Declustering results in the eastern domains

Elements 0P1EAST 0P75EAST
Cell Size
(m)
Declustered
Mean (g/t)
Differences Cell Size
(m)
Declustered
Mean (g/t)
Differences
AU 40x40x40 0.55 14% 40x40x40 3.77 -2%
AS 40x40x40 991.2 -21% 50x50x50 1 0024.8 -21%
AG 30x30x30 0.91 -4.5% 30x30x30 8.23 0.0%
CU 30x30x30 60.1 2.8% 30x30x30 91.6 0.0%
C 30x30x30 1.89 -0.2% 30x30x30 - -
SB 30x30x30 33.3 3.1% 30x30x30 62.5 0.0%
S 30x30x30 1.24 -69.6% 30x30x30 1.80 -6.4%

14.11 Outliers  

The element grades have a skewed distribution within each domain as shown by the coefficient of variation shown in, Table 14-8 and

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Table 14-9. The histograms show a long tail for high grades. In the eastern domains and Minor domain, the grade distributions of the elements other than gold and arsenic are not well represented due to a limited number of samples.

MAIN domain, Gold grade distribution is strongly skewed with a coefficient of variation of 5.2. Significant high grade samples are skewing the Gold grade distribution. In particular two samples have been identified above 200 g/t are relatively isolated from the main distribution and have a strong impact on the statistics and variography (Figure 14-13). They are located approximately 425 m below surface at the edge of the domain (Figure 14-14).

Figure 14-13:     MAIN – Left and Middle Au histogram: from 0 to 20 g/t and from 20 to 1000 g/t and Right: Variogram cloud, Red and blue: outliers

Figure 14-14:     MAIN – Au samples location – Left: XoY view, Right: XoZ view

Red: outliers

Further analysis was undertaken for each element within each domain. According to the influence of the outliers on the variography and their locations, a top cut was applied or not. When the outliers have a strong impact, a top cut was applied. Values above the top cut were replaced by the top cut threshold.
Table 14-8 and

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Table 14-9 summarise the treatment of the outliers per domain.

Table 14-8:     Outlier treatment in veins domains

Elements MAIN MINOR
# Samples Declustered
Coeff.
Variation
Top cut
threshold
(g/t)
# Samples Declustered
Coeff.
Variation
Top cut
threshold (g/t)s
AU 1 114 5.2 50 4 615 1.7 50
AS 719 0.9 - 147 0.8 10 000
AG 147 1.0 - 28 0.8 6.5
CU 148 2.5 - 28 1.0 300
C 43 0.7 2 31 0.2 2
SB 57 1.0 - 9 0.8 50
S 121 0.7 - 47 0.4 2.1

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Table 14-9:     Outlier treatment in western domains

Elements 0P1WEST 0P75WEST
# Samples Declustered
Coeff.
Variation
Top cut
threshold
(g/t)
# Samples Declustered
Coeff.
Variation
Top cut
threshold (g/t)s
AU 11 091 2.5 50 4 615 1.7 -
AS 5 365 1.7 - 2 440 1.1 -
AG 1 369 2.2 - 499 2.0 -
CU 1 443 2.0 - 505 1.5 -
C 96 0.7 - 229 0.8 -
SB 298 0.8 - 233 0.9 -
S 263 0.9 - 388 0.7 -

Table 14-10:     Outlier treatments in eastern domains

Elements 0P1EAST 0P75EAST
# Samples Declustered
Coeff.
Variation
Top cut
threshold
(g/t)
# Samples Declustered
Coeff.
Variation
Top cut
threshold (g/t)s
AU 3 408 3.3 40 386 1.1 -
AS 1 188 2.5 - 102 0.9 12 000
AG 192 1.2 5.5 11 0.7 -
CU 190 1.1 - 11 0.4 -
C 31 0.2 2 0 - -
SB 73 0.4 - 11 0.4 -
S 46 1.0 1.9 71 0.5 1.9

14.12

Drillhole Types

The drilling dataset used for the estimation consists of 677 drillholes including 1 RAB hole for 1.91 m composite length, 93 DD holes totalling 6107.8 m composite length and 583 RC holes totalling 14,497 m composite length. 256 RC holes are from grade control RC drillholes and are called RCGC. Figure 14-15 shows in red the location of the RCGC on the projected northing section.

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Figure 14-15:      AU samples collected in the RC drillholes, Red: RCGC, Blue: RC samples above 1,084 mRL and Green: remaining RC

The statistical comparison of drillhole types was undertaken using weighted composite data and outlier treatment.

Figure 14-16 compares the differences Au mean grade between RC above 1,084 mRL and RCGC. RCGC were not found in the eastern domains. The Au mean grade compares relatively well between both drilling types at the exception of MINOR domain. In MINOR domain, Figure 14-17 indicates that the Au distribution from RC samples above 1,084 mRL is not well defined after 24 g/t compares to the distribution within the RCGC samples. SRK considers that there is no bias between the RC and RCGC samples.

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Figure 14-16:      Comparison of weighted and top-cut Au mean grade between RC and RCGC above 1084 mRL

Figure 14-17:      MINOR – Au distribution – Left: RC above 1,084 m, Right: RCGC

Samples from RCGC drillhole type were combined with all the samples from RC drillhole type. From now on any reference to RC drillhole type includes RCGC drillhole type.

Figure 14-18 compares the Au mean grade between the DD and RC within each domain. However, the differences are less than 10% in most domains, the statistics fluctuate with the number of samples and are influenced by high grade values. Two third of the composite samples are from RC samples. Small domains with low number of samples tend to have more discrepancies between data types. SRK considers that there is no bias between the RC and DD samples.

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Figure 14-18:      Comparison of unweighted Au mean grade per drilling type

The samples from DD, RC and RCGC were used for the estimation.

14.13

Summary Statistics

This section compares the mean gold grade, declustered grade and top-cut grade within each domain (Table 14-11).

Table 14-11:      Comparison of mean gold grade (g/t) within each domain

Domains Gold mean grade (g/t)
Unweighted Weighted Top cut applied
0P1WEST 0.54 0.47 0.47
0P75WEST 4.61 3.11 3.11
0P1EAST 0.64 0.55 0.54
0P75EAST 3.71 3.77 3.77
MAIN 6.50 5.67 4.13
MINOR 5.72 3.26 3.20

14.14 Variography  

Spatial variability was studied for the two main elements gold and arsenic within each domain. The variography of gold and arsenic at Maud Creek was completed using the Isatis software, produced by Geovariance.

The choice of variogram directions were constrained by a combination of drilling data, geological continuity and domains. The structures observed were poor in most domains. To improve the variography study, the Gold grade was transformed into a Gaussian value within each domain while the arsenic was only transformed into a Gaussian value in domain 0P1EAST and 0P1WEST. The transformation was completed using a punctual Gaussian anamorphosis. The number of Hermite polynomial was adjusted according to the domain and element, varying between 30 and 70.

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The variogram was computed on the 1 m Gaussian composites within each domain separately. A model was fitted to the Gaussian variogram and back transformed to a model representing the composite data.

The parameters for the variogram model of the gold and arsenic for all the domains are given in Table 14-12 and Table 14-13. As expected, the Au element is fairly variable with a nugget that accounts for at least 44% of the total sill.

In MAIN domain, the best model is reasonably well described by two spherical structures and a nugget for both elements. The nugget for the Au variogram account for 74% of the total sill while the nugget for the arsenic variogram account for only 4%. The first structure is relatively short. Figure 14-19 shows the variogram fitting within the MAIN domain for gold and arsenic.

Table 14-12:      Variogram model characteristics for Au

Domains Rotation Plane Geologist Structures Sills Range
AZI DIP PITCH U V W
0P1WEST 0 67 161 Nugget 0.97
Spherical 1 0.44 30 30 5.6
0P75WEST 0 67 90 Nugget 16.42      
Spherical 1 4.97 35 35 2.8
Spherical 2 3.92 35 35 12
Spherical 3 2.31 90 35 12
0P1EAST 0 65 95 Nugget 2.7
Spherical 1 0.62 20 20 7
0P75EAST Omni-directional Nugget 14.23
Spherical 1 3.48 110 110 -
MAIN 0 65 116 Nugget 625.12      
Spherical 1 162.15 40 40 20
Spherical 2 60.73 160 100 20
MINOR 0 65 116 Nugget 13.35      
Spherical 1 9.46 40 40 20
Spherical 2 7.79 160 100 20

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Table 14-13:      Variogram model characteristics for Arsenic

Domains Rotation (degree)
Geologist Plane
Structures Sills (g/t) Range (m)
AZI DIP PITCH U V W
0P1WEST 0 100 -180 Nugget 266,079      
Spherical 1 456,877 40 20 10
Spherical 2 402,390 200 80 50
0P75WEST 0 67 180 Nugget 1 000 000      
Spherical 1 5 200 000 25 40 9
Spherical 2 2,000,000 140 40 40
0P1EAST 0 70 -180 Nugget 1 840 481      
Spherical 1 2 526 700 50 40 35
Spherical 2 2 001 163 100 40 35
0P75EAST Omni-directional Spherical 1 11 050 000 19 19 -
MAIN 0 85 170 Nugget 300,000      
Spherical 1 5 500 000 35 35 15
Spherical 2 2 000 000 100 80 15
MINOR Omni-directional Nugget 800 000
Spherical 1 4 000 000 40 40 40

Figure 14-19:      MAIN variogram fitting – left: Au, Right: As

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14.15 Estimation  

14.15.1

Gold and Arsenic

The Ordinary Kriging (OK) method was selected to estimate gold and arsenic within each domain. Domain boundaries are considered hard, with two exceptions.

Composites from the MAIN were allowed to be used in the estimation of the MINOR domain (soft boundary, although the domains are not in direct contact).

   

 

To address the potential under call in ounces where the 0.75 g/t Leapfrog grade shells did not connect, a two stage process was used to estimate the 0.1 g/t Au domain. Firstly the 0.1 g/t Au domain was estimated using a hard boundary constrained between 0.1 g/t Au to 0.75 g/t Au material. Secondly, and overriding the first estimation, a soft boundary was estimated using all of the material located outside the main vein (primary). A tight search ellipse orientated perpendicular to the dip direction was employed, to ensure only those blocks within the 0.1 g/t Au domain, directly adjacent to the main vein, were informed during the second estimation (The overall effect of this methodology was to increase the 0.1 g/t Au domain from an average grade of 0.48 g/t to 0.7 g/t Au, with some extra 140 koz at zero cut of.)

Identified outliers were replaced by the top cut when they exceeded a threshold distance of 15 m from the block centroid. Within 15 m the full uncut value was used.

The estimation was done using one neighbourhood search size specific to each domain. The maximum number of samples selected for the estimation is a compromise between the best local smoothed estimates, using many samples to optimise the slope of regression minimise negative weights and the best global grade and tonnage curve, de-smoothed, estimate using fewer samples. Details of the search ellipsoids for gold are shown in Table 14-14 and Table 14-15. Details of the search ellipsoids for arsenic are shown in Table 14-16. Note that the maximum number of samples is given by the number of sectors multiplied by the optimum number of sample per sector.

Table 14-14:      Neighbourhood parameters – Au estimated by OK

Domains 0P1WEST  0P75WEST 0P1EAST   0P75EAST   MAIN   MINOR  
Rotation            
X Angle 0 0 0 0 0 0
Y Angle 67 67 65 65 67 67
Z Angle 161 90 95 95 116 116
Search Ellipsoid            
X max 400 400 400 400 600 600
Y max 400 400 400 400 600 600
Z max 200 200 200 200 300 300
Number of sectors 8 8 10 8 1 1
Minimum samples 8 8 8 8 8 8
Optimum samples per sector 4 4 3 4 32 32
Minimum distance between data 0.7
Cut-off            
Threshold 50   40   200 50
Distance 15   15   15 15

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Table 14-15:      Neighbourhood parameters – Au estimated by OK – 0P1WEST soft boundary

Domains 0P1WEST (Soft boundary)
Rotation  
X Angle 0
Y Angle 67
Z Angle 161
Search Ellipsoid  
X max 400
Y max 400
Z max 10
Number of sectors 8
Minimum samples 4
Optimum samples 4
Minimum Number samples per line 2
Cut-off  
Threshold 50
Distance 15

Table 14-16:      Neighbourhood parameters – AS estimated by OK

Domains 0P1WEST 0P75WEST   0P1EAST   0P75EAST   MAIN   MINOR  
Rotation            
X Angle 0 0 0 0 0 0
Y Angle 100 67 70 65 85 67
Z Angle -180 90 -180 95 170 116
Search Ellipsoid            
X max 400 400 400 400 600 600
Y max 400 400 400 400 600 600
Z max 200 200 200 200 300 300
Number of sectors 8 4 4 8 4 1
Minimum samples 40 1 1 40 1 40
Optimum samples per sector 10 8 20 10 10 400
Minimum distance between data 0.7
Maximum Distance without sample 50 90
Cut-off            
Threshold       12000   10000
Distance       15   15

14.15.2

Other elements

The Inverse Distance of order 2 (ID2) method was selected to estimate the other elements in the western domains and MAIN while the Nearest Neighbourhood (NN) method was selected to estimate in the eastern domains and MINOR. ID2 and NN characteristics are as follow:

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Domain boundaries are considered hard, at the exception of the contact between MINOR and MAIN when estimating in the MINOR domain and the eastern domains.
     
  Identified outliers were replaced by the top cut if outside the threshold distance.
     
The search orientation and size were similar for all domains . The number of data used for the estimation is limited by the neighbourhood parameters. Details of the search ellipsoids are shown in Table 14-17.
     
Due to the limited number of data, a restriction on the distance without samples was applied meaning that not all the block cells were informed during the ID2 or NN estimation.

Table 14-17:      Neighbourhood parameters

Domains 0P1WEST 0P75WEST 0P1EAST 0P75EAST MAIN MINOR
Rotation  
X Angle 0
Y Angle 90
Z Angle 180
Search Ellipsoid  
X max 400
Y max 400
Z max 200
Number of sectors 4
Minimum samples 10
Optimum samples 10
Maximum Distance without sample 60 60 30 30 60 50

14.16

Density

There are three periods of density sampling over the life of the project as recorded in the density database supplied by Newmarket Gold. Around 1991 Places made 43 measurements from the WD series holes from 0.1 to 0.3 m lengths of core. Around 1995 Kalmet took 481 measurements from the MD series holes from 0.1 m to 0.3 m intervals of core. In 2011 Newmarket Gold took 2.145 measurements from 0.1 m lengths of core from the MC series holes. The Newmarket Gold measurements used some half core and some full core.

There are a total of 2.669 density measurements available, 740 of which are in mineralization (as defined by the estimation domains). Oxide states show significant differences in density (Table 14-18 and Table 14-19). There is no practical difference in densities within the fresh material for lithology or estimation domain (Table 14-20, Table 14-21and Table 14-22). The densities shown in Table 14-18 have been used to inform both the resource and waste model.

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Table 14-18:      Density statistics by oxide state all data

Oxide state Count Density
1 (completely oxidised) 3 2.11
2 (oxide / transition) 35 2.60
3 (transition) 53 2.72
4 (transition/ fresh) 39 2.77
5 (fresh) 2538 2.80

Table 14-19:      Density statistics by oxide state in all estimation domains

Oxide state Count Density
1 (completely oxidised) 2 2.10
2 (oxide / transition) 23 2.57
3 (transition) 16 2.65
4 (transition/ fresh) 3 2.77
5 (fresh) 696 2.79

Table 14-20:      Density statistics by estimation domain within all estimation domains in Fresh rock

Domain Count Density
0.1 halo 249 2.80
0.75 halo 327 2.80
Veins 120 2.78

Table 14-21:      Density statistics by lithology in Fresh rock all data

Lithology Count Density
Tuff 1016 2.79
Sediments 640 2.59
Minor Vein 30 2.76
Major Vein 90 2.78
Dolerite 762 2.85

Table 14-22:      Density statistics by lithology in estimation domains in in Fresh rock

Lithology Count Density
Tuff 387 2.81
Sediments 160 2.77
Minor Vein 31 2.76
Major Vein 90 2.78
Dolerite 29 2.82

14.17

Validation

Following the estimation, the final model was reviewed and validated. The estimation validation consists of statistical comparison, visual comparison and swath plots. The estimation validation of the OK includes also the review of the kriging output variables such as the regression slopes.

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The regression slopes of the Au estimates were relatively poor with a high number of values lower than 0.5. Figure 14-20 shows the results for the gold and arsenic within the MAIN domain.

Figure 14-20:      Regression slope histogram for Au estimates (left) and AS estimates (right)

The gold estimation by OK was re-run by increasing the maximum allowable number of data used in the search neighbourhood to 400. Increasing the maximum number of samples used for estimation in the neighbourhood definition improves the local block OK performance and results in lower slopes of regression overall (Figure 14-21) but over smooths the block distribution such the grades and tonnages at higher cut-offs are not realistic (Figure 14-23). For identification purposes in this section of the report, the check run OK is referred to Local OK while the gold estimation is called Global OK. Table 14-23 compares the global statistics between the composite top cut data and the two gold estimates. Global statistics showed reasonable good correlation at zero cut-off between the estimates and the data, slightly better for the Global OK estimates.

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Figure 14-21:      Regression slope histogram for Main Au Local check estimate

Table 14-23:      Comparison between declustered top cut Au composite and Au estimates per domain at zero cut-off

Domains Variable # Samples
/blocks
Minimum Maximum
(Cut max)
(Declustered)
Mean
%
Difference

0P1WEST
Composite 11 091 0.001 58.18 (50) 0.47  
Local OK 28 831 0.13 7.13 0.48 2.1
Global OK 28 831 0.11 7.14 0.46 -2.1

0P75WEST
Composite 4 615 0.05 282.83 3.11  
Local OK 8 650 0.99 25.70 3.33 7.1
Global OK 8 650 0.95 28.61 3.34 7.4

0P1EAST
Composite 3 408 0.01 76.07 (40 0.54  
Local OK 12 592 0.18 6.03 0.58 7.4
Global OK 12 592 0.11 7.55 0.54 0.0

0P75EAST
Composite 386 0.01 25.38 3.77  
Local OK 992 1.71 10.91 4.06 7.7
Global OK 992 1.29 11.13 3.98 5.6

MAIN
Composite 1 114 0.005 619.43 (50) 4.13  
Local OK 7 136 0.63 91.87 4.32 4.6
Global OK 7 136 0.06 103.36 4.30 4.1

MINOR
Composite 224 0.01 79.67 (50) 3.20  
Local OK 1 880 0.228 24.874 3.65 14.1
Global OK 1 880 1.499 22.599 3.96 23.8

Within the MAIN domain, 40 m swath plot in X, Y and Z directions showed a reasonably good correlation within area well sampled (Figure 14-22). In areas with many composite samples, the OK has smoothed the grade while in area with limited composite samples the OK tends to overestimate. The Local OK is mostly over-estimating the grade compared to the Global OK approach.

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Figure 14-22:      MAIN – Au Global OK – Swath plots

The gold estimations were then compared against the theoretical block distribution defined by a change of support calculation on the Gaussian anamorphosis of the sample distribution. Figure 14-23 compares the estimates with the theoretical block distribution for gold in the MAIN domain. The Local OK displays a sharp tonnage and grade gradient between gold grade cut-off of 3g/t and 4g/t. The grade tonnage curve for Global OK shows a better correlation with the theoretical block distribution. The metal tends to be consistently under-estimated compared to the theoretical block distribution. The theoretical block distribution is considered to predict tonnage, grade and metal closer to the expected global recovery. Therefore the Global OK is considered a better estimate at this stage of the mining project. All further references to the gold estimate in this report refer to the Global OK Au estimate.

From Left to right: Tonnage, Grade, Metal, Green = Global OK, Black= Local OK and Red= Theoretical Block distribution.

Figure 14-23:      MAIN –gold – Grade tonnage comparison Reasonable Prospects of economic extraction

In assessing the criteria for reasonable prospects of economic extraction both open pit and underground scenarios were considered. With respect the scattered lower grade mineralization contained within the eastern domains (0p1E and 0p75E) near surface a simple pit optimisation using the optimistic parameters in Table 14-21 at twice the current gold spot price did not generate a pit of practical size on the eastern domains. All material in the eastern domains is not considered to have reasonable prospects of economic extraction and does not appear in the Mineral Resource.

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Table 14-24:      Pit optimisation parameters for evaluation of Mineral Resource classification

Parameter Value
Gold Price (twice current spot) AUD2 830 /Oz
Processing Cost (Whittle) AUD58.4 /t mill feed
Recovery Oxide and Transition 81%
Recovery Fresh 90%
Overall wall slope angles 34 degrees

With respect to the underground potential the grade is reasonably consistent down to approximately 650 mRL below which it drops significantly. All material below 650 mRL is not considered to have reasonable prospects of economic extraction and does not appear in the Resource.

14.18

Classification

Classification is based on a combination of drill spacing, geological interpretation confidence, proximity to the previously mined open pit, reasonable prospects of economic extraction and grade. The classification areas are coherent zones and do not contain isolated blocks of lower classifications within them.

Measured is defined by the main, minor and 0p75W domains above 950 mRL and between plunging north and south boundaries that approximate the limits of the closer drill spacing (Figure 14-24). Indicated (Figure 14-25) is defined by:

  1.

The 0p1 domain external to the Measured above 950 mRL; and

     
  2.

The approximate limits of the 20 m by 20 m drilling.

The inferred is the remaining material above 650 mRL within approximately 50 m of drilling and the areas with low geological confidence in the orientations of the controls on mineralization and assumed structures (Figure 14-26).

Figure 14-24:      Long section facing west displaying measured blocks for all domains

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Figure 14-25:      Long section facing west displaying indicated blocks for all domains

Figure 14-26:      Long section facing west displaying inferred blocks for all domains

14.19 Mineral Resource Tonnage and Grade  

Given the Mineral Resource is amenable to open cut and underground mining a split set of cut-offs is used for reporting. An elevation limit of 950 mRL has been used for the depth limit for open cut reporting as this is 50 m below the simplistic whittle pit optimisation depths generated with optimistic revenues.

A cut-off of 0.5 g/t Au is defined as the base case (Table 14-25), a comparison at 1.0 g/t Au is also included (Table 14-26). The open pit Mineral Resource is exclusive of the underground Mineral Resource. The Mineral Resources are stated here for the Maud Creek deposit with an effective date of 15 March 2016.

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Table 14-25:      Open pit Mineral Resource above 950 mRL at 0.5 g/t Au cut-off base case

Classification Tonnage
(Kt)
Grade
(Au g/t)
Contained Metal
(KOz Au)
Measured 1 067 5.59 192
Indicated 1 100 2.14 76
Measured and Indicated 2 167 3.84 268
Inferred 531 1.41 24

Table 14-26:      Open Pit Mineral Resource above 950 mRL at 1.0 g/t Au cut-off – for comparison only

Classification Tonnage
(Kt)
Grade
(Au g/t)
Contained Metal
(KOz Au)
Measured 1 067 5.59 192
Indicated 1 100 2.14 76
Measured and Indicated 2 167 3.84 268
Inferred 232 2.36 18

The underground Mineral Resource consists only of material below 950 mRL. The base case is stated at 1.5 g/t Au cut-off (Table 14-27). A comparison at 2.0 g/t Au cut-off is provided in Table 14-28. The underground Mineral Resource is exclusive of the open pit mineral resource.

Table 14-27:      Underground Mineral Resource below 950 mRL at 1.5 g/t Au cut-off – base case

Mineral Resource
Classification
Tonnage
(Kt)
Grade
(Au g/t)
Contained Metal
(KOz Au)
Measured - - -
Indicated 4 326 3.28 456
Measured and Indicated 4 326 3.28 456
Inferred 1 451 2.65 124

Table 14-28:      Underground Mineral Resource below 950 mRL at 2.0 g/t Au cut-off – for comparison only

Mineral Resource
Classification
Tonnage
(Kt)
Grade
(Au g/t)
Contained Metal
(KOz Au)
Measured - - -
Indicated 3 490 3.65 410
Measured and Indicated 3 490 3.65 410
Inferred 1 026 3.04 100

It should be pointed out the mineral resource estimate is categorized as Measured, Indicated and Inferred as defined by the CIM guidelines for resource reporting. Mineral resources do not demonstrate economic viability, and there is no certainty that these mineral resources will be converted into mineable reserves once economic considerations are applied. The Measured, Indicated and Inferred mineral resource estimate has been prepared in compliance with the standards of NI 43 – 101 by Danny Kentwell, FAusIMM.

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15

Mineral Reserve Estimate

In July of 2013 Mineral Reserves were estimated based on the Mineral Resource and processing route at that time.

Due to the changes in the Mineral Resource model, evaluation of a revised processing route and the potential change in the saleable product from gold Dore to Gold-rich concentrate there is no Mineral Reserve to report at this time.

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16

Mining Methods

Based on the geological review there is potential for the deposit to be exploited by conventional open pit and underground mining methods.

The final underground option will be determined by the availability of pastefill supplied from an on-site processing plant. If mineralization is trucked to Union Reefs, a cemented aggregate fill method will be utilised. This will dictate the selected mining method and extraction sequence.

The Union Reefs option provides a processing stream and value to be realised from the for the oxide component of the Mineral Resource that is not available under the stand-alone plant option and will influence the mine design.

16.1

Geotechnical

In conjunction with the Mineral Resource review, a comprehensive review of the available data and analysis to determine open pit and underground geotechnical design guidelines. The following summarises the work undertaken and findings presented in Appendix A.

An assessment of overall slope angles and underground mining parameters has been undertaken using geological and geotechnical drilling data supplied by the client. The analysis provides good early-stage design guidelines of the geotechnical properties of the rock mass. The typical geotechnical conditions on site can be summarised as follows:

The Hangingwall Tuffs are typically massive but may be locally bedded. Hangingwall tuffs are also affected by the numerous shears present in the Hangingwall, resulting in reduced strength, increased fracture frequency and graphitic and/or chloritic alteration of the rock mass.

   

 

The Footwall Sediments consist of low to medium strength thinly bedded or laminated mudstone and siltstone, and medium to thickly bedded sandstone. Zones of intense shearing with chlorite and graphite alteration occurring in the 5 to 10 meters below the mineralized zone where the sediments are commonly black, highly graphitic and/or chloritic, very weak and fissile.

   

 

The competency of the mineralized zone can be expected to be variable with competent, partially silicified mineralized zones separated by zones of intensely sheared rock .

   

 

The distribution of the various fault configurations is not understood at this stage and this should be one of the main focus for subsequent field investigations.


16.1.1

Level of Confidence

The perceived level of confidence in the different data streams available were rated subjectively, as shown in Table 16-1. A five-point rating scale of Very Low – Low – Moderate – High – Very High has been used.

At this stage of the project development, at least a Low rating should be expected for all items, with all items requiring further investigation. As the study progresses, the confidence levels will improve. Aspects for which no data is currently available or represent a key concern have been flagged as Very Low.

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Table 16-1:      Qualitative risk assessment of study components.

Data Confidence level
Empirical: rock mass characterisation Low
Structural: major structures Low to Moderate
Structural; rock mass structures Low
Rock mass strengths Very low
Rock material strengths Very low
Groundwater conditions Very low
Slope angle recommendations Low

16.1.2

Project Risks and Opportunities

The main project risks relate to the absence of data required for the development of design parameters. The areas where data is required are:

Minor geological structures – additional pit mapping/photogrammetry and diamond drilling with geotechnical and structural logging are necessary to overcome the current orientation bias and increase the data density throughout the study area.

   

 

Rock mass characterisation and strength – further drilling and evaluation is needed to improve confidence in the rock mass properties of all the domains, in particular within the major fault zones.

   

 

Rock material strength – a representative laboratory testing programme needs to form part of the next stage of investigation. This should allow the identification and evaluation of design mechanical properties which can be used for numerical analysis of pit designs at the next stage of the study.

   

 

Improved understanding of the project groundwater conditions is needed to understand the interaction between the site hydrogeology and the pit walls.


16.1.3

Review of Geotechnical data

Geotechnical data was collected from drill holes drilled between 1990 and 2011. The geotechnical data was collected from the orebody and the rock mass adjacent to the orebody, with the more recent drilling collecting data throughout the wider area. Recent drilling was also more likely to include a wider range of geotechnical parameters. Sediments located in the footwall of the deposit are not well represented within the overall geotechnical dataset. Additional geotechnical drillhole data will be required for future studies. Geotechnical laboratory testing will be required from samples recovered from future drilling.

16.1.4

Rock mass characterisation

Geotechnical domains have been determined for both the open pit and underground areas at the Maud Creek project based on geological units, weathering, structural setting and rock mass quality. Figure 16-1 shows the different geotechnical domains selected for the open pit analysis. For underground component of the study, stope stability assessments considered domains within the orebody and 10 m either side of the orebody only, whilst broader domains were used to characterise the rock mass for ground support requirements.

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Figure 16-1:      Plan view of existing pit showing the geotechnical domains used for rock mass characterisation

16.1.5

Underground Design Considerations

Stope Design

Maximum allowable Hydraulic Radii (HR) have been determined for the different geotechnical domains based on empirical analysis completed. The surface controlling stope sizes will be the footwall as it will be exposed as the footwall across most of the production areas. A summary of recommended stope sizes is shown in Table 16-2.

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Table 16-2:      Empirical Stope Design Summary

Domain Stope
Surface
N’      Hydraulic Radius (HR) Strike Length
(m)
Stope
Height (m)
Unsupported Supported
Sediments
FW- A
FW 0.9 – 3.27 2.3 – 3.8 6.5 – 8.0 10 25
Sediments
FW – B
FW 1.83 – 18.58 3.0 – 7.1 7.2 – 11.3 15 25
Sediments
FW- C
FW 0.6 – 30.39 2.0 – 8.4 6.1 – 12.5 20 - 30 25
ORE Backs 1.66 – 26.89 2.9 – 8.1 7.0 – 12.1 15 10 - 35
Tuff - HgW HgW 9.82 – 79.93 5.5 – 12.0 8.0 – 15.0   25

Stope dilution may be controlled by designing stable beams between the stope footwall and the shear zones, reduction in stope strike spans or cable bolt reinforcement of stope hangingwall or footwall.

Ground Support Design

Ground support requirements were assessed using empirical methods and are summarised in Table 16-3.

Table 16-3:      Ground Support Requirements

Geotechnical Domain Excavation Type Minimum Ground Support Requirements (Range)

Footwall Sediments

Main Decline

2.4 m Galvanised Friction bolts (46 mm diameter Split Sets)

Bolt spacing 1.0 m – 1.6 m

Surface support ranges from Fibrecrete (75 mm) to Mesh covering backs and walls to grade line

Footwall Sediments

Level Accesses

2.4 m Galvanised Friction bolts (46 mm diameter Split Sets)

Bolt spacing 1.0 m – 1.6 m

Surface support ranges from Fibrecrete (75 mm) to Mesh covering backs and walls to grade line

Vein (Orebody)

Ore Development

2.4m Galvanised Friction bolts (46 mm diameter Split Sets)

Bolt Spacing 1.4 m – 1.8 m

Mesh support covering backs extending to grade line

Hanging wall Tuff
Development

Other Capital
Development

2.4 m Galvanised Friction bolts (46 mm diameter Split Sets)

Bolt Spacing 1.5 m

Fibrecrete (40 mm) to Mesh covering backs and walls to grade line

Crown Pillar assessment

Caving and potential subsidence was identified as a possible hazard during this study. An empirical assessment was completed to determine the stability of the crown pillar located between the top of the stoping and the base of the proposed open pit.

A number of mitigation measures are available and include limiting the number of stopes open at a given time and limiting stope widths to 15 m with a 10 m thick crown pillar. Alternately increasing the crown pillar thickness would allow for wider stopes in the transverse stoping area. Cemented backfill is recommended for the stoping area located immediately below the open pit and highly weathered zones.

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Further investigation of the crown pillar coupled with numerical modelling would allow optimisation of the crown pillar thickness.

Backfill

Pastefill will be required at Maud Creek in areas where the mining method and extraction sequence dictate such as any area where a top down sequence is adopted and for transverse stoping.

A detailed backfill study should be completed during the prefeasibility study stage. Factors to be considered during the pastefill study include the suitability of tailings and binders to form pastefill, the required strengths required for stable fill exposures and curing times.

16.1.6

Open pit Slope Stability Analysis

The preliminary open pit design recommendations have been based upon the geotechnical domains summarised in Table 16-4.

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Table 16-4:      Preliminary Slope Design Parameters

Domain Geotechnical
Safety Berm
Width
(m)
Ramp
Width
(m)
Bench
Height
Cases
(m)
Depth
Range
(m)
Bench
Height
(m)
Bench
Face
Angle
Bench
Stack
Height
(m)
Bench
Stack
Angle
Inter-
Ramp
Angle
Limiting
Overall
Slope
Angle
Hanging
Wall
N/A 24 All 0 – 25 12.5 55° 25 46° 38°
12.5 25 - 50 12.5 55° 25 46° 38°
50 - 100 60° 50  45° 33° 34°
50 - 150 75 44° 35° 36°
50 - 250 36° 36°
25 25 - 50 25 55° 25  55° 46°  
50 - 100 60°

50  54° 41° 43°
50 - 150 75 53° 40° 42°
50 - 250 42° 43°
Southern 15 24 All 0 – 25 12.5 55° 25 46° 38°
12.5 25 - 50 12.5 55° 25 46° 38°
50 - 100 50  42° 34° 35°
50 - 150 75 41° 35° 36°
50 - 250 34° 35°
25 25 - 50 25 55° 25 55° 46°
50 - 100 50  50° 38° 39°
50 - 150 75 48° 40° 41°
50 - 250
39° 40°
Footwall
and
GCFZ
15 N/A All 0 – 25 12.5 55° 25 46° 38°
12.5 25 – 50 12.5 55° 25 46° 38°
50 – 100 50  42° 35° 33°
50 – 150 75 41° 36° 35°
50 – 250 34°
25 25 – 50 25 55° 25 55° 46°
50 – 100 50  50° 40° 37°
50 – 150 75 48° 41° 40°
50 – 250 39°

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17

Recovery Methods

Maud Creek sulphide (fresh) mineralization are refractory to direct cyanidation but respond well to flotation. There is significant but variable gravity-recoverable gold throughout. The gold is largely locked in sulphides and requires an oxidation process for downstream processing. Furthermore, the presence of preg-robbing carbonaceous material renders the ore ‘double-refractory’.

The Maud Creek mineralization includes oxide, transitional and sulphide (fresh) material. The oxide material is generally amenable to direct cyanidation and could be treated through the existing Union Reef mill, the transitional material is less amenable to conventional cyanide leaching (variable). Alternatively, this transitional material may be treatable by flotation. Flotation recovery could be further improved following a sulphidisation pre-treatment process. Controlled potential sulphidisation (CPS) technology has been successfully applied on several gold and copper/gold mines, and may be beneficial at Maud Creek. Some preliminary testing would be worthwhile.

The main body of metallurgical testing and other development work was conducted from 1994 - 98 for Kalmet Resources NL and Kilkenny Gold NL. Some minor work was done for Harmony Gold Operations Ltd in 2003, and then further testing and piloting were undertaken for Terra Gold Mining Ltd in 2006. Along the way there have been several engineering and consulting reports in support of the metallurgical development of the project. In particular, John W MacIntyre and Associates Pty Ltd supervised the earlier testing, starting in late 1996 and culminating in a detailed report: Metallurgical Evaluation of the Maud Creek Project in September 1998.

There is also an option to modify the Union Reef mill to allow sulphide treatment (campaign treatment). This plant currently treats oxide material from the Cosmo underground mine by gravity recovery and CIL. It has excess crushing capacity and two underutilised grinding mills at the current throughput rates on a hard underground ore. It could treat >500 ktpa of feed. The addition of a flotation plant would enable parallel treatment of sulphides. Though conceptually possible, there is an economic trade-off due to the trucking distance.

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18

Project Infrastructure


18.1

Site Access

A suitable access road connecting the Maud Creek project site to the Stuart Highway is required to facilitate access. Three options are illustrated in

Figure 18-1.

The existing unsealed access road to the Maud Creek site into Katherine, heading south from the mine was used during previous mine operations to haul material from site. The access road is linked to the Stuart Highway by Ross Road, a sealed public road which is regularly used by horticultural landholders in the area. The access road crossed Gold Creek just south of the existing mine pit and during the wet season, the road is often impassable and experiences significant damage.

The existing access track would need to be upgraded and have a crossing installed over Gold Creek to enable transport operations to continue during the wet season both into and from site. This route has been approved by the Northern Territory Government as a Right of Way and General Service easement. A second approved Right of Way and General Service easement runs north from the Mine to Gorge Road. This provides the site with a second access point to Katherine.

Figure 18-1:      Site Access Roads

18.2

Power Supply

The town of Katherine is linked by a 132 kV transmission line to the Darwin-Katherine Interconnected System which supplies regulated electricity to the region. The transmission and distribution of power in the Northern Territory is the responsibility of the government owned Power and Water Corporation (PWC).

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18.3

Water

The raw water supply sources for the site include the:

  Groundwater seepage into the mine pit; and
     
Potential runoff over catchments associated with the mine pit, waste rock dump, tailings facility infrastructure area, and sedimentation ponds.

The total catchment area reporting to the Maud Creek site of 25.74 km2 was divided into three major sub-catchments, as shown in Figure 18-2. The areas of each sub-catchment are outlined in Table 18-1.

Table 18-1:      Areal extent of catchment areas

ID Area (km2)
1 0.616
2 1.337
3 23.786

Figure 18-2:      Catchment areas in Maud Creek site

Water management of the site will need to be incorporated as a key factor in any future plans, to keep clean water clean and direct the contact water to appropriate containment system.

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19

Market Studies and Contracts

This section is not applicable at this time.

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20

Environmental Studies, Permitting, and Social or Community Impact


20.1

Environment and Social Aspects and Impacts


20.1.1

Social and Economic Context

The Maud Creek Project lies within the Town of Katherine local government area (LGA 72200), which occupies an area of some 7417 square kilometers (Figure 20-1) and within the broader Katherine region (336,674 km2,Figure 20-2). The traditional owners of the land, the Jawoyn people, have occupied the Katherine region for thousands of years.

Figure 20-1:      Katherine municipality local government area (ABS, 2011 census)

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Figure 20-2:      Katherine region (shown in darker tan)

Source: Katherine Land Use Plan, 2014

The town of Katherine is the fourth largest population centre in the Northern Territory, after Darwin, Palmerston and Alice Springs. The town was formally gazetted in 1926. At the 2011 census, the population of the Katherine municipality was estimated at slightly under 11,000 people, of whom about 24% identified as indigenous (Aboriginal or Torres Strait Islander). In younger age groups, the percentage of Aboriginal people is higher (Figure 20-3). Compared to the Australian population as a whole, the Katherine community is relatively young, with a median age of 31. The Katherine population is characterised by a high degree of mobility, with a relatively large proportion of the population changing place of residence (either within the region or interstate) between consecutive censuses.

Figure 20-3:      Katherine LGA population, by age group (2011 census)

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Katherine is an important regional centre, providing government services (health, education, transport, communications and business development functions) to the wider region. A range of basic utilities, services and community infrastructure is available, including:

  A 60- bed hospital; various community and private medical clinics;
     
A range of public and private education and training providers from pre-primary to university level; an airport (shared with the RAAF Base Tindal);
     
  Police, fire and emergency services;
     
  Water supply, sewerage and waste disposal facilities, and
     
  Cultural, sport and outdoor leisure facilities, such as the Nitmuluk National Park.

The modern economy of the Katherine region has traditionally been dominated by the agricultural and pastoral sectors, but in the past few decades mining, public administration and safety (including defence), tourism and construction (largely related to major resource and infrastructure projects) have also been important contributors to the regional economy. Although mining has been by far the greatest contributor to gross regional product in the past few years (Figure 20-4), it is not a major employer in the Katherine municipality (Figure 20-5).

Unemployment and labour force participations rates in Katherine are, on average, similar to those in Australia as a whole, although unemployment among young people (those aged 24 years and below) is conspicuously higher than average unemployment rates in the general Australian community (Figure 20-6 and Figure 20-7). In 2011 about 41% of the Katherine population aged 15 and over had some type of post-secondary school qualification, mostly at Certificate or Diploma levels.

Figure 20-4:      Gross regional product, by industry – Katherine NT (2012)

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Figure 20-5:      Employment by industry sector – Katherine LGA (ABS data)

Figure 20-6:      Unemployment rates – Katherine LGA (ABS data)

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Figure 20-7:      Labour force participation rate, Katherine LGA (ABS data)

Median weekly household income reported at the last census in the Katherine LGA was approximately AUD1,424. Median weekly rental and mortgage repayments at the 2011 census were AUD200 and AUD433, respectively. Current median rentals range from around AUD320/ week for a unit to AUD450/ week for a house (https://www.realestate.com.au/neighbourhoods/katherine-0850-nt, accessed 14 Jan 2016). Slightly over 50% of the Katherine population lives in rented accommodation.

Approximately 3860 private dwellings exist in the Katherine municipality, of which about 8.5% (329 dwellings) were unoccupied at the time of the 2011 census. Housing types include single dwellings (~58%); semi-detached structures, townhouses, units and apartments (~17%); and a range of other accommodation, such as caravans, tents and improvised accommodation. Northern Territory government budget forecasts predict continuing strong demand for housing in Katherine, with relatively low vacancy rates (NT Government budget, 2013-2014). The Town of Katherine Land Use Plan (2014), anticipates the need for an additional 81ha of urban residential land and 180 ha of rural lifestyle lots (rural, rural living and rural residential) to accommodate residential development to beyond 2026. In the main, the areas being considered for additional residential development lie to the south and northwest of the Katherine urban centre and are unlikely to encroach on the Maud Creek project area, however, evolving land use patterns may need to be taken into account as part of transport planning for the project.

The Katherine community comprises a number of different groups, including non-resident tourists and seasonal workers, part-time residents (FIFO workers), employees of the RAAF Base Tindal and their families, and full time residents. Stakeholder engagement for the Maud Creek project will need to take account of these diverse stakeholders, as well as the needs and expectations of Aboriginal traditional owners and their representative bodies, along with the requirements of government stakeholders at local, regional and territory level.

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20.1.2

Surface Water

The Maud Creek Project area is traversed by Gold Creek, an ephemeral tributary of Maud Creek, which flows into the Katherine River upstream of the water supply extraction point for the Katherine Township. The distance between the proposed mine operations area and the confluence with the Katherine River is approximately 10.5 km. The Katherine municipality sources its water from a combination of groundwater bore and the Katherine River. The town water supply use of water from the Katherine River is sometimes constrained by the high turbidity in surface watercourses during the wet season. Neither Maud Creek nor Gold Creek contribute significant (or possibly, any) flow to the Katherine River during the dry season, although remnant pools persist along the watercourses throughout the dry season. Hydrological modelling conducted in connection with Terra Gold’s Maud Creek proposal (discussed in Section 4.5) concluded that the channels of both Gold Creek and Maud Creek would be likely to experience over bank flows on several occasions in every wet season. Modelling conducted by SRK is consistent with the earlier assessment (Figure 20-8).

Surface water quality in the project area has been characterised to a limited extent. Dry season sampling conducted by previous tenement holders between 1968 and 1998 recorded generally good quality water, of near-neutral pH, relatively low salinity and low concentrations of most trace metals (URS, 2008). The EIS prepared in connection with the Terra Gold Maud Creek proposal suggested that surface waters in the project area may occasionally show elevated concentrations of copper or lead, presumably as a result of mineralization in the catchment, however additional monitoring would be required to demonstrate this convincingly.

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Figure 20-8:      Estimated flood extents in absence of engineering controls

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20.1.3

Groundwater

Groundwater beneath the Maud Creek site occurs primarily within fractured rock (Tollis formation) aquifers and unconsolidated alluvial sediments. Groundwater recharge is expected to occur from infiltration of rainfall, especially in areas of outcropping quartz/quartz breccia. Some recharge would also occur seasonally through the bases of watercourses. The proposed mine operations area lies to the north of the Tindall limestone formation, upon which the Town of Katherine, and a range of local enterprises, rely for water supply (Figure 20-9).

Figure 20-9:      Groundwater aquifers near Katherine (from Katherine land use plan, 2014)

The groundwater table in the Maud Creek area generally lies at shallow depth (between 1 and 6 m below ground surface, corresponding to an average elevation of approximately 141m AHD). The seasonal variation in the depth to water is expected to be in the range of 2 to 4 m. The general direction of groundwater flow is to the northeast, away from the Tindal limestone aquifer.

Baseline monitoring conducted in connection with Terra Gold’s EIS (URS, 2008) reported the dry season groundwater level at the confluence of Gold and Maud Creeks to be approximately 6.9 m below ground level (approximately 4 m below the creek bed). Limited if any contribution of groundwater to local stream flow is likely during the dry season, as the groundwater levels will generally lie below the beds of the watercourses.

Groundwater quality is typically fresh, with total dissolved solids concentrations below 600 mg/L. Groundwater chemistry is dominated by calcium and magnesium bicarbonates and the groundwater pH is slightly alkaline, with an average pH around pH8. Information in Terra Gold’s Maud Creek EIS suggested that the local groundwater may have naturally elevated concentrations of arsenic and/or selenium; however, this conclusion appears to be based on limited water quality testing and would need to be confirmed through further groundwater monitoring.

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20.1.4

Native Flora and Vegetation

As part of the EIS, Terra Gold commissioned field surveys of flora and vegetation for its Maud Creek proposal. The botanical investigations, which were carried out in March and April 2007, identified 153 plant species of which 19 were introduced weeds. None of the 153 plant species recorded during the 2007 were listed under the Commonwealth Environment Protection and Biodiversity Conservation Act 1999 (EPBC Act) as vulnerable, threatened, rare or endangered. One species, Tephrosia humifusa, recorded at 5 sites during the field surveys, is listed as Near Threatened (NT) under the Territory Parks and Wildlife Conservation Act 2006. The 2007 survey reported noted that Tephrosia humifusa appeared to be common locally, occurring at over one quarter of the floristic study sites and in a range of different habitat types. At the time of the Terra Gold EIS, Tephrosia humifusa was known to occur at 5 different locations in the Northern Territory. The Maud Creek survey records represented a range extension of the species. Current government records shown 19 occurrences of Tephrosia humifusa in the Northern Territory (not including at Maud Creek, referFigure 20-10). The plant remains on the Northern Territory endangered species list (2012).

Figure 20-10:      Distribution of Tephrosia humifusa (Atlas of Living Australia, accessed 15/01/2016)

Northern Territory government databases show a number of protected flora species in area to the north of the Maud Creek project area, but none within the mine tenement (data from http://nrmaps.nt.gov.au/) (Figure 20-11).

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Figure 20-11:      Conservation significant flora

(data from http://nrmaps.nt.gov.au/)

Ten distinct vegetation communities were mapped in the Maud Creek Project area during the 2007 surveys (Figure 20-12). The vegetation communities included open forest, woodland and open woodland assemblages. None of the vegetation types was considered of special conservation significance: all occur widely outside the project area and none is thought to support specially protected flora or fauna species. No vegetation communities protected under the EPBC Act were recorded during the 2007 surveys.

There is currently no mechanism for listing Threatened Ecological Communities under NT legislation. However, one threatened ecological community that occurs in the general project locality is listed as Endangered under the Commonwealth EPBC Act. This is the Arnhem Plateau Sandstone Scrubland Complex, which has had protected status since 2011. The Arnhem Plateau Sandstone Scrubland Complex is known to occur in parts of the Nitmiluk (Katherine Gorge) National Park. It is not likely to occur within the Maud Creek mine tenement. This would need to be confirmed as part of future project assessment and permitting.

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Figure 20-12:      Vegetation communities

(From Crawford & Metcalf, 2007)

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20.1.5

Native Fauna and Habitats

Terrestrial fauna habitats

The most recent field fauna surveys of the Maud Creek project area were conducted in April and May 2007. Previous surveys had been carried out in 1994, 1996 and 1997 (EMS, 2007). The 2007 surveys recorded a range of habitat types, including open woodland, grassland and riparian habitats. Distinctive and geographically limited habitat types underlain by limestone karst or limestone outcrops were noted to the west of the mine tenement, in an area that had been proposed for an access road. No limestone outcrop was recorded within the mine tenement itself. Overall, the fauna habitats within the proposed mine operations areas were not characterised by significant biodiversity. Many areas had been disturbed by past mining activities, grazing pressure, weed incursion and disturbance from swamp buffalo, cattle, pigs, feral donkey and fires. Riparian habitats, which are not locally extensive, have been degraded by large numbers of swamp buffalo. Habitats within the mine tenement are unlikely to support significant biodiversity. None of the habitats present in the project are likely to provide significant habitat for water birds or shore birds. The limestone outcrop habitats to the west and south of the mine operations areas are less heavily impacted and may act as locally significant refugia and areas that support conservation significant wildlife, including bats and a variety of invertebrate species such as land snails.

Terrestrial vertebrate fauna

A total of 144 native and six introduced terrestrial vertebrate species were recorded during the 2007 surveys. These included 30 mammals, 91 birds, 18 reptiles, and 11 amphibian species. A number of varanid (monitor lizard) species observed during previous surveys of the area were absent during the 2007 surveys. This may be the result of predation and displacement by cane toads, which were introduced to the area in about 2001. Cane toads were the most common amphibian observed during the surveys.

The majority of fauna recorded or expected to occur in the Maud Creek Project area are widespread in northern Australia. Two bird species of conservation significance – the red goshawk (Erythrotriorchis radiatus) and the Australian bustard (Ardeotis australis) have been observed within the mine tenement. Both of these are relatively widespread within the NT. Six species that are considered Near Threatened under the according to the Territory Parks and Wildlife Conservation Act 2006 were also recorded in or very near to the proposed mine operations area ((data from http://nrmaps.nt.gov.au/) Figure 20-13). They included: the bush stone-curlew (Burhinus grallarius), hooded parrot (Psephotus dissimilis), grey falcon (Falco hypolelucos), northern nailtail wallaby (Onychogalea unguifera), Arnhem sheathtail bat (Taphozous kapalgensis), orange leafnosed bat (Taphozous georgianus) and western chestnut mouse (Pseudomys nanus). No significant migratory bird species were recorded on or near the mining tenement. (The EPBC-listed rainbow bee-eater (Merops ornatus) was recorded, but this species is very common and is unlikely to be significantly impacted.)

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Figure 20-13:      Conservation significant fauna

(Data from http://nrmaps.nt.gov.au/)

Terrestrial invertebrate fauna

Land snails were collected and identified at all the limestone outcrop areas surveyed to the west of the mine tenement (along the access road alignment previously proposed by Terra Gold). Some of the snails observed in the limestone outcrop habitats (Xanthomelon sp) were also collected from open forest habitats in the project area. This species is thought to be relatively widespread, but occurs at low densities within its range (EMS, 2007). The most common snail species collected was Torresitrachia weaberana, a moderately commonly species known to occur in areas between Katherine and Kununurra. One of the snail species collected was an undescribed genus and undescribed species that is known to be limited in range to limestone karst in the Tindal – Cutta Cutta region. This species was considered to be of regional conservation significance. Recent technical studies on snails in the Katherine region have proposed that endemic land snails can be used as bio-indicators for the health of some ecological assemblages and may themselves be under threat of extinction (Braby et al, 2011, Willan et al, 2009).

Aquatic fauna and habitats

Fifteen species of freshwater fishes were recorded in Gold Creek and Maud Creek during the 2007 field surveys. The western rainbowfish (Melanotaenia australis) was the most abundant species in samples and was the only species present at all sites. Other common species included the banded grunter (Amniataba percoides), spangled grunter (Leiopotherapon unicolor) and northern trout gudgeon (Mogurnda mogurnda) (URS, 2008).

No listed threatened aquatic species were recorded in recent or previous field surveys in the project area. Although the EPBC Act lists the freshwater sawfish (Pristis microdon) as a vulnerable species that could potentially occur in the vicinity of the project area, no suitable habitat for this species occurs in proximity to the proposed mine operations it. Neither is the species known to occur in the Katherine River.

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None of the freshwater fishes recorded during past surveys of the Maud Creek project area are listed as threatened in the NT or under Commonwealth legislation. The species occurring Maud Creek are all common and widespread forms that are well adapted to the variable instream conditions that characterise the system Aquatic habitat quality along Gold Creek was described as ‘significantly impaired’ by trampling and other disturbance by swamp buffalo, donkeys, feral pigs and cattle. Neither Maud Creek nor Gold Creek are considered likely to support high levels of biodiversity in the project area, although the riparian habitats may afford a level of shelter and temporary feeding and growing habitat. The watercourse may provide a migration path for aquatic species during the wet season (URS, 2008).

20.1.6

Air Quality, Noise & Vibration

The proposed mine operations area is located at least 5 kilometers from the nearest ‘sensitive receptor’ (residential premises, for example). Providing conventional environmental controls on dust and noise emissions are implemented during construction and operations, it is unlikely that significant impacts on environmental amenity will arise.

There is some risk that offsite traffic associated with the project will give rise to public concern. Arrangements for controlling impacts of product haulage, workforce traffic and vehicle movements for transport of fuel and reagents will need to be described as part of the project’s environmental impact assessment.

20.1.7

Conservation Areas

The closest conservation area to ML30260 is the Nitmiluk (Katherine Gorge) National Park, owned by the Jawoyn Aboriginal Land Trust and jointly managed with the Parks and Wildlife Commission (Figure 20-14). The park is notable for its scenery, cultural values and ecological integrity. The proposed mining operations are unlikely to give rise to direct impacts on Nitmiluk National Park, although there is some potential for indirect impacts on the park, particularly if the Maud Creek Project includes the upgrade and use of the general service easement running north from the mine to Gorge Road. Increased use of the northern access route has the potential to increase the risk of spreading weeds and/or pathogens and may exacerbate the adverse environmental impacts caused by animal pests. Improved road access could also increase the risk of bushfires, by making the area generally more available to the public.

It is unlikely that the mining operations would intrude on parts of the park that are routinely accessed by visitors, although some consideration of light spill and noise (blasting, offsite traffic) may be required as part of a future environmental impact assessment.

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Figure 20-14:      Locations of parks and reserves near Maud Creek

20.1.8

Heritage Values

Heritage surveys relating to both Aboriginal and non-Aboriginal values of the Maud Creek mine site were conducted as part of baseline studies for the Maud Creek EIS prepared by Terra Gold (URS, 2008, refer Figure 20-15).

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The studies identified sixteen Aboriginal archaeological sites, including 14 stone artefact scatters and three stone quarries (one the quarries was associated with an artefact scatter). Each site was rated as having high, moderate or low significance, based on rarity, intactness and age of the site. In all, over 240 stone artefacts were identified. Sixteen of the artefact scatters were considered to have high significance, three were classified as having moderate significance and five were described as having moderate to low significance. Two of the three stone quarries were described as highly significant.

A number of additional sites were identified on the Maud Creek tenements subsequent to the completion of Terra Gold’s EIS. The current Newmarket Gold Cultural Heritage Environmental Management Plan (2015) identifies 18 Aboriginal sacred sites. Disturbance of any Aboriginal sacred site would require formal authorisation (refer Section 20.2.8) .

Two non-Aboriginal heritage sites were identified during baseline surveys for Terra Gold’s Maud Creek project. One site comprised a series of historic alluvial gold mine diggings at the eastern side of the project area. The site was described as having significant heritage value. The second non-Aboriginal heritage site was a feature from an historic settlement to the northeast of the project area. The site consists of a raised stone hearth and a foundation made from cement, gravel and earth and edged with cobbles of rock. A variety of other artefacts were identified (shards of stoneware and earthenware, preserved meat tins, tobacco tins and matchboxes). These were generally rated as having a moderate to low significance.

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Figure 20-15:      Aboriginal and non-Aboriginal heritage sites (Figure 11.1 from URS 2008)

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20.1.9

Mine Closure and Revegetation

The EIS prepared for Terra Gold’s Maud Creek Project included a preliminary discussion of mine rehabilitation objectives and outcomes. The approach proposed by Terra Gold was to rehabilitate disturbed land to allow future use primarily for pastoral purposes, with some areas (such as waste rock dumps) potentially allocated for conservation purposes. It was suggested that water in pit voids might be used for watering livestock. The closure strategy also proposed to use pit voids for disposal of reactive materials remaining at the ROM. No closure provision was made for rehabilitation of tailings or waste rock storage facilities, as the project did not include these elements. A significant review of mine rehabilitation and closure will be required as part of project assessment and permitting, both to reflect changing community expectations and to align with evolving government regulations and policy. Long-term management of mine wastes and of water in pit voids is likely to be a central issue for mine closure design. Newmarket Gold should expect an increased focus on verifiable rehabilitation performance targets.

20.1.10

Potential Impacts

Although a range of environmental impacts is possible as a result of implementing the Maud Creek Project, the aspects most likely to attract the attention of stakeholders are:

Protection of surface water and groundwater quality, especially from contaminants contained in tailings and other mineral wastes;
     
  Avoidance of culturally significant areas;
     
Potential for impacts on public safety and/or amenity from mine traffic in or near the town of Katherine;
     
Control of indirect impacts (weeds, fire, feral animals) on the environmental values of Nitmiluk National Park; and
     
  Newmarket Gold’s ability to deliver acceptable mine rehabilitation outcomes .

A summary of project aspects and impacts is presented in Table 20-1.

A more detailed analysis and formal risk assessment of these aspects and impacts will be required as part of project assessment and permitting.

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Table 20-1:      Preliminary aspects and impacts analysis

  Social /
economic
 Public
health &
safety
 Aboriginal
heritage
 Surface
water
 Groundwater  Soil quality /
land capability
 Flora /
vegetation
 Fauna &
habitats
 Air quality
Noise  &
vibration
Key issues
Land access & clearing
for mine development
X X X X X X

Need to minimise footprint overall; implementation of weed / pathogen hygiene procedures; limiting impacts to riparian systems;  avoidance of sacred sites; operational controls to minimise dust generation, erosion and noise.

Land access & clearing
for access road(s)
X X X X X X

Need to confirm avoidance/ impact minimisation to short range endemic fauna (eg land snails) and high conservation value vegetation assemblages. Most likely not a significant issue if western access route is not adopted.

Mine dewatering X X X

Strong potential for public concern about possible contamination (discharge of water) and/or abstraction of water

Mine operations: haulage
& traffic within mine site
X X X

Unlikely to generate significant impacts / concern: routine operational practices to minimise footprint and control noise & dust.

Mine operations: offsite
haulage & traffic
X X X X

Has potential to generate significant concern relating to traffic  impacts (dust, noise, public safety, amenity, impacts on public infrastructure.

Blasting, excavation,
heavy equipment operation
X X

Unlikely to generate significant impacts: routine operational  practices to minimise footprint and control noise & dust.

Handling and stockpiling
of waste rock
X X X X

Key concerns likely to relate to i) management of reactive (acidic generating) or erodible materials and ii) ability to rehabilitate land

Ore processing X X X X

MMP will need to demonstrate appropriate systems for controlling noise, dust, process effluents.


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  Social /
economic
Public
health &
safety
Aboriginal
heritage
Surface
water
Groundwater Soil quality /
land capability
Flora /
vegetation
Fauna &
habitats
Air quality Noise &
vibration
  
Key issues
Tailings storage X X X

Key concerns likely to relate to i) management of reactive (acid generating, saline or metalliferous) or erodible materials; ii) ability to rehabilitate land; iii) risks of dam failure (especially with respect to surface water impacts) during or after mine operations.

Establishment and
operation of mine
accommodation
X

May generate concerns about housing availability / cost; other impacts relating to social integration and increased traffic. Also potential for social benefits.

Waste generation(non-
process wastes)
X X

Unlikely to be a major concern, providing conventional environmental management controls are in place for handling and disposal of sewage, domestic waste and industrial waste.


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20.2

Regulatory Approvals


20.2.1

Mineral Titles Act

With the exception of certain prescribed substances (such as uranium), all minerals located in the Northern Territory (except for certain prescribed substances including uranium) are the property of the Northern Territory Crown. The Mineral Titles Act took effect in November 2011, replacing the Mining Act. The Mineral Titles Act provides a framework for granting and transferring mineral titles to authorise exploration for and extraction of minerals. It also establishes a basis for authorising ancillary activities related to mining.

The Mineral Titles Act recognises seven categories of mineral title, of which the most relevant to the Maud Creek Project is a mineral lease (ML). Newmarket Gold was granted a mineral lease (ML30260) over the project area on 14 April 2014. The lease over the 106 ha tenement is valid until 13 April 2024 and can be extended, subject to certain conditions (including expenditure requirements and compliance with tenement conditions).

Grant of a mineral lease gives the tenement holder exclusive rights for the development of mineral projects on land within the title boundaries. Subject to the tenement holder obtaining an authorisation under the Mining Management Act, the grant of a mineral ease confers rights to conduct mining and processing of minerals, as well as a range of mining related activities, including:

  Exploration for minerals;
     
  Treatment of tailings; and
     
  Storage of waste.

Additionally, the Mineral Titles Act provides the holders of mineral titles strong rights of access and important entitlements related to the taking and use of water. For example, a mineral title holder may apply for an access authority to enter land which is outside the title area for the purpose of constructing infrastructure or for the taking of water.

The Mineral Titles Act also provides a basis for assigning liability for mine rehabilitation. An entity taking ownership of a tenement on which disturbance exists (as is the case at Maud Creek) assumes the risks and responsibility for that tenement, irrespective of whether it causes any additional mining disturbance to the land.

20.2.2

Mining Management Act

Before any mining activity can occur on a granted mining title the intending operator of the mining activity must apply for and be granted an authorisation under the Mining Management Act. The Mining Management Act (MMA) effectively constrains rights conferred by the Minerals Title Act (MTA) by prohibiting mining activities until the intending operator (which may or may not be the tenement holder) has:

Been granted an authorisation (supported by an approved mining management plan (MMP)), and
     
Lodged a security to cover the full costs of rectifying any environmental harm arising from mining activities and for final rehabilitation of the affected area (including rehabilitation of any legacy disturbance).

Authorisations granted under the Mining Management Act typically cover all mining (and related) activities required for project implementation. In some circumstances, separate authorisations may be granted for activities (operation of an explosives storage facility, accommodation village, power station) which the mine operator does not itself wish to operate and manage.

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Mining management plans (MMP) are binding legal instruments through which pollution and environmental harm are controlled on mine tenements. Operators of mining projects are exempt from some licensing requirements that would normally apply to comparable activities on non-mining land. The MMP instead serves regulate resource usage (taking of water) and other environmental aspects of mining (and related activities) within the bounds of the granted tenement. It is an offence to contravene the environmental obligations arising from an approved MMP, if the activities result in environmental nuisance, harm or pollution. More generally, the Act prohibits the release of wastes or contaminants unless the release is done in accordance with an approved MMP.

20.2.3

Waste Management and Pollution Control Act 2009

The Waste Management and Pollution Control Act is administered by the NT EPA. The Act regulates the collection, transport, storage, treatment and disposal of listed wastes. Wastes – including both process and non-process wastes – arising from mining activities are generally exempt from licensing requirements under this Act and are instead regulated through the approved mining management plan.

20.2.4

Water Act 1992

Mining operations in the Northern Territory are not required to seek licences for abstraction and use of water under the Northern Territory Water Act, providing their use, storage and management of water is done in accordance with an approved mining management plan. Mine tenement holders are entitled to construct bores, and to take or divert water (not including water in dam or wells belonging to others) and to use the water for mining and mineral processing.

Pollution provisions of the Water Act generally do not apply to mining operations, except in the event that contamination or pollution moves beyond the tenement boundary (either through an intentional or accidental discharge). Mine operators are exempt from discharge licensing requirements that would normally apply under the Water Act, to the extent that:

  The contaminant or waste results from carrying out of a mining activity;
     
The waste or contaminant discharge is confined within the land on which the mining activity is carried out; and
     
  The discharge is done in accordance with an approved mining management plan.

In circumstances where a discharge of a waste or contaminant (for example, water from mine dewatering) is likely to move beyond the boundaries of the mine tenement, a discharge licence would be required. As an example, Newmarket Gold holds a waste discharge licence (WDL 138-02) to authorise release of wastewater from its Union Reefs site to Wellington Creek and the McKinlay River. Newmarket Gold has implemented comprehensive environmental management systems for several of its NT operations: similar arrangements would be appropriate at Maud Creek.

A waste discharge licence may be required for activities that could result in seepage that could cause pollution of an aquifer. The implications of this for storage of tailings and waste rock and for the development of permanent pit voids would need to be discussed with NT regulators.

20.2.5

Aboriginal Land Rights (Northern Territory) Act 1976 (Cwlth)

Contemporary use of the term ‘traditional owner’ largely derives from the Aboriginal Land Rights (Northern Territory) Act 1976 (ALRA). The ALRA established ways for Aboriginal people to claim land in the territory on the basis that they were the traditional Aboriginal owners of the land. Further, the ALRA gave traditional Aboriginal land owners the right to control access to their land, including the right to withhold consent for exploration activities on their land. This Act has limited, if any, relevance to the Maud Creek Project, as mining tenure has already been granted and the project lies on freehold land that is not subject to the control of any Aboriginal Land Trust.

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20.2.6

Native Title Act 1993 (Cwlth)

The Native Title Act 1993 establishes a legal framework for the recognition and protection of native title. It also establishes an administrative system for:

  Determining claims to native title;
     
Validating past acts which might otherwise be invalidated because of the existence of native title; and
     
  Developing processes and standards for future dealings which could affect native title.

The Maud Creek Project is located on freehold land, and therefore is not exposed to native title claims. There are no registered or determined native title claims over the project area.

20.2.7

Heritage Act 2011

The Heritage Act 2011 declares all Aboriginal archaeological places and objects heritage places. Other historically or culturally significant places (for example, buildings and heritage objects) are also protected under the Heritage Act if they are listed on the NT Heritage Register. No non-Aboriginal heritage sites formally registered on the Australian Heritage database are known to occur in the project area. No formal authorisation would be required in relation to disturbance of the non-Aboriginal heritage sites and Terra Gold made no commitments to the protection or conservation of these sites.

The Act makes provision for approval of works to salvage Aboriginal archaeological sites, subject to the consent of the appropriate Traditional Owner or Site Custodian. It is normally a requirement that appropriate studies be conducted to assess the character, extent and value of a site before salvage. A person proposing to disturb a heritage site would be expected to commit to a programme to prevent or minimise damage and to make provision for suitable curation of any artefacts. For Aboriginal sites, suitable curation typically involves return of the object(s) to the Traditional Owners or protection if artefacts are outside of the operational footprint.

20.2.8

Northern Territory Aboriginal Sacred Sites Act 1989

The Aboriginal Sacred Sites Act provides the legal basis for protection of all Aboriginal sacred sites, irrespective of whether the site is formally registered. Carrying out work – or even accessing – a sacred site requires a consent in the form of a certificate issued by the administering agency, the Aboriginal Areas Protection Authority (AAPA).

A certificate (reference number D89/199:90/804 (Doc: 68552) C2009/266 (supersedes C2007/072)) was issued by the AAPA to Crocodile Gold on 8 October 2009. The certificate authorises access to / disturbance of nominated Aboriginal sites within the Maud Creek mine site (ML30260 – formerly MLN1978), as well as some other sites on other Crocodile Gold tenements named on the certificate. The consent is subject to a range of implementation conditions, including a condition excluding works or access to designated sites. No works of any kind are permitted at sites SS5369-69 (restricted work area 1, 5369-32 (restricted work area 2) or 5369-27. Two of the exclusion areas (restricted work areas 1 and 3) lie ML30260, refer Figure 20-16.

A further implementation condition (Condition 3) says, This certificate shall lapse and be null and void, if the works in question or the proposed use is not commenced within 24 months of this certificate. Newmarket Gold has advised that some of the works approved under the AAPA certificate (specifically, exploration drilling) was carried out in 2011, prior to the expiry of the AAPA consent. Accordingly, the consent is still valid.

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With regards to the NT Aboriginal Sacred Sites Act, Newmarket Gold have the requisite permit (AAPA Authority Certificate C2009/266 superseding C2007/072), provided the project footprint remains within the Subject land as defined in the Authority Certificate, and avoids the two exclusion areas, then no further permitting is required.

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Figure 20-16:      Restricted work areas (excerpt from AAPA certificate, June 2007)

20.3

Waste Rock and Ore Geochemistry


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20.3.1

Available Data

Drillhole Database

The current drillhole assay database (Database_08_05_2015) contains assay data for Au, As, Cu, Pb, Zn, Cr, and Ni. A smaller extended assay dataset was obtained from drilling in 2011, which includes Ca, Mg, Ag, Bi, Fe, Hg, K, Mn, Mo, Na, S, Sb, Sn, Ti and Zr.

Total sulphur can sometimes be used as an indicator of acid generating potential - the assumption being that all sulphur is present in the form of potentially reactive, acid generating sulphides.

The sulphur assay sampling density was low, with 2,140 samples collected from 21 drillholes (Figure 20-17) relative to the 2,129 drillholes included in the database.

The sulphur assay data was separated into waste-grade and ore-grade assuming a gold cut-off value of 0.1 mg/kg. Sulphur content in the waste rock (n=1,831) assays ranged from <0.05 -8.55% S (mean 0.16% S), while in the ore-grade assays sulphur ranged between <0.05 -10.7% S (mean 0.92% S). In the waste rock lithologies the highest sulphur contents were present within the dolerite (8.55% S), sediments (SLST, 3.75% S), Cambrian Basalt (3.55% S), tuff (2.75% S) and veins (2% S).

Figure 20-17:      Drillhole database S assay data

Notes: The figure also shows the proposed open-cut pit (grey), underground workings (blue) and modelled deposit (green).

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Detailed Geochemical Characterisation

Previous geochemical characterisation of waste rock (33 samples), ore materials (11 samples) from the weathered (or oxide) zone and the fresh (unweathered) zone, and soil (2 samples) from the Maud Creek deposit was undertaken in 1996 (Graeme Campbell & Associates, 1997). The assessment was prepared in support of an Environmental Impact Statement prepared by Kalmet Resources NL (Kalmet), to identify materials that may generate Acid Mine Drainage (AMD) and/ or impact water quality and revegetation.

The sampled materials included 33 waste rock samples, 11 ore samples and 2 soil samples. The waste rock and ore samples were collected from the weathered and fresh rock zones – a breakdown of the lithologies sampled is given in Table 20-2.

Table 20-2:     Sampled materials (GCA, 1997)

Waste
Rock/Ore
Weathering
Zone

Kalmet Lithology

SRK Lithology No. of
Samples
Waste
Rock
Weathered
Zone

Hangingwall Mafic Tuff

Tuff 10

Footwall Sediments Undifferentiated

Sediments (SDST) 1

Footwall Sandstone

Sediments (SDST) 3
Fresh Zone

Hangingwall Mafic Tuff

Tuff 11

Footwall Sediments Undifferentiated

Sediments (SDST) 4

Footwall Sandstone

Sediments (SDST) 4
Ore Weathered
Zone

Stockwork Quartz - Tuff Hosted

Vein 1

Stockwork Quartz - Sediment Hosted

Sediment (MSED) 1

Massive Quartz - Tuff Hosted

Vein 1
Fresh Zone

Stockwork Quartz - Graphitic
Sediment Hosted

Vein (2), MSED (1) 3

Stockwork Quartz - Tuff Hosted

Tuff 1

Massive Quartz - Graphitic Sediment Hosted

Vein (1), Tuff (1) 2

Massive Quartz - Tuff Hosted

Tuff (1), Vein (1) 2

The assessment included the following test work on all 44 samples:

  pH and EC (paste/slurries; 1:2 solid : water.)
     
  Total S.
     
  Acid Neutralisation Capacity (ANC)1.
     
  Net Acid Producing Potential (NAPP).

The following additional test work was carried out smaller sub-sets of samples:

 

Sulphate-sulphur (SO4-S) and sulphide-sulphur (14 samples).

     
 

Net Acid Generation (NAG) (15 samples).

     

Multi-element analyses, (12 samples) (digest not specified) including the following elements: Ag, Al, As, B, Ba, Bi, C, Ca, Cd, Co, Cr, Cu, F, Fe, Hg, K, Mg, Mn, Mo, Na, Ni, P, Pb, Sb, Se, Sn, Sr, Th, Tl, U, V, Zn).

________________________________
1
Modified ANC test (unspecified number of tests on waste rock) using dilute sulphuric acid to measure readily available ANC (ANCC).

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Saturation-Extract (SE) tests (7 waste rock samples) to assess the soluble salts content (Na, K, Mg, Ca, Cl and SO4) and the solubility of As and Sb.

20.3.2

Sample Representivity

The drill-hole assay samples with sulphur data are generally located in rock volumes that lie outside of proposed open pit and underground workings (Figure 20-17).

With respect to samples submitted for detailed characterisation, six of the uppermost samples occur within depths ranging from 2-3 m below ground level (mbgl) to 174-175 mbgl; and include 7 samples from the weathered (or oxide) zone and 19 from the fresh zone2. No samples have been collected below 175 mbgl.

In conclusion, current datasets do not adequately represent the volumes to be mined as part of the currently proposed open pit and underground workings.

20.3.3

Geochemical Characteristics

Results generated from the detailed characterisation are summarised in Table 20-3. Note that, as discussed in the previous section, the samples studied do not give good representation of material that could be mined as part of currently proposed open pit and underground workings. However, the data obtained have value in that they can be used to infer some characteristics of materials located within the topmost 175 m.

Most of the waste rock samples contained an excess of ANC and would be classed as NAF (Figure 20-17). Highest ANC was offered by the Hangingwall Mafic Tuff samples; however, note that only a portion (10-20%) of ANC was considered readily available for reaction. The balance of acid potential and neutralising capacity is closest in the case of the Footwall Sediments and Footwall Sandstone samples, including three samples that were classed as having uncertain acid generating potential and two samples that classed as PAF.

Multi-element analyses results were used to calculate Global Abundance Indices (GAI) to identify elements that are enriched relative to average crustal abundance concentrations. In both waste rock and ore, elements identified as being enriched included arsenic, antimony, chromium, nickel, tin, boron and silver.

Saturation-Extract (SE) tests (9 samples waste rock; 2 samples ore) gave alkaline extraction solutions (pH 8.6 -9.2) . Of the enriched elements listed above, only arsenic and antimony were included in the leach analysis suite – giving dissolved concentrations of arsenic up to 0.56 mg/L (one of the ore samples) and antimony up to 0.046 mg/L (one of the waste rock samples). The results suggest that both arsenic and antimony could be leachable from material to be mined at Maud Creek. [It is noted that the existing pit lake water was found to contain elevated arsenic concentrations (185-201 μg/L; relative to the ANZECC (2000) freshwater trigger value of 13 μg/L).]

______________________________
2
More recent geological interpretation of the weathering zone extents would place 15 of the waste rock samples within the oxide (weathered) zone, 1 within the transitional zone and 17 in the fresh zone.

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Figure 20-18:      Geochemical classification of waste rock samples using NPR

Note: The plot includes lines to show where the NPR (ANC/MPA) values equal 1 and 3, indicating boundaries between PAF, UC and NAF regions on the plot. A line to show where the ANC/MPA value equals 2 is also shown, as values greater than 2 are considered to be unlikely to be problematic with respect to AMD (DITR, 2007).

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Table 20-3:      Summary of Available Results, Detailed Geochemical Characterisation Waste Rock and Ore

Parameter Range (Waste Rock) Range (Ore) Comment
Paste pH 7.6-9.4 6-9.1  
Paste EC, μS/cm 38-1800 34-490  
Total sulphur, % <0.01-2.1%
<0.01-0.11 % (weathered zone)
<0.01-2.1 % (fresh zone)
<0.01-3.2
<0.01-0.51 % (weathered zone)
0.38-3.2 % (fresh zone)
Where sulphur speciation test work was available (fresh zone waste rock, and ore), sulphide was the dominant form of sulphur, accounting for between 94-99% of the total sulphur content.
Acid neutralising capacity,
kgH2SO4/t
280-540 (Hangingwall mafic tuff)
1.9-6.7 (other lithologies in the weathered zone)
25-53 (other lithologies in the fresh zone)
<0.5-380
<0.5-160 (sediment-hosted ore)
The readily available ANC3 of the Hangingwall Mafic Tuff from the weathered and fresh zones were assessed to range up to 50 kgH2SO4/t and 20 kgH2SO4/t, respectively.
Net acid producing potential
(NAPP), kgH2SO4/t
-539 to -23 (Hangingwall mafic tuff)
Close to zero (Footwall sediment and sandstone)
-350 to -1 (most samples)
8.4 (Graphitic Sediment Hosted
Massive Quartz sample)
Net acid generation (NAG) testing
NAG pH
NAG acidity, kgH2SO4/t
2.5-9.3
<0.5 - 36
7.9-10.4
<0.5

5 of 6 waste rock samples and 9 ore samples generated no acidity.

One waste rock sample (Footwall Sediment) was acid generating (NAG pH 2.5; NAG acidity, 36 kgH2SO4/t)

Geochemical Classification
NPR
AMIRA
NAF (26), UC (5), PAF (2)
NAF (4), UC(NAF) (1), PAF (1 – footwall sediments)
NAF (7), UC (4), PAF (1)
NAF (9), UC(NAF) 1

_____________________________________________
3
Readily available ANC was assessed using a modified ANC test – details not of test work method not specified.

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20.4

Mineralised Waste Management

It is understood that there is an existing stockpile comprising approximately 300,000 m3 oxidised, non-mineralised, waste rock. Any future project is likely to generate additional stockpiles that would need to be managed.

Geochemical characterisation datasets are limited. There is insufficient information to determine the potential for acid, metalliferous drainage from the wastes with any degree of confidence. While oxidised material is currently considered to be non-acid forming (NAF); further data are required to verify that this would be true of oxidised volumes to be mined as part of current plans. Even if classed as NAF, the potential for leaching of elements such as arsenic, antimony and selenium under pH neutral, oxidising conditions should be evaluated.

Some important issues with respect to waste rock management requirements are:

Determination of the volumes of NAF and PAF-classed materials that would require management on the surface.

   

 

For high sulphide wastes (whether PAF or NAF-classed), waste storage facilities should be designed so that potential base seepage and surface run-off are managed. This would mitigate the potential for water quality impacts from acidic, metalliferous drainage (PAF wastes) or neutral saline drainage (NAF-classed sulfidic wastes).

   

 

Should material be identified that reacts at a very high rate (e.g. high sulphide content material), then it may be advisable to use dump designs and construction methodologies that control and minimise oxidation rates. The objective would be to minimise accumulation of reaction products during storage on the surface. Such products could dissolve in contact waters when stored material is returned underground and is inundated as groundwater rebounds . Minimisation of accumulation rates on the surface will minimise potential impacts on groundwater quality later.


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21

Capital and Operating Costs

This section is not applicable at this time.

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22

Economic Analysis

This section is not applicable at this time.

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23

Adjacent Properties

Other projects in the region managed and owned by Newmarket Gold include: –

The Pine Creek Gold Project (Figure 23-1), located approximately 190 km south southeast of Darwin accessible by the Stuart Highway, directly to the west of the Township of Pine Creek.
     
The Union Reefs Gold Project (Figure 23- 2), located approximately 180 km south-southeast of Darwin accessible by the Stuart Highway, directly to the north of the Township of Pine Creek.
     
The Burnside Gold & Base Metals Project (Figure 23-3), located approximately 130 km to the South-southeast of Darwin accessible by the Stuart Highway.
     
The Moline Gold and Base Metals Project (Figure 23-4), located approximately 40 km to the Northeast of the Township of Pine Creek, accessible by the Kakadu Highway.
     
The Yeuralba Gold and Base Metals Project (Figure 23-5), located approximately 45 km to the northeast.

Figure 23-1:      Pine Creek Gold Project Tenements

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Figure 23-2:      Union Reefs Gold Project Tenements


Figure 23-3:      Burnside Gold Project Tenements

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Figure 23-4:      Moline Gold Project Tenements

Figure 23-5:      Maud Creek & Yeuralba Gold Project Tenements

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24

Other Relevant Data and Information

The filing of this Technical Report to support the release of the Mineral Resource estimate includes a review of the geotechnical conditions and metallurgical aspects of the Mineral Resource that will support ongoing work on the Project that includes a PEA.

The PEA is considering two processing routes that will underpin the decision on mining method and have an impact on the particular infrastructure requirements.

The two processing routes includes a stand-alone processing plant at Maud Creek and trucking ore to the Union Reefs processing plant. Under both options a gold-rich concentrate would be produced

No other relevant/material data has not been included in the Technical Report.

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25

Interpretation and Conclusions

The modelling of vein and grade volumes for the 2015 estimate takes a very different approach to the 2012 model as described in the sections above. The 2015 model incorporates a detailed structural, vein and lithological model in the construction of the various estimation domains. This was a deliberate decision as it was SRK’s understanding from discussions with Newmarket Gold and from reading previous reports that there were potential deficiencies in the very linear grade only approach previously used. Concerns had been expressed in some previous reports that insufficient attention had been paid to the geology and that the previous models may have diluted a high grade, geologically controlled core to the main zone thereby creating a model that underestimated grade and overestimated tonnages at economic cut-offs.

There are a number of differences between the 2012 and 2015 modelling approach. The 2015 model uses the following:

Pure geology to define the main and minor vein domains. The 2012 used grade only. Consequently the 2015 vein model contains considerably lower tonnage and slightly elevated grade in comparison.

     

Grade halos to capture both high and low grade outside the geological veins. This captures low grade material that was not modelled in 2012 which may be of value in an open pit scenario.

     

Orientation controls on the grade halos derived by the combined fault/lithology contact model resulting in multiple orientations and fattening around fault and contact intersections.

The 2015 model is more robust in terms of its geological basis and this has led to a slightly higher grades but a reduction in contained gold. Only further drilling can define true connectivity of the mineralization in widely spaced areas.

A discussion on the risks and opportunities in the Mineral Resource model as discussed in Table 25-1.

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Table 25-1:      Mineral Resource Model Risks and Opportunities

Project Element Economic Risk Level Comment Opportunity
Database – Exploration data Low

Historical and recent data have been re-collated and re-validated for this Mineral Resource estimate.

Assaying Low

QAQC for recent and older assaying shows no material issues. Arsenic assaying has incomplete coverage.

Additional assaying for Arsenic may be beneficial depending on the processing method.

Surveying Low

Both collar surveys and downhole surveys completed to a high level of accuracy for recent drilling. Representative collars resurveyed for older drilling with no significant discrepancies.

Geology Low

Detailed logging and interpretation together with evidence from both regional structural features and detailed in pit mapping informs the geological understanding

Additional drilling may be able to add detail to the interaction of structures controlling mineralization at depth.

Geological modelling Low

A detailed structural and lithological model has been built and incorporated into the estimation domain construction.

Additional drilling may be able to add detail to the interaction of structures controlling mineralization at depth.

Resource Estimation Low

Ordinary kriging cross checked and validated with theoretical grade tonnage curves and alternative search parameters has been used.

The project may benefit from simulation studies or non-linear estimates if detailed studies at selective mining unit block sizes are required in the future.

The absence of suitable data has led to low-confidence in the geotechnical conditions. Additional data will be required to improve confidence and refine decisions on mining methods and the mine design. The mining method studies are linked to the decision on the location of the processing plant and the availability of pastefill.

The Maud Creek mineralization has been subject to extensive metallurgical testing. However, there remain some gaps that should be covered for future studies that could reduce the future design risks if it was decided to undertake additional test work earlier. This would include, lithological domain characteristics, SAG milling parameters, testing samples from depth, flotation test work and concentrate production and specification.

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26

Recommendations

Drilling Campaign

Infill drilling in the parts of the mineral resource currently classified as Indicated and Inferred would enable an upgrade of the Mineral Resource Classification. An approximate meterage and cost to complete this from surface down to 850 mRL is provided in Table 26-1 assuming the drilling takes place from surface. A number of sections of the geological model remain open down dip with good grades seen in the last hole down dip. Extension drilling is recommended to test these areas. Metres and costs to complete these are shown in Table 26-1 assuming drilling from surface. Costs are based on RC collars and 50m diamond drill tails.

Table 26-1:      Recommended drilling

Target Current
exploration
status
Potential End
of 2016 Status
Description Drilling
(Meters)
Total Cost
(AUD)
Infill Drilling Indicated and/or Inferred Measured
and/or Indicated
Increase
confidence in estimated
Mineral Resource
9 200 770 000
Extension
drilling
Down dip or
along strike
from current
Mineral Resource
Indicated
and/or Inferred
Close off or extend
Mineral Resource volumes
2 200 200 000

Geotechnical

If an additional phase of geotechnical work is undertaken, SRK recommends that a drilling program be considered to provide additional geotechnical data and infrastructure for hydrogeological test work. These programmes should be combined where possible to assist in any further resource definitions requirements.

Based on further drilling and geotechnical data, further studies should be undertaken to improve the confidence of the understanding of the geotechnical domains and to provide a basis for the improving the design guidelines.

Geometallurgy

Insufficient information was available from the metallurgical reports to create preliminary spatial domains for the physical processing parameters, such as grindability. The metallurgical report Core Process Engineering Report No. 140-001 outlined JK Drop Tests which had been conducted, resulting in a Bond ball Work Index of 18-19 kWh/tonne for the main ore lode. Previous geology reports, as well as a site visit to inspect the core, gave an indication of the mineralogy of the Maud Creek deposit, which has an impact on the hardness of the ore. From a geometallurgical perspective, between and within each ore domain there will likely be a range of hardness values which will need to be established. Quartz alteration at Maud Creek has been identified as varying between the primary lode (high percentage quartz veining), moderate (footwall and hangingwall lodes stock work veining) to low (sandstone and tuff country rocks). However, variations in silica content within these lodes will definitely occur, as alteration boundaries are normally pervasive across lithology boundaries. A possible way to better define the hardness parameters is the use of proxies. If silica analysis is included in any future assay testing of the mineralized zones, this can be used to identify target areas for JK drop weight tests. A regression calculation between the A*b result, Bond work indices and other comminution parameters versus the silica content can then be determined, allowing a predictive model of the ore hardness to be created.

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SRK Consulting Page 197

Mineral analysis would need to be conducted to ensure the silica content is reflecting quartz alteration which has a high hardness (Moh’s scale 7) and not feldspar (Moh's scale 5-6) or mica minerals (Moh’s scale 2-3) which have lower hardness and varying crystallography, resulting in different grinding behaviours. The Geology of the Maud Creek Gold Deposit and Maud Creek Reconciliation report by AngloGold in 2000 identified plagioclase laths and alkali feldspar minerals in the Andesite unit located to the south of the deposit, but not within the ore lodes. However, sericite (a type of muscovite) and chlorite alteration (most likely the mineral clinocore) are both quite prevalent in all three mineralized zones, so their prevalence would need to be established as well. Similarly, the extent of the hematite and marcasite alteration noted in all three mineralized zones (present as pseudomorphs of pyrite in the oxide zone) needs to be quantified, as all three will have moderate-high hardness (6-7), but also a brittle tenacity upon breakage, which will influence the ore grindability.

The Maud Creek mineralization is going to be very hard and abrasive so finding a correlation with proxies such as silica, for example, is recommended by SRK. Equally, determining correlations between arsenic, sulphur and gold to help generate the flotation relationship correlations, and approximated the amount of arsenic in the final concentrate is also recommended. The arsenic and sulphur contents will also be a good indicator of recovery.

FAIR\KENT\mota CGC001_Maud Creek Resource Report_NI_43 101_Rev1.Docx 21 March 2016



SRK Consulting Page 198

27

References

Bremner, P and Edwards, M 2012. Report on the Mineral Resource and Mineral Reserve of the Maud Creek Gold Project.

Pakalnis, R C, Poulin, R and Hadjigeorgiou, J, 1995. Quantifying the cost of dilution in underground mines, Mining Engineering, 47(12): 1136-1141.

Scoble, M.J., Moss, A., 1994. Dilution in underground bulk mining: Implications for production management, mineral resource evaluation II, methods and case histories, Geological Society Publication No. 79, pp. 95-108.

FAIR\KENT\mota CGC001_Maud Creek Resource Report_NI_43 101_Rev1.Docx 21 March 2016



SRK Consulting Appendices

Appendices

FAIR\KENT\mota CGC001_Maud Creek Resource Report_NI_43 101_Rev1.Docx 21 March 2016



SRK Consulting Appendix A

Appendix A: Geotechnical Report

FAIR\KENT\mota CGC001_Maud Creek Resource Report_NI_43 101_Rev1.Docx 21 March 2016



Kirkland Lake Gold Ltd.: Exhibit 99.24 - Filed by newsfilecorp.com

Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

REPORT ON THE

MINERAL RESOURCES & MINERAL RESERVES

OF THE

STAWELL GOLD MINE

In the

State of Victoria

Australia

FOR

NEWMARKET GOLD INC.

Effective Date December 31st 2015

Dated March 16th 2016

Justine Tracey, BScH MAusIMM (CP)
Senior Resource Geologist
 
 

Wayne Chapman, BEng MAusIMM (CP)   Mark Edwards, BSc MAusIMM (CP)  
Technical Manager   General Manager Exploration  
     
   



Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

REPORT ON THE MINERAL RESOURCES & MINERAL RESERVES OF THE STAWELL GOLD MINE,

In the State of Victoria Australia

FOR NEWMARKET GOLD INC.

CAUTIONARY NOTE WITH RESPECT TO FORWARD LOOKING INFORMATION

This document contains “forward-looking information” as defined in applicable securities laws. Forward looking information includes, but is not limited to, statements with respect to the future production, costs and expenses of the project; the other economic parameters of the project, as set out in this technical report, including expansion plans; the success and continuation of exploration activities, including drilling; estimates of mineral reserves and mineral resources; the future price of gold; government regulations and permitting timelines; requirements for additional capital; environmental risks; and general business and economic conditions. Often, but not always, forward-looking information can be identified by the use of words such as “plans”, “expects”, “is expected”, “budget”, “scheduled”, “estimates”, “continues”, “forecasts”, “projects”, “predicts”, “intends”, “anticipates” or “believes”, or variations of, or the negatives of, such words and phrases, or statements that certain actions, events or results “may”, “could”, “would”, “should”, “might” or “will” be taken, occur or be achieved. Forward-looking information involves known and unknown risks, uncertainties and other factors which may cause the actual results, performance or achievements to be materially different from any of the future results, performance or achievements expressed or implied by the forward-looking information. These risks, uncertainties and other factors include, but are not limited to, the assumptions underlying the production estimates not being realized, decrease of future gold prices, cost of labor, supplies, fuel and equipment rising, the availability of financing on attractive terms, actual results of current exploration, changes in project parameters, exchange rate fluctuations, delays and costs inherent to consulting and accommodating rights of local communities, title risks, regulatory risks and uncertainties with respect to obtaining necessary permits or delays in obtaining same, and other risks involved in the gold production, development and exploration industry, as well as those risk factors discussed in Newmarket Gold Inc.’s latest Annual Information Form and its other SEDAR filings from time to time. Forward-looking information is based on a number of assumptions which may prove to be incorrect, including, but not limited to, the availability of financing for Newmarket Gold Inc.’s production, development and exploration activities; the timelines for Newmarket Gold Inc.’s exploration and development activities on the property; the availability of certain consumables and services; assumptions made in mineral resource and mineral reserve estimates, including geological interpretation grade, recovery rates, price assumption, and operational costs; and general business and economic conditions. All forward-looking information herein is qualified by this cautionary statement. Accordingly, readers should not place undue reliance on forward-looking information. Newmarket Gold Inc. and the authors of this technical report undertake no obligation to update publicly or otherwise revise any forward-looking information whether as a result of new information or future events or otherwise, except as may be required by applicable law.

NON-IFRS MEASURES

This technical report contains certain non-International Financial Reporting Standards measures. Such measures have non standardized meaning under International Financial Reporting Standards and may not be comparable to similar measures used by other issuers.

i



Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

TABLE OF CONTENTS PAGE
   
1 EXECUTIVE SUMMARY
1
     
  1.1 PROJECT DESCRIPTION AND OWNERSHIP 2
  1.2 GEOLOGY & MINERALIZATION 3
  1.3 EXPLORATION, DEVELOPMENT AND OPERATIONS 4
  1.4 MINERAL RESOURCES AND MINERAL RESERVES 4
    1.4.1 STAWELL GOLD MINES 4
    1.4.2 UNDERGROUND 6
    1.4.3 BIG HILL SURFACE 7
    1.4.4 LOW GRADE STOCKPILES 8
  1.5 INTERPRETATION 10
  1.6 CONCLUSION AND RECOMMENDATIONS 10
       
2 INTRODUCTION AND TERMS OF REFERENCE 12
     
  2.1 INTRODUCTION 12
  2.2 TERMS OF REFERENCE 12
  2.3 AUTHORS’ QUALIFICATIONS & RESPONSIBILITIES 13
  2.4 DEFINITIONS 14
  2.5 MINERAL RESOURCE AND RESERVE DEFINITIONS 16
    2.5.1 MINERAL RESOURCES 16
    2.5.1.1 INFERRED MINERAL RESOURCE 17
    2.5.1.2 INDICATED MINERAL RESOURCE 17
    2.5.1.3 MEASURED MINERAL RESOURCE 18
    2.5.1.4 MINERAL RESOURCE AND MINERAL RESERVE CLASSIFICATION 18
    2.5.2 PRELIMINARY FEASIBILITY STUDY 20
    2.5.3 FEASIBILITY STUDY 20
    2.5.4 MINERAL RESERVES 20
    2.5.4.1 PROBABLE MINERAL RESERVE 21
    2.5.4.2 PROVEN MINERAL RESERVE 21
         
3 RELIANCE ON TECHNICAL EXPERTS – MINERAL RESOURCE ESTIMATES 23
     
  3.1 HISTORICAL INFORMATION 23
       
4 PROPERTY DESCRIPTION & LOCATION 25
  4.1 LOCATION 25
  4.2 PROPERTY DESCRIPTION 25
  4.3 LEGISLATION AND PERMIT 27
  4.4 ROYALTY AND ENCUMBERANCES 29
  4.5 ENVIRONMENTAL LIABILITIES 29
  4.6 STAWELL GOLD MINES LOCAL SURVEY GRID REFERENCE 30
     
5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE & PHYSIOGRAPHY 32
   
  5.1 ACCESSIBILITY 32
  5.2 CLIMATE 32
  5.3 LOCAL RESOURCE 32
  5.4 INFRASTRUCTURE 32
  5.5 PHYSIOGRAPHY 33
  5.6 MINING PERSONNEL 33
  5.7 PROCESSING FACILITIES 33

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Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

6 HISTORY 35
       
  6.1 HISTORICAL AND MODERN PRODUCTION 36
       
7 GEOLOGICAL SETTING AND MINERALIZATION 38
       
  7.1 REGIONAL GEOLOGY 38
  7.2 LOCAL GEOLOGY AND PROPERTY 39
    7.2.1 STRATIGRAPHY AT STAWELL GOLD MINES 39
    7.2.1.1 MAGDALA BASALT 40
    7.2.1.2 LEVIATHAN FORMATION 40
    7.2.1.3 MAGDALA FACIES 41
    7.2.1.4 FELSIC INTRUSIONS 41
    7.2.1.5 STAWELL GRANITE 41
    7.2.1.6 LAMPROPHYRIC INTRUSIONS 42
    7.2.2 STRUCTURAL HISTORY AT STAWELL 42
    7.2.2.1 EARLY, DUCTILE DEFORMATION 42
    7.2.2.2 LATE, BRITTLE DEFORMATION 42
    7.2.3 STAWELL MINE GEOLOGICAL ARCHITECTURE 45
  7.3 MINERALIZATION 50
   
8 DEPOSIT TYPES 53
   
  8.1 DEPOSIT TYPES AND MINERALIZATION 53
  8.2 ORE TYPES BY AREA 54
    8.2.1 FEDERAL ALBION AND FEDERAL ALBION SOUTH 54
    8.2.2 BELOW SCOTCHMAN 250 55
    8.2.3 UPPER SOUTH FAULT 2 57
    8.2.4 MAGDALA S6000 59
    8.2.5 MARINERS 59
    8.2.6 BIG HILL AND UPPER LEVELS 60
    8.2.1 AURORA B 63
   
9 EXPLORATION 65
     
  9.1 CURRENT EXPLORATION 69
    9.1.1 EXPENDITURE 72
  9.2 MAGDALA WEST FLANK 72
    9.2.1 FEDERAL ALBION SOUTH (FAS) 72
    9.2.2 UPPER SOUTH FAULT 2 (USF2) 72
    9.2.3 250L BELOW SCOTCHMANS 73
    9.2.4 486L MID NORTH MAGDALA 73
    9.2.5 GOLDEN GIFT OFFSET 73
    9.2.6 UPPER SOUTH FAULT 5 (USF5) 73
  9.3 MAGDALA EAST FLANK 74
    9.3.1 AURORA B 74
    9.3.2 AURORA A 74
  9.4 WONGA PIT (BRUMMIGANS) 75
  9.5 REGIONAL EXPLORATION 75
  9.6 ONGOING EXPLORATION PROGRAM 78
    9.6.1 FEDERAL ALBION SOUTH 82
    9.6.2 UPPER SOUTH FAULT 2 (USF2) 82
    9.6.3 486L MID-NORTH MAGDALA 82
    9.6.4 475L FEDERAL ALBION 82
    9.6.5 MARINERS 109 DOWN PLUNGE EXTENSION 82
    9.6.6 AURORA B 83
    9.6.7 MORAY 83

iii

Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

  9.6.8 BRUMMIGANS STRUCTURE 83
  9.6.9 GERMANIA 83
  9.6.10 WAL WAL 83
  9.7 EXPLORATION MINERAL TENURE 84
     
10 DRILLING 87
   
  10.1 STAWELL GOLD MINES MINERAL RESOURCE DEFINITION PROCESS 87
  10.2 DRILLING PROCESS 91
  10.3 DRILL SPACING 93
  10.4 DRILLHOLE ORIENTATION 94
  10.4.1 UNDERGROUND 94
  10.4.2 SURFACE 96
  10.5 COLLAR SURVEY CONTROL 98
  10.6 DOWNHOLE SURVEY CONTROL 99
  10.7 DOWNHOLE SURVEY QUALITY CONTROL 101
  10.8 DIAMOND DRILL CORE PROCESSING 103
  10.9 LOGGING 103
  10.10 CORE RECOVERY 104
       
11 SAMPLE PREPARATION, ANALYSIS AND SECURITY 105
     
  11.1 REVERSE CIRCULATION DRILLING SAMPLING 105
  11.2 DIAMOND DRILLING SAMPLING 105
  11.2.1 DIAMOND DRILL CORE SAMPLES 106
  11.2.2 RELIABILITY OF SAMPLES 107
  11.3 ASSAY LABORATORIES 107
  11.3.1 STAWELL GOLD MINE LABORATORY 107
  11.3.2 WMC BALLARAT ASSAY LABORATORY 108
  11.3.3 AMDEL LABORATORY 108
  11.3.4 AMINYA LABORATORY 108
  11.3.5 INTERTEK GENALYSIS 108
  11.3.6 ALS LABORATORY GROUP 109
  11.4 SAMPLE PREPARATION 109
  11.4.1 SPLITTING USING A VIBRATING FEED CONE SPLITTER 109
  11.5 SAMPLE SECURITY 111
  11.6 ASSAY METHODS 111
  11.7 DATABASE STORAGE AND INTEGRITY 112
  11.8 QUALITY ASSURANCE/QUALITY CONTROL 112
  11.8.1 QA/QC CHECKS AND ACTIONS 113
    11.8.1.1 STANDARDS 113
    11.8.1.2 BLANKS 116
    11.8.1.3 LABORATORY SPLITS 117
  11.8.2 LABORATORY DUPLICATE ASSAYS 119
  11.8.3 STAWELL GOLD MINES STANDARD REFERENCE MATERIAL 122
  11.8.4 QA/QC FOR THIS TECHNICAL REPORT 122
  11.8.5 OPINIONS ON SAMPLING 124
  11.8.6 RECOMMENDATIONS 124
       
12 DATA VERIFICATION 125
     
13 MINERAL PROCESSING AND METALLURGICAL TESTING 126
     
  13.1 MINERAL PROCESSING 126
  13.2 METALLURGICAL TESTWORK 126
     
14 MINERAL RESOURCE ESTIMATES 129

iv

Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

  14.1 INTRODUCTION 129
    14.1.1 STOPE RECONCILIATION 133
  14.2 MANUAL 2D MINERAL RESOURCE AND MINERAL RESERVE ESTIMATION METHODOLOGIES 133
  14.3 COMPUTER 3D MINERAL RESOURCE AND MINERAL RESERVE ESTIMATION METHODOLOGIES 135
    14.3.1 INTRODUCTION 135
    14.3.2 DATA TYPES 135
    14.3.3 GEOLOGICAL INTERPRETATION 135
    14.3.3.1 FEDERAL ALBION AND FEDERAL ALBION SOUTH 135
    14.3.3.2 BELOW Scotchmans 250 136
    14.3.3.3 UPPER SOUTH FAULT 2 136
    14.3.3.4 MAGDALA S6000 137
    14.3.3.5 MARINERS 138
    14.3.3.6 BIG HILL AND UPPER LEVELS 139
    14.3.3.7 AURORA B 140
    14.3.4 GEOLOGICAL MODELLING 141
    14.3.5 WEATHERING MODEL 142
    14.3.6 MINERAL RESOURCE INTERPRETATION BY AREA 143
    14.3.6.1 FEDERAL ALBION SOUTH 143
    14.3.6.2 BELOW SM250 144
    14.3.6.3 UPPER SOUTH FAULT 2 144
    14.3.6.4 FEDERAL ALBION 145
    14.3.6.5 MAGDALA S6000 145
    14.3.6.6 MARINERS 146
    14.3.6.7 BIG HILL AND UPPER LEVELS 146
    14.3.6.8 AURORA B 147
    14.3.7 BLOCK MODEL DIMENSIONS 147
    14.3.8 BLOCK MODEL CODING 150
    14.3.9 DRILLHOLE CODING 150
    14.3.10 COMPOSITING AND STATISTIC 150
    14.3.11 GEOSTATISTICAL PARAMETERS 154
    14.3.12 BULK DENSITY 166
    14.3.13 MODEL VALIDATION 168
    14.3.14 MINERAL RESOURCE CLASSIFICATION 171
    14.3.15 LOGGING 171
    14.3.16 DATA SPACING AND DISTRIBUTION 171
    14.3.17 ORIENTATION OF DATA IN RELATION TO GEOLOGICAL STRUCTURE 172
    14.3.18 GEOLOGICAL INTERPRETATION 172
    14.3.19 DEPOSIT DIMENSIONS 172
    14.3.20 ESTIMATION AND MODELLING TECHNIQUES 172
    14.3.21 MOISTURE 172
    14.3.22 EXTERNAL FACTORS AFFECTING EXTRACTION 173
    14.3.23 BULK DENSITY 173
    14.3.24 CLASSIFICATION 173
    14.3.25 SELECTIVITY ASSUMPTIONS 173
    14.3.26 RESOURCE AUDITS OR REVIEWS 173
    14.3.27 DISCUSSION OF RELATIVE ACCURACY/CONFIDENCE 174
    14.3.28 MINERAL RESOURCE STATEMENT 174
    14.3.29 RECOMMENDATIONS 177
  14.4 LOW GRADE STOCKPILE 177
    14.4.1 MT MICKE 178
     
15 MINERAL RESERVES 180
   
  15.1 INTRODUCTION 180
  15.2 MINERAL RESERVE ESTIMATE UNDERGROUND 180
    15.2.1 MINERAL RESERVE ESTIMATE PROCESS 181

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Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

    15.2.2 MINE RESERVE DESIGN UNDERGROUND 181
    15.2.3 UNDERGROUND DESIGN AND RESERVE PARAMETERS 181
    15.2.4 GOLD CUT-OFF GRADES UNDERGROUND 185
    15.2.5 DEPLETION AND RESULTS 185
  15.3 MINERAL RESERVE ESTIMATE BIG HILL 186
    15.3.1 MINERAL RESERVE ESTIMATE 186
  15.4 MINERAL RESERVE ESTIMATE SURFACE STOCKPILES 188
     
16 MINING METHODS 190
     
  16.1 INTRODUCTION 190
  16.2 MINE DESIGN UNDERGROUND 190
    16.2.1 MINING METHOD DESCRIPTION 190
    16.2.2 UNDERGROUND MINING EQUIPMENT 191
  16.3 MINE DESIGN SURFACE MINING 192
    16.3.1 MINING METHOD DESCRIPTION 192
    16.3.2 SURFACE MINING SCHEDULE 193
    16.3.3 SURFACE MINING EQUIPMENT 193
17 RECOVERY METHODS 197
   
  17.1 MINERAL PROCESSING 197
     
18 PROJECT INFRASTRUCTURE 200
   
  18.1 SURFACE INFRASTRUCTURE 200
  18.2 TAILINGS AND STORAGE FACILITIES 200
  18.3 WASTE DUMPS 201
  18.4 UNDERGROUND MINE INFRASTRUCTURE 203
  18.5 POWER 203
  18.6 WATER 204
  18.7 EXISTING PUBLIC INFRASTRUCTURE 204
     
19 MARKET STUDIES AND CONTRACTS 206
   
  19.1 MARKETS 206
  19.2 GOLD PRICE 206
     
20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT 207
   
  20.1 ENVIRONMENTAL RESEARCH 207
  20.2 MONITORING 209
  20.3 PERMITTING 210
  20.4 COMMUNITY ENGAGEMENT 211
  20.5 REHABILITATION AND CLOSURE 211
     
21 CAPITAL AND OPERATING COSTS 213
   
  21.1 UNDERGROUND COSTS 213
    21.1.1 CAPITAL COSTS 213
    21.1.2 OPERATING COSTS 214
  21.2 SURFACE COSTS 215
    21.2.1 CAPITAL COSTS 215
    21.2.2 OPERATING COSTS 216
    21.2.2.1 LOAD AND HAUL OPERATING COSTS 217
    21.2.2.2 ANCILLARY EQUIPMENT OPERATING COSTS 220
    21.2.2.3 DRILL AND BLAST OPERATING COSTS 220
    21.2.2.4 EQUIPMENT LEASING AND HIRE 221

vi

Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

22 ECONOMIC ANALYSIS 222
     
  22.1 UNDERGROUND ECONOMIC ANALYSIS 222
  22.2 BIG HILL ECONOMIC ANALYSIS 222
    22.2.1 PRODUCTION AND REVENUE 222
    22.2.2 OPERATING COSTS 222
    22.2.3 CASHFLOWS 223
       
23 ADJACENT PROPERTIES 226
     
24 OTHER RELEVANT DATA AND INFORMATION 227
     
  24.1 BIG HILL MINERAL RESOURCE ESTIMATE 230
  24.2 BIG HILL MINERAL RESERVE ESTIMATE 231
  24.3 GEOLOGY, MINING, METALLURGY AND PROCESSING 232
  24.4 ENVIRONMENTAL, PERMITTING, SOCIAL AND COMMUNITY CONSIDERATIONS 234
       
25 INTERPRETATION AND CONCLUSIONS 236
     
26 RECOMMENDATIONS 238
     
27 REFERENCES 239
     
28 SIGNATURE PAGE 242

vii

Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

TABLES   PAGE
     
TABLE 1-1 STAWELL GOLD MINES MINERAL RESOURCE AS AT DECEMBER 31, 2015 5
TABLE 1-2 MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015  
TABLE 1-3 UNDERGROUND MINERAL RESOURCE ESTIMATION AS AT DECEMBER 31, 2015 6
TABLE 1-4 UNDERGROUND MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015 7
TABLE 1-5 BIG HILL SURFACE MINERAL RESOURCE AS AT DECEMBER 31, 2015 7
TABLE 1-6 BIG HILL SURFACE MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015 8
TABLE 1-7 LOW GRADESTOCKPILE MINERAL RESOURCE AS AT 31 DECEMBER 2015 8
TABLE 1-8 LOW GRADE STOCKPILE MINRAL RESERVE CLASSIFICATION AS AT 31 DECEMBER 2015 9
TABLE 2-1 TECHNICAL REPORTING RESPONSIBILITIES 14
TABLE 2-2 DEFINITIONS AND ABBREVTIATIONS 16
TABLE 3-1 SITE EXPERTS WHO CONTRIBUTED TO THE TECHNICAL REPORT 23
TABLE 9-1 EXPENDITURE ($AUD) BY YEAR FOR THE PERIOD JANUARY 2015 TO DECEMBER 2015 72
TABLE 9-2 SUMMARY OF PAST EXPLORATION ON REGIONAL PROSPECTS WITHIN SGM HELD TENEMENTS 77
TABLE 9-3 FORECAST EXPLORATION DRILLING ($A) FOR 2016 79
TABLE 9-4  PROPOSED EXPLORATION EXPENDITURE (A$1,000,000) FOR THE PERIOD 2017- 2018 81
TABLE 9-5 REGIONAL TENEMENT INFORMATION 86
TABLE 10-1 STAWELL GOLD MINES GEOLOGICAL PROCESSES AND APPROXIMATE DRILL SPACINGS 90
TABLE 11-1 DRILL STATISTICS FOR STAWELL GOLD MINES UNDERGROUND DURING APRIL 2012 -DECEMBER 2015 106
TABLE 11-2 DIAMOND DRILL CORE SAMPLE INTERVAL STATISTICS FOR SAMPLES TAKEN DURING 2015 107
TABLE 11-3 RANGE OF STANDARDS USED AT STAWELL GOLD MINES (G/T AU) 114
TABLE 11-4 QA/QC ASSAY RESULTS 123
TABLE 11-5 REPEATED ASSAY FOR ALL CONTRIBUTING MINERAL RESOURCE AREAS 123
TABLE 13-1 STAWELL GOLD MINES RECOVERY IN PERCENT AND PREG-ROB INDEX BY ORE SOURCE 127
TABLE 13-2 ACTUAL VERSUS EXPECTED RECOVERY – ALL ORE TYPES MONTHLY RECOVERY RECONCILIATION 127
TABLE 13-3 METALLURGICAL LEACH TEST WORK RESULTS FOR FEDERAL ALBION ORE SAMPLES 128
TABLE 14-1 SUMMARY OF MINERAL RESOURCE MODELS UPDATED IN 2015 129
TABLE 14-2 SUMMARY OF RESOURCE MODELS COMPLETED PRIOR TO 2015 THAT CONTRIBUTE TO THE MINERAL RESOURCE 129
TABLE 14-3 SUMMARY OF ORE SOURCE CONTRIBUTION FOR 2015 133
TABLE 14-4 BLOCK MODEL DIMENSIONS AND MODEL SET UP 149
TABLE 14-5 STATISTICAL SUMMARY BY AREA 152
TABLE 14-6 STATISTICAL SUMMARY FOR HIGH CUT COMPOSITES, GOLD G/T AU 153
TABLE 14-7 STAWELL GOLD MINES GEOSTATISTICAL SEARCH PARAMETERS 159
TABLE 14-8 STAWELL GOLD MINES GEOSTATISTICAL VARIOGRAM PARAMETERS 165
TABLE 14-9 COMPILATION OF DENSITY G/CM3,APPLIED BY RESOURCE MODEL AREA 168
TABLE 14- 10 MINERALIZED DOMAINS MODEL VALIDATION 171
TABLE 14- 11 STAWELL GOLD MINES RESOURCE AS AT 31 DECEMBER 2015 174
TABLE 14- 12  STAWELL GOLD MINES UNDERGROUND RESOURCE AS AT 31 DECEMBER 2015  175
TABLE 14- 13 BIG HILL SURFACE RESOURCE ESTIMATION AS AT DECEMBER 31, 2015 176
TABLE 14-14 LOW GRADE STOCKPILE RESOURCE ESTIMATION AS AT 31 DECEMBER 2015 178
TABLE 15-1 STAWELL GOLD MINES MINERAL RESERVE CLASSIFICATION – EFFECTIVE DECEMBER 31, 2015 180
TABLE 15-2 MINING METHOD SELECTION 183
TABLE 15-3 RECOVERY AND DILUTION FACTORS FOR THE UNDERGROUND RESERVE BLOCKS AT STAWELL GOLD MINES 184 
TABLE 15-4 GOLD CUT-OFF GRADES UNDERGROUND 185
TABLE 15-5 MINABLE SHAPE OPTIMISER – BIG HILL DILUTION 187
TABLE 15-6 GOLD CUT-OFF GRADES, BIG HILL 188
TABLE 15-7 LOW GRADESTOCKPILE CUT OFF CALCULATION 189
TABLE 16-1 UNDERGROUND MINING MAJOR MOBILE EQUIPMENT– EFFECTIVE DECEMBER 31, 2015 192
TABLE 16-2 PRODUCTIVITY ESTIMATES 194

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Newmarket Gold
Stawell Gold Mines

TABLE 17-1 STAWELL GOLD MINES TOTAL GOLD RECOVERED JANUARY 2015 TO DECEMBER 2015 199
TABLE 20-1 ENVIRONMENTAL MONITORING REQUIREMENTS FOR STAWELL GOLD MINES 210
TABLE 21-1 SGM 2016 CAPITAL COST SUMMARY 213
TABLE 21-2 SGM 2016 OPERATING COST SUMMARY 214
TABLE 21-3 CAPITAL COST ESTIMATE BIG HILL 215
TABLE 21-4 OPERATING COST ESTIMATE BIG HILL 216
TABLE 22-1 BIG HILL PRODUCTION AND REVENUE SUMMARY 222
TABLE 22-2 BIG HILL OPERATING COST SUMMARY 223
TABLE 24-1 BIG HILL FINANCIAL ANALYSIS RESULTS BASED ON A USD$1,225/OZ PRICE AND A US$:A$ EXCHANGE RATE OF 0.87 227
TABLE 24-2 CAPITAL COST SUMMARY 227
TABLE 24-3 OPERATING COST SUMMARY 227
TABLE 24-4 UNIT COSTS SUMMARY 228
TABLE 24-5 OPERATING PLAN SUMMARY 228
TABLE 24-6 BIG HILL PROCESSING SUMMARY 228
TABLE 24-7 GOLD PRICE SENSITIVITY ANALAYSIS 228
TABLE 24-8 COST SENSITIVITY ANALYSIS 228
TABLE 24-9 CAPITAL COST SUMMARY AND PRE-PRODUCTION CAPITAL COSTS 229
TABLE 24- 10  OPERATING COSTS SUMMARY 230
TABLE 24- 11  BIG HILL MINERAL RESOURCES ESTIMATE AS OF MARCH 2014 230
TABLE 24- 12  BIG HILL ORE RESERVES AS OF MARCH 2014 231
TABLE 24- 13  ADDITIONAL OPERATING COSTS FOR COMMUNITY AND ENVIRONMENTAL CONSIDERATIONS 235

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Stawell Gold Mines

FIGURES PAGE
   
FIGURE 1-1 MAP HIGHLIGHTING LOCATION OF STAWELL GOLD MINES 3
FIGURE 2-1 RELATIONSHIP BETWEEN MINERAL RESERVES AND MINERAL RESOURCES 19
FIGURE 4-1 LOCATION MAP OF THE MIN5260 LEASE. THE GRID IS LATITUDE AND LONGITUDE (GDA94) 26
FIGURE 4-2 AERIAL VIEW OF THE MIN5260 LEASE INDICATING UNDERGROUND AND PIT LOCATIONS 27
FIGURE 4-3 LOCAL MINE GRID USED AT STAWELL GOLD MINES 31
FIGURE 6-1 ANNUAL GOLD PRODUCTION SINCE 1984 36
FIGURE 6-2 ANNUAL PROCESSING PLANT PRODUCTION SINCE 1984 37
FIGURE 7-1 IMAGE SHOWING LACHLAN FOLD BELT, LOCATING STAWELL ON THE WESTERN BOUNDARY 39
FIGURE 7-2 D1 TO D5 DUCTILE AND BRITTLE EVOLUTION OF THE STAWELL SYSTEM. FROM MILLER ET AL. 2006 43
FIGURE 7-3 EVOLUTION OF THE STAWELL SYSTEM FROM 420 TO 380 MA (MODIFIED FROM MILLER & WILSON 2004A) 4
FIGURE 7-4 FMINE GEOLOGY CROSS-SECTION HIGHLIGHTING ARCHITECTURE OF THE MAGDALA AND GOLDEN GIFT
OREBODIES
46
FIGURE 7-5 PLAN VIEW GEOLOGICAL INTERPRETATION OF THE STAWELL STRUCTURAL AND STRATIGRAPHIC ARCHITECTURE AT 1000MRL 47
FIGURE 7-6 STAWELL GOLD MINES WEST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF THE
MINERALIZED ORE BLOCKS 
48
FIGURE 7-7 STAWELL GOLD MINES EAST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF THE MINERALIZED
ORE BLOCKS 
49
FIGURE 7-8 EXAMPLE OF CENTRAL LODE MINERALIZATION 50
FIGURE 7-9 EXAMPLE OF BASALT CONTACT MINERALIZATION 51
FIGURE 8-1 SCHEMATIC CROSS SECTION SHOWING LOCATION OF FEDERAL ALBION SOUTH AREA 54
FIGURE 8-2 FTYPICAL CROSS-SECTION THROUGH FEDERAL ALBION AREA SHOWING CURRENT INTERPRETED GEOLOGICAL
SETTING 
55
FIGURE 8-3 GENERAL CROSS SECTION OF THE BELOW SCOTCHMANS AREA 57
FIGURE 8-4 GENERAL CROSS SECTION OF THE USF2 AREA 58
FIGURE 8-5 TYPICAL CROSS-SECTIONS THROUGH MARINERS AREA SHOWING CURRENT INTERPRETED GEOLOGICAL SETTING 60
FIGURE 8-6 SECTION THROUGH MARINERS AND ALLENS LODES STRUCTURAL COMPLEXITY AND OFFSETTING 61
FIGURE 8-7 SECTION 5840 THROUGH SOUGH PIT SHOWING IRON DUKE AND MAGDALA LODES 62
FIGURE 8-8 SECTION THROUGH THE MAGDALA MINERALIZATION SHOWING THE INTERPRETED GEOLOGY BELOW 63
FIGURE 8-9 TYPICAL CROSS-SECTION THROUGH AURORA B SHOWING CURRENT INTERPRETED GEOLOGICAL SETTING 64
FIGURE 9-1 MINE GEOLOGY LONGITUDINAL SECTION OUTLINING NEAR MINE EXPLORATION ON WEST FLANK FROM JANUARY 2015 TO DECEMBER 2015 66
FIGURE 9-2 MINE GEOLOGY LONGITUDINAL SECTION OUTLINING NEAR MINE EXPLORATION ON EAST FLANK FROM JANUARY 2015 TO DECEMBER 2015 67
FIGURE 9-3 PROJECT LOCATION PLAN 68
FIGURE 9-4 MINE GEOLOGY LONGITUDINAL SECTION OUTLINGING PROPOSED NEAR MINE EXPLORATION FOR 2016, WEST FLANK 70
FIGURE 9-5 MINE GEOLOGY LONGITUDINAL SECTION OUTLINGING PROPOSED NEAR MINE EXPLORATION FOR 2016, EAST FLANK 71
FIGURE 9-6 REGIONAL TENEMENTS ELS 3008, 5443 AND 5474 SUPERIMPOSED ON AEROMAGNETIC TMI IMAGE OUTLINING REGIONAL EXPLORATION PROSPECTS 78
FIGURE 9-7 REGIONAL TENEMENTS ELS 3008, 5443 AND 5474 SUPERIMPOSED ON AEROMAGNETIC TMI IMAGE OUTLINING ONGOING REGIONAL EXPLORATION FOR 2016 80
FIGURE 9-8 STAWELL GOLD MINES REGIONAL VICTORIA TENEMENT AREAS 85
FIGURE 10-1 SGM DIAMOND DRILLING PROCESS FLOWSHEET 92
FIGURE 10-2 TYPICAL SLUDGE DRILLING FANS COMPLETED FROM AN ORE DEVELOPMENT DRIVE 94
     
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Stawell Gold Mines

FIGURE 10-3 LONGITUDINAL SECTION VIEW (TOP), CROSS SECTION VIEW (BOTTOM LEFT) AND PLAN VIEW (BOTTOM RIGHT) OF THE FAS MINERALIZED DOMAIN AND ALL UNDERGROUND DIAMOND DRILLHOLES USED TO CONSTRAIN THE MOST RECENT FAS MINERAL RESOURCE ESTIMATE 95
FIGURE 10-4 PLAN VIEW (TOP LEFT), CROSS SECTION VIEW (TOP RIGHT) AND LONGITUDINAL SECTION VIEW (BOTTOM) OF THE USF2 MINERALIZED DOMAINS SHOWING ALL UNDERGROUND DIAMOND DRILLHOLES USED TO CONSTRAIN THE MOST RECENT MINERAL RESOURCE ESTIMATE 96
FIGURE 10-5 PLAN VIEW (TOP) AND CROSS SECTION VIEWS (A-B AND C-D) 97
FIGURE 10-6 PLAN SHOWING BIG HILL PIT USED FOR RESOURCE REPORTING, DRILL COVERAGE AND DRILL ORIENTATIONS  98
FIGURE 10-7 PLOT OF DRILLHOLE SURVEY METHOD 100
FIGURE 10-8 MAGNETIC DECLINATION CORRECTION AS CURRENTLY APPLIED TO STAWELL GOLD MINES DRILLHOLE DATA  101
FIGURE 10-9 EXAMPLE OF THE CHECK PLOTS USED TO CORRECT DOWNHOLE SURVEY AZIMUTH INFORMATION 102
FIGURE 10-10 AN EXAMPLE OF THE DIAMOND DRILLCORE PHOTOGRAPHS STORED DIGITALLY FOR ALL DIAMOND DRILLCORE 103
FIGURE 11-1 STAWELL GOLD MINES DRILL CORE SAMPLE PREPARATION, ASSAY AND QA/QC FLOWSHEET 110
FIGURE 11-2 ASSAY STANDARD PERFORMANCE FOR FAS DATA (STANDARD DEVIATION TO THE EXPECTED VALUE) G/T AU  115
FIGURE 11-3 OR200 STANDARD PERFORMANCE FOR FAS MAY 2015 RESOURCE MODEL UPDATE 116
FIGURE 11-4 OR204 STANDARD PERFORMANCE FOR FAS MAY 2015 RESOURCE MODEL UPDATE 116
FIGURE 11-5 BLANK STANDARD PERFORMANCE FOR FAS MAY 2015 RESOURCE MODEL UPDATE 117
FIGURE 11-6 PRECISION OF THE CRUSHER FOR THE MAY 2015 FAS RESOURCE MODEL UPDATE 118
FIGURE 11-7 SAMPLE SPLITS (CRUSH SIZE FRACTION SPLIT) COMPARISON FOR MAY 2015 FAS RESOURCE MODEL UPDATE  118
FIGURE 11-8 PRECISION OF THE LAB DUPLICATE DATA FOR MAY 2015 FAS RESOURCE MODEL UPDATE 120
FIGURE 11-9 ASSAY DUPLICATE COMPARISON OF THE LAB DUPLICATE DATA FROM MAY 2015 FAS RESOURCE MODELUPDATE 120
FIGURE 11-10 LOG-LOG GRAPHICAL REPRESENTATION OF THE ASSAY REPEATS FROM MAY 2015 FAS RESOURCE MODEL UPDATE 121
FIGURE 14-1 WEST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF MINERAL RESOURCE AND MINERAL RESERVE AREAS AS OF 31 DECEMBER 2015 131
FIGURE 14-2 EAST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF MINERAL RESOURCE AND MINERAL RESERVE AREAS AS OF 31 DECEMBER 2015  132
FIGURE 14-3 STOPE RECONCILIATION 133
FIGURE 14-4 CLASSIFICATION OF MINED MATERIAL FOR 2015 134
FIGURE 14-5 LONGSECTION OF BIG HILL DOMAIN 601 SHOWING COMPOSITES WHICH HAVE A PLUNGE 154
FIGURE 14-6 PLAN OF BIG HILL DOMAIN BIG 602 SHOWING COMPOSITES WHICH HAVE NO PREFERRED PLUNGE 155
FIGURE 14-7 LOCATION OF STAWELL LOW-GRADE STOCKPILE 179
FIGURE 16-1 BIG HILL SURFACE MINING METHOD SCHEMATIC 192
FIGURE 16-2 BIG HILL SURFACE MINING SCHEDULE 193
FIGURE 16-3 EXCAVATOR REQUIREMENTS BIG HILL 195
FIGURE 16-4 HAUL TRUCK REQUIREMENTS BIG HILL 196
FIGURE 17-1 SGM TREATMENT PLANT FLOW SHEET 198
FIGURE 18-1 PLAN SHOWING THE LOCATION OF MIN 5260, STAWELL GOLD MINES OPERATIONAL INFRASTRUCTURE 202
FIGURE 21-1 EXCAVATOR REQUIREMENTS BIG HILL 219
FIGURE 21-2 HAUL TRUCK REQUIREMENTS BIG HILL 220
FIGURE 22-1 BIG HILL OPERATING EXPENDITURE PROFILE 223
FIGURE 22-2 BIG HILL CASHFLOW PRE-TAX NPV 224
FIGURE 22-3 BIG HILL FOUR CASH STREAMS 225
FIGURE 24-1 BIG HILL SURFACE PLAN 233

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December 2015
Newmarket Gold
Stawell Gold Mines

  1

EXECUTIVE SUMMARY

This technical report has been prepared for Newmarket Gold Inc. (Newmarket Gold), the beneficial owner of the Stawell Gold Mines. Newmarket Gold is listed on the Toronto Stock Exchange (NMI).

It has been prepared in accordance with the requirements of the National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101) and updates material changes to the Mineral Resource and Mineral Reserve position as of December 31, 2015.

The Mineral Resources and Mineral Reserve estimate for Stawell Gold Mines is a summation of a number of individual estimates for various orebodies and geographically constrained areas. All of these estimates are contained within the Mining Lease MIN5260.

The Stawell Gold Deposit was discovered in the mid 1850’s during the Victorian gold rush, which saw the discovery and exploitation of significant deposits at Bendigo and Ballarat. Mining activity eventually ceased in the 1920’s and after a prolonged period of sporadic exploration, mining operation recommenced in 1981. Mining operations and various levels of exploration and resource development activities have been continuous since 1981 and as such the project has significant past production and development history, which is discussed in this technical report and also utilized during the compilation of the Mineral Resource and Mineral Reserve estimates.

Since the publication of the previous NI 43-101 Technical Report on the Stawell Gold Mines prepared by Justine Tracey, MAusIMM (CP) and Wayne Chapman, MAusIMM (CP) dated March 24, 2015 (the 2014 Stawell Technical Report), Newmarket Gold has drilled and re-estimated the Mineral Resource for some of the deposits within the Stawell underground mine. Using these models, Mineral Reserves have been calculated and are reported within for those Mineral Resources that warrant the financial considerations required to fulfill the requirements of a Mineral Reserve.

In June 2014 the Big Hill Surface Resource and Reserve were reported in the NI 43-101 Technical Report on the Big Hill Enhanced Development Project at Stawell Gold Mines prepared by Dean Basile, MAusIMM (CP) and Stuart Hutchin, MAIG, MAusIMM, dated June 2014 (the Big Hill Technical Report). The ongoing reporting of this Mineral Resource and Reserve is contained within this report.

This technical report has been prepared by a number of the Stawell Gold Mines personnel. The report utilizes information available within Newmarket Gold’s technical reports, published geological papers and internal Mineral Resource and Mineral Reserve documents completed by members of the Stawell Gold Mines mine geological and mine engineering teams.

Justine Tracey, Mark Edwards and Wayne Chapman are qualified persons as defined by NI 43-101 and accept overall responsibility for the preparation of all sections of this technical report including the preparation of the Mineral Resources as reported in Section 14. All information presented in this technical report was prepared in accordance with the requirements of NI 43-101 and is in the format prescribed by that instrument.

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Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

  1.1

PROJECT DESCRIPTION AND OWNERSHIP

Stawell Gold Mines is located in the Australian State of Victoria, 250km northwest of Melbourne and 2km from the Township of Stawell, Figure 1-1.

Stawell Gold Mines principal approval is its Mining Lease MIN5260, (see Figure 4-2) issued by the Victorian State Government under the Sustainable Development Act. This MIN5260 lease (centroid coordinates of 142.80° E and 37.06° S, GDA94) encompasses both the Magdala and Wonga Mines and is located both under and around the Township of Stawell with an area of approximately 1,000.58 Ha.

Stawell is a historic goldfield that produced 2.7 million ounces of gold between 1853 and 1926 from both alluvial and hard rock sources. In 1981, Stawell Gold Mines was re-opened by the Western Mining Corporation (WMC)/Central Norseman Gold joint venture with commencement of the Magdala decline. By 1984, the operation had expanded with the construction of a processing facility and subsequent commencement of an open cut operation at the Wonga Mine (2 kilometers south of Magdala). The Wonga Open Cut operated from 1984 to 1987 and produced 778,847 tonnes recovering 69,159 ounces of gold. The Davis Open Cut operated from 1987 to 1989 and produced 154,525 tonnes for 8,992 recovered ounces of gold.

In December 1992, the operation was acquired in a 50/50 joint venture by Mining Project Investors Pty Ltd. (MPI) and Pittston Mineral Ventures (Pittston). The joint venture continued until 2004, during which time there was a record of continued exploration success with discovery of additional mineralized deposits that were subsequently mined.

In February 2004, MPI acquired Pittston’s 50% share of the project. In November 2004, a de-merger of the MPI gold business came into effect, and Leviathan Resources Ltd. (Leviathan) was floated in December 2004. The resource drilling into the Golden Gift Deposit initially identified seven areas of mineralization offset from each other due to late faulting. From the increased geological understanding of the Golden Gift Deposit, it was clear in the mine planning process that two declines were required, the GG5 and GG3 declines, to access the ore zones for continuity of supply.

In January 2007 Perseverance Corporation Limited (Perseverance) acquired Leviathan. Perseverance was acquired by Northgate Minerals Corp. (Northgate) on February 18, 2008. Northgate was acquired by AuRico Gold Inc. (AuRico) in October 2011. Crocodile Gold Corp (Crocodile Gold) completed their acquisition of Stawell Gold Mines from AuRico on May 4, 2012. On July 14, 2015 a merger between Newmarket Gold Inc. and Crocodile Gold was completed to form Newmarket Gold.

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Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

FIGURE 1-1 MAP HIGHLIGHTING LOCATION OF STAWELL GOLD MINES

  1.2

GEOLOGY & MINERALIZATION

The Stawell Goldfield is located in the western Stawell Zone of the Lachlan Fold Belt. Interpretations from the Victorian Geological Survey present a thin skinned tectonics model where the Moyston Fault is an east dipping basal detachment, which has juxtaposed higher metamorphic grade rocks of the Stawell Zone against lower grade Cambrian rocks of the Delamarian Glenelg Zone. The west dipping Stawell Fault, Coongee Break and other parallel west dipping faults represent back thrusts from the Moyston Fault. The Stawell-Wildwood corridor therefore represents a significant structural high in an up-thrown block of deeper stratigraphy between the Coongee Break and Pleasant Creek Fault.

Intruded into this sequence are the Stawell Granite and a number of felsic and mafic intrusions. The stratigraphy at Stawell is divided into three principal units: Magdala Basalt, Albion Formation and Leviathan Formation. The dominant feature at Stawell is the 1.2km wide, doubly plunging, northwest-striking Magdala Basalt dome. The Magdala Basalt is made up of a series of basalt noses, interpreted to be flow sheets, which now dip to the southwest and plunge to the northwest. Areas of sedimentation are present between the basalt noses and are locally termed ‘waterloos’.

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Technical Report
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Newmarket Gold
Stawell Gold Mines

There are three mineralization styles at Stawell, being Magdala (separated into west and east flanks), Golden Gift and Wonga. The Magdala and Golden Gift ore types are hosted within the Magdala Volcanogenic. Within the Magdala Deposit there are three main ore types; Central Lode, Basalt Contact Lodes, and Magdala Stockwork Lodes. The east flank mineralization introduces a new ore type: the Hampshire Lode.

Gold occurs as free gold, frequently associated with pyrite, pyhrotite and arsenopyrite.

  1.3

EXPLORATION, DEVELOPMENT AND OPERATIONS

Since the closure of the lower portion of Stawell Gold Mines in 2013, production has focused on mining of remnant Mineral Resource and mining pillars in areas above 1065 mRL. The underground mining plan over this period has been based on an extraction target of around 400 - 500kt/pa with a grade range of 2.0 - 3.0 g/t Au. Surface mining plans include the Big Hill project area with ongoing permitting considerations through 2016.

The 2016 mine budget based upon higher confidence material consists of development, mining and processing of 495kt at 2.53 g/t Au of underground mill feed supplemented by 442kt of surface oxide stockpiles for production of 38k ounces.

The mine plan is based upon similar Mineral Resource conversion and discovery as the previous three years, and results in an ongoing underground mine scenario while surface mining operations undergo permitting.

Exploration is carried out in the near mine environment in order to supplement resources that can be accessed from current underground workings with a minimum of development work required. Underground and surface exploration programs are ongoing and have recently been successful in adding inferred Mineral Resources in the Aurora B zone.

  1.4

MINERAL RESOURCES AND MINERAL RESERVES


  1.4.1

STAWELL GOLD MINES

The total Mineral Resource estimate for the Stawell Gold Mines operations is listed in Table 1-1.

This Stawell Gold Mines Mineral Resource is categorized as Underground (Table 1-3), Big Hill Surface (Table 1-5) and Low Grade Stockpiles (Table 1-7). These are all located on the same Mining Lease and share the Processing Facility.

 Stawell Gold Mines Resource 
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Measured 56 2.56 5
Indicated 4,063 1.85 241
Total (Measured and Indicated only) 4,119 1.86 246
Inferred 1,164 3.16 118

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Technical Report
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Newmarket Gold
Stawell Gold Mines

TABLE 1-1 STAWELL GOLD MINES MINERAL RESOURCE AS AT DECEMBER 31, 2015

NOTES:

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:


  a.

Gold price of A$1,500/oz

     
  b.

Cut-off Grade applied was variable for surface Mineral Resources. Grades used were as follows:


  0.44 g/t for Big Hill surface
  0.35 g/t for surface Low Grade Stockpiles

  c.

Cut-off Grade applied was variable for underground Mineral Resources. Grades used were as follows:


  2.0 g/t for Mariners and Big Hill outside of current pit optimisation
  2.3 g/t for all remaining underground Mineral Resources

4.

Surface and Underground Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines. Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Mineral Resources under NI 43-101.

   
5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

The quantity and grade of the reported inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to define the inferred Mineral Resources as indicated or measured Mineral Resources and it is uncertain if further exploration will result in upgrading them to an indicated or measured Mineral Resources category.

   
7.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
8.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Mineral Reserve estimate for the Stawell Gold Mines Operations is listed in Table 1-2. This total Mineral Reserve is categorized as Underground (Table 1-4), Big Hill Surface (Table 1-6) and Low Grade Stockpiles (Table 1-8).

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven 51 2.49 4
Probable 3,428 1.46 162
Total Mining Reserve 3,479 1.48 166

TABLE 1-2 MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by NI 43-101.

   
2.

Mineral Reserves were estimated using the following economic parameters:


  a.

Gold price of A$1,450/oz

     
  b.

Cut-off Grade applied was variable for underground depending upon width, mining method and ground conditions.

     
  c.

Cut-off Grade applied to Big Hill surface was 0.4 g/t Au

     
  d.

Cut-off Grade applied to surface LG Stockpiles was 0.35 g/t Au

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Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

3.

Underground Mineral Reserve estimates were prepared by Stawell Gold Mines personal under the guidance of Wayne Chapman, Technical Manager Stawell Gold Mines. Mr Chapman is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 11 years of relevant mining engineering experience and is the Qualified Person for Mineral Reserves under NI 43-101.

   
4.

Big Hill Surface Mineral Reserve estimates were prepared by Mining One personal under the guidance of Mark Edwards. Mr Edwards is the General Manager for Exploration, Newmarket Gold and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 11 years of relevant mining engineering experience and is the Qualified Person for Mineral Reserves under NI 43-101.

   
5.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.


  1.4.2

UNDERGROUND

The Mineral Resource estimate for the Stawell underground mine is listed in Table 1-3.

Stawell Underground Mineral Resource 
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Measured 56 2.56 5
Indicated 669 3.49 75
Total (Measured and Indicated only) 725 3.42 80
Inferred 1,118 3.24 116

TABLE 1-3 UNDERGROUND MINERAL RESOURCE ESTIMATION AS AT DECEMBER 31, 2015

NOTES:

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:


  a.

Gold price of A$1500/oz

     
  b.

Cut-off Grade applied was variable for surface Mineral Resource . Grades used were as follows:


  0.44 g/t for Big Hill surface
  0.35 g/t for surface Low Grade Stockpiles

  c.

Cut -off Grade applied was variable for underground Mineral Resources. Grades used were as follows:


  2.0 g/t for Mariners and Big Hill outside of current pit optimisation
  2.3 g/t for all remaining underground Mineral Resources

4.

Surface and Underground Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines. Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Mineral Resources under NI 43-101.

   
5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

The quantity and grade of the reported inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to define the inferred Mineral Resources as indicated or measured Mineral Resources and it is uncertain if further exploration will result in upgrading them to an indicated or measured Mineral Resource category.

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Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

7.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
8.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Mineral Reserve estimate for the Stawell underground mine is listed in Table 1-4.

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven 51 2.49 4
Probable 305 2.47 24
Total Mining Reserve 356 2.47 28

TABLE 1-4 UNDERGROUND MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by NI 43-101.

   
2.

Mineral Reserves were estimated using the following economic parameters:


  a.

Gold price of A$1,450/oz

     
  b.

Cut-off Grade applied was variable for underground depending upon width, mining method and ground conditions.


3.

Underground Mineral Reserve estimates were prepared by Stawell Gold Mines personal under the guidance of Wayne Chapman, Technical Manager Stawell Gold Mines. Mr Chapman is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 11 years of relevant mining engineering experience and is the Qualified Person for Mineral Reserves under NI 43-101.

   
4.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.


  1.4.3

BIG HILL SURFACE

The Mineral Resource estimate for the Surface Big Hill Project at Stawell Gold Mines is listed in Table 1-5.

 Stawell Underground Resource 
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Measured      
Indicated 2,971 1.68 160
Total (Measured and Indicated only) 2,971 1.68 160
Inferred 46 1.15 2

TABLE 1-5 BIG HILL SURFACE MINERAL RESOURCE AS AT DECEMBER 31, 2015

NOTES:

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:


  a. Gold price of A$1,500/oz
  B. Cut-off Grade applied for Big Hill Surface Mineral Resource is 0.44 g/t Au

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Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

4

Surface Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines. Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Mineral Resources under NI 43-101.

   
5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

The quantity and grade of the reported inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to define the inferred Mineral Resources as indicated or measured Mineral Resources and it is uncertain if further exploration will result in upgrading them to an indicated or measured Mineral Resource category.

   
7.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
8.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Mineral Reserve estimate for the Big Hill surface project is listed in Table 1-6.

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven      
Probable 2,700 1.51 132
Total Mining Reserve 2,700 1.51 132

TABLE 1-6 BIG HILL SURFACE MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by NI 43-101.

   
2.

Mineral Reserves were estimated using the following economic parameters:


  a.

Gold price of A$1,450/oz

     
  b.

Cut-off Grade applied was 0.4 g/t Au


3.

Big Hill Surface Mineral Reserve estimates were prepared by Mining One personal under the guidance of Mark Edwards, General Manager Exploration Newmarket Gold. Mr Edwards is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 18 years of relevant mining experience and is the Qualified Person for Mineral Reserves under NI 43-101.

   
4.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.


  1.4.4

LOW GRADE STOCKPILES

The Mineral Resource estimate for the Stawell surface low grade stockpiles is listed in Table 1-7.

Stawell Low Grade Stockpile Mineral Resource 
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Measured - - -
Indicated 423 0.43 6
Total (Measured and Indicated only) 423 0.43 6
Inferred - - -

TABLE 1-7 LOW GRADESTOCKPILE MINERAL RESOURCE AS AT 31 DECEMBER 2015

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NOTES:

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:


  a.

Gold price of A$1,500/oz

     
  b.

Above 0.35 g/t Au cut-off


4.

Surface Low Grade Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines. Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Mineral Resources under NI 43-101.

   
5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
7.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Mineral Reserve estimate for Stawell Surface Low Grade Stockpiles is listed in Table 1-8.

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven - - -
Probable 423 0.43 6
Total Mining Reserve 423 0.43 6

TABLE 1-8 LOW GRADE STOCKPILE MINRAL RESERVE CLASSIFICATION AS AT 31 DECEMBER 2015

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Reserves were estimated using the following economic parameters:


  a.

Gold price of A$1,450/oz

     
  b.

Above 0.35 g/t Au cut-off


3.

Surface Low Grade Mineral Reserve estimates were prepared by Stawell Gold Mines personal under the guidance of Wayne Chapman, Technical Manager Stawell Gold Mines. Mr Chapman is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 11 years of relevant mining engineering experience and is the Qualified Person for Reserves under NI 43-101.

   
4.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

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  1.5

INTERPRETATION

The data on which this updated Mineral Resource and Mineral Reserve statement is based has been collected utilizing the quality systems, procedures and processes. There are very high standards of data collection and storage utilized at Stawell Gold Mines and the assay quality is supported by QA/QC documentation and verifiable data. Data management systems are in place to ensure long term security of all geological information collected on site.

The gold grade estimates are based on high quality assay datasets of diamond drill core that has been spatially located, sampled and assayed using sound industry standard practices.

There is available an extensive coverage of diamond drilling reaching a drill spacing of 15m x 15m in areas that are subject to grade control drilling. Additionally face mapping information and “sludge sample” holes logged for geology are available to construct geological models for all Mineral Resource areas. The key control on Mineral Resource estimation is an accurate definition of the constraining geological models. Estimation of grade within the domains, whilst still very important, is of secondary importance to the first order geological domaining. The geological personnel have a sound understanding of the mineralized system and have good practices in place to ensure quality models are produced.

In addition to the quality control and data verification procedures, the Qualified Persons preparing the Mineral Resource estimates have further validated the data upon extraction from the database prior to resource interpolation. This verification used MineSight software as the primary tool to identify data problems. This allowed the omission of holes if they were of questionable quality, for example due to low quality sample techniques or incomplete assaying. When coupled with the more mechanical check processes ensuring high quality is entering the database in the first place, these checks were effective in allowing the Qualified Persons to be confident that the data was geologically coherent and of appropriate quality.

  1.6

CONCLUSION AND RECOMMENDATIONS

Ongoing mining operations in the upper levels of Stawell Gold Mines through both low grade resource remodelling and mining and adjusted operational costs have provided confidence in a mine budget targeting 38koz for 2016. Continued western flank low grade resource investigation, Big Hill surface mining permitting and underground eastern flank exploration provide opportunity beyond 2016.

The Stawell Mineral Resource increased in inventory over the past 12 months. A small increase in Indicated ounces in the Magdala Western Flank mineralisation does not reflect the extent of conversion during the year. This is a result of a short time frame between delivery of the resource and mine scheduling, thus material that was converted during the reporting period is already in production in the same reporting period.

An increase to the Inferred resource in the reporting year is result of targeted drilling of the unmined margins of the Magdala orebody (Federal Albion South, Below Scotchmans 250 and Upper South Fault 2) and exploration drilling on the Eastern Flank (Aurora B).

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Prior to the reporting period the a significant portion of the Inferred Mineral Resource inventory underwent conversion to Indicated Mineral Resource in order to sustain Mine Operations. Drill programs through 2015 have enabled this material to be replaced through the course of 2015. Continued drilling and investigation into the Inferred resource areas in 2016 gives capacity for conversion to Indicated and understand into along strike and down plunge potential to the mineralisation.

Mineral Reserve has matched depletion. Big Hill Surface Mineral Reserve was reduced in line with the adjusted Big Hill mining plan.

A summary of the recommendations:

A structural review of all available drill core to understand the down plunge position and potential of the mineralized shoots on the eastern flank Aurora B. Undertake further diamond drilling on the Eastern Flank for exploration of mineralization extent of the Aurora B zone.

Undertake infill diamond drilling on the upper southern extents of priority lodes on the western flank to confirm the assumptions of geological continuity inherent in the current estimate.

Continue to build geological models over the Magdala orebody where there is no current digital model to aid in targeting and geological understanding.

Undertake targeted resource drilling on the faulted extremities of the mineralization (above the Scotchmans Fault and below the South Fault).

Continue underground channel sampling and digital capture of the results to assist with determination of wireframe extents and aid the consideration of recoverable reserves.

Continue to review the performance of the Mineral Resource estimate through regular reconciliation between geological modelling, mining and the processing facility.

 

Continue further permitting applications for the Big Hill surface mining project.

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  2

INTRODUCTION AND TERMS OF REFERENCE


  2.1

INTRODUCTION

The purpose of this technical report on the Stawell Gold Mines is to support public disclosure of Mineral Resource and Mineral Reserve estimates at the Stawell Gold Mines as at December 31, 2015. This technical report is prepared in accordance with NI 43-101.

Newmarket Gold is a Canadian-listed gold mining and exploration company with three operating mines in Australia – the Fosterville and Stawell Gold Mines in the State of Victoria and Cosmo Gold Mine in the Northern Territory.

The Stawell Gold Deposit was discovered in the mid 1850’s during the Victorian gold rush, which saw the discovery and exploitation of significant deposits at Bendigo and Ballarat. Mining activity eventually ceased in the 1920’s and after a prolonged period of sporadic exploration, mining operations recommenced in 1981. Mining operations and various levels of exploration and Mineral Resource development activities have been continuous since 1981 and as such, the project has significant past production and development history which is discussed in this technical report and also utilized during the compilation of the Mineral Resource and Mineral Reserve estimates.

The Authors have prepared this technical report for Newmarket Gold in respect of the underground Mineral Resource and Reserve of the Stawell Gold Mines, which is located adjacent to the Township of Stawell, Western Victoria, Australia. The underground operation has been in operation since 1981 and has produced in excess of 2.31 million ounces of gold during this period.

Crocodile Gold (now Newmarket Gold) acquired the Stawell Gold Mines in May 2012 through an acquisition of AuRico’s Australian assets. Newmarket Gold owns a 100% interest in the Stawell Gold Mines including the Big Hill Project with an A$2.00 per gold ounce royalty payable to Minerals Ventures of Australia. Effective Januray 1st, 2016 Newmarket will pay Alamos Gold Inc. a 1% NSR royalty on all production from Stawell Gold Mines.

This technical report includes a geological overview of the Stawell Gold Mines, including a description of the geology, mineralization, key occurrences and deposits. It also provides an update on Mineral Resources and Mineral Reserves, and makes recommendations on additional exploration and development drilling which has the potential to upgrade Mineral Resource classifications and to augment the Mineral Resource base.

  2.2

TERMS OF REFERENCE

This technical report supersedes and updates the 2014 Stawell Technical Report and the Big Hill Technical Report. This technical report is inclusive of all Mineral Resource and Mineral Reserves contained within the Stawell Gold Mines operations.

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This technical report has been prepared in accordance with the requirements of the NI 43-101 and updates changes to the Mineral Resource and Mineral Reserve position of the Stawell Gold Mines as of December 31, 2015.

The Mineral Resources estimate for the Stawell Gold Mines is a summation of a number of individual estimates for various orebodies or various geographically constrained areas. All of these estimates are contained within the Mining Lease MIN5260. Details of the locations and geographical constraints of the various orebody components as of December 31, 2015 are given in Section 8. The Mineral Reserves estimate for the Stawell Gold Mines is a summation of a number of individual estimates for various methods of extraction of Mineral Resource areas contained within Mining Lease MIN5260.

  2.3

AUTHORS’ QUALIFICATIONS & RESPONSIBILITIES

This technical report conforms to the CIM Standards on Mineral Resources and Mineral Reserves referred to in NI 43-101. The qualified persons (together, the Authors) who supervised the preparation of this technical report are:

Wayne Chapman, MAusIMM CP (Mining), Technical Manager, Stawell Gold Mines, Victoria, Newmarket Gold. Wayne Chapman has over 11 years of relevant experience and has worked at Stawell Gold Mine for the past 3.5 years.

   

Mark Edwards, MAusIMM CP (Geology), General Manager Exploration, Newmarket Gold. Mark Edwards has over 18 years of relevant experience and regularly has worked at Stawell Gold Mine over the past 4 years.

   

Justine Tracey, BScH Geology, MAIG, MAusIMM CP (Geology), Senior Resource Geologist, Stawell Gold Mines, Victoria, Newmarket Gold. Justine Tracey has over 13 years of relevant experience and has worked at Stawell Gold Mines as a Resource geologist for 6 years.

Responsibilities for the preparation of certain sections of this technical report have been assigned to individual authors as shown in Table 2-1. Technical Reporting Responsibilities of this technical report and such individual authors are not responsible for sections of this technical report other than those indicated in this table.

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Newmarket Gold
Stawell Gold Mines

EPA

Environmental Protection Agency

FAS

Federal Albion South

FX

Foreign exchange

g or gm

Gram (s)

g/t

Grams per tonne

GDA94

Geocentric Datum of Australia (usually referred to as GDA94, or just GDA) is a coordinate system for Australia

GEOCD

Geological Rock code

GWM-Water

Grampians-Wimmera-Mallee Water

ha

Hectare (10,000 m2)

Historical Resource

Non-compliant Mineral Resource as reported in publically available documentation. In no terms is this type of Mineral Resource to be included or quantified but is noted in this technical report to reflect previous work that has been completed on deposits outside the current listing in this Mineral Resource statement

IRR

Internal Rate of Return

JORC Code

Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves prepared by the Joint Ore Reserves Committee of the Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Minerals Council of Australia, as amended

JV

Joint Venture

LLD

Lower Limit of Detection

kg

Kilogram(s)

km

Kilometer(s)

Ma

Million Years

m

Meter (s)

MGA

Map Grid of Australia

ML

Milliliters

Mt

Million tonnes

Mtpa

Million tonnes per annum

MMP

Mine Management Plan

mRL

Mine grid reduced level (meters)

NATA

National Association of Testing Authorities

NMI

Newmarket Gold Inc.

NPV

Net Present Value

Oz

Troy ounces (31.1035 g)

Oz/an

Ounce (gold) per annum

%

Per cent by weight

ppb

Parts Per Billion

ppm

Parts Per Million

QA/QC

“Quality Assurance – Quality Control”

PEA

Preliminary Economic Assessment

Qualified Person

“Qualified Person” has the meaning ascribed to such term in NI43-101

RAB

Rotary Air Blast drillhole

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RC Reverse Circulation Drillhole
RL Reduced Level
ROM Run of Mine ore pad
T or t Metric tonne (2,204 lbs)
TSF Tailings Storage Facility
UG Under Ground
USF1 Upper South Fault 1
$C Canadian Dollar
US$ United States Dollar
Volc Volcanogenic
°C Degrees Celsius

TABLE 2-2 DEFINITIONS AND ABBREVTIATIONS

  2.5

MINERAL RESOURCE AND RESERVE DEFINITIONS


  2.5.1

MINERAL RESOURCES

Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured categories, Figure 2-1. An Inferred Mineral Resource has a lower level of confidence than that applied to an Indicated Mineral Resource. An Indicated Mineral Resource has a higher level of confidence than an Inferred Mineral Resource but has a lower level of confidence than a Measured Mineral Resource.

A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction.

The location, quantity, grade or quality, continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.

Material of economic interest refers to diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals.

The term Mineral Resource covers mineralization and natural material of intrinsic economic interest which has been identified and estimated through exploration and sampling and within which Mineral Reserves may subsequently be defined by the consideration and application of Modifying Factors. The phrase ‘reasonable prospects for eventual economic extraction’ implies a judgment by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction. The Qualified Person should consider and clearly state the basis for determining that the material has reasonable prospects for eventual economic extraction. Assumptions should include estimates of cutoff grade and geological continuity at the selected cut-off, metallurgical recovery, smelter payments, commodity price or product value, mining and processing method and mining, processing and general and administrative costs. The Qualified Person should state if the assessment is based on any direct evidence and testing.

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Interpretation of the word ‘eventual’ in this context may vary depending on the commodity or mineral involved. For example, for some coal, iron, potash deposits and other bulk minerals or commodities, it may be reasonable to envisage ‘eventual economic extraction’ as covering time periods in excess of 50 years. However, for many gold deposits, application of the concept would normally be restricted to perhaps 10 to 15 years, and frequently to much shorter periods of time.

    2.5.1.1

INFERRED MINERAL RESOURCE

An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity.

An Inferred Mineral Resource has a lower level of confidence than that applying to an Indicated Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.

An Inferred Mineral Resource is based on limited information and sampling gathered through appropriate sampling techniques from locations such as outcrops, trenches, pits, workings and drill holes. Inferred Mineral Resources must not be included in the economic analysis, production schedules, or estimated mine life in publicly disclosed Pre-Feasibility or Feasibility Studies, or in the Life of Mine plans and cash flow models of developed mines. Inferred Mineral Resources can only be used in economic studies as provided under NI 43-101.

There may be circumstances, where appropriate sampling, testing, and other measurements are sufficient to demonstrate data integrity, geological and grade/quality continuity of a Measured or Indicated Mineral Resource, however, quality assurance and quality control, or other information may not meet all industry norms for the disclosure of an Indicated or Measured Mineral Resource. Under these circumstances, it may be reasonable for the Qualified Person to report an Inferred Mineral Resource if the Qualified Person has taken steps to verify the information meets the requirements of an Inferred Mineral Resource.

    2.5.1.2

INDICATED MINERAL RESOURCE

An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics are estimated with sufficient confidence to allow the application of Modifying Factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit.

Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing and is sufficient to assume geological and grade or quality continuity between points of observation.

An Indicated Mineral Resource has a lower level of confidence than that applying to a Measured Mineral Resource and may only be converted to a Probable Mineral Reserve.

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Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such as to allow confident interpretation of the geological framework and to reasonably assume the continuity of mineralization. The Qualified Person must recognize the importance of the Indicated Mineral Resource category to the advancement of the feasibility of the project. An Indicated Mineral Resource estimate is of sufficient quality to support a Pre-Feasibility Study which can serve as the basis for major development decisions.

    2.5.1.3

MEASURED MINERAL RESOURCE

A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit.

Geological evidence is derived from detailed and reliable exploration, sampling and testing and is sufficient to confirm geological and grade or quality continuity between points of observation.

A Measured Mineral Resource has a higher level of confidence than that applying to either an Indicated Mineral Resource or an Inferred Mineral Resource. It may be converted to a Proven Mineral Reserve or to a Probable Mineral Reserve.

Mineralization or other natural material of economic interest may be classified as a Measured Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such that the tonnage and grade or quality of the mineralization can be estimated to within close limits and that variation from the estimate would not significantly affect potential economic viability of the deposit. This category requires a high level of confidence in, and understanding of, the geology and controls of the mineral deposit.

    2.5.1.4

MINERAL RESOURCE AND MINERAL RESERVE CLASSIFICATION

The CIM Definition Standards provide for a direct relationship between Indicated Mineral Resources and Probable Mineral Reserves and between Measured Mineral Resources and Proven Mineral Reserves. In other words, the level of geoscientific confidence for Probable Mineral Reserves is the same as that required for the in situ determination of Indicated Mineral Resources and for Proven Mineral Reserves is the same as that required for the in situ determination of Measured Mineral Resources. Figure 2-1, displays the relationship between the Mineral Resource and Mineral Reserve categories.

Figure 2-1 sets out the framework for classifying tonnage and grade/quality estimates so as to reflect different levels of geological confidence and different degrees of technical and economic evaluation. Mineral Resources can be estimated by a Qualified Person, with input from persons in other disciplines, as necessary, on the basis of geoscientific information and reasonable assumptions of technical and economic factors likely to influence the eventual prospect of economic extraction. Mineral Reserves, which are a modified sub-set of the Indicated and Measured Mineral Resources (shown within the dashed outline in Figure 2-1), require consideration of modifying factors affecting profitable extraction, including mining, processing, metallurgical, economic, marketing, legal, environmental, infrastructure, social and governmental factors, and should be estimated with input from a range of disciplines. Additional test work, e.g. metallurgy, mining, environmental is required to reclassify a resource as a reserve.

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In certain situations, Measured Mineral Resources could convert to Probable Mineral Reserves because of uncertainties associated with the modifying factors that are taken into account in the conversion from Mineral Resources to Mineral Reserves. This relationship is shown by the dashed arrow in Figure 2-1 (although the trend of the dashed arrow includes a vertical component, it does not, in this instance, imply a reduction in the level of geological knowledge or confidence). In such a situation these modifying factors should be fully explained. Under no circumstances can Indicated Resources convert directly to Proven Reserves.

In certain situations previously reported Mineral Reserves could revert to Mineral Resources. It is not intended that re-classification from Mineral Reserves to Mineral Resources should be applied as a result of changes expected to be of a short term or temporary nature, or where company management has made a deliberate decision to operate in the short term on a non-economic basis. Examples of such situations might be a commodity price drop expected to be of short duration, mine emergency of a non-permanent nature, transport strike etc.

FIGURE 2-1 RELATIONSHIP BETWEEN MINERAL RESERVES AND MINERAL RESOURCES

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  2.5.2

PRELIMINARY FEASIBILITY STUDY

The CIM Definition Standards requires the completion of a Pre-Feasibility Study as the minimum prerequisite for the conversion of Mineral Resources to Mineral Reserves.

A Pre-Feasibility Study is a comprehensive study of a range of options for the technical and economic viability of a mineral project that has advanced to a stage where a preferred mining method, in the case of underground mining, or the pit configuration, in the case of an open pit, is established and an effective method of mineral processing is determined. It includes a financial analysis based on reasonable assumptions on the Modifying Factors and the evaluation of any other relevant factors which are sufficient for a Qualified Person, acting reasonably, to determine if all or part of the Mineral Resource may be converted to a Mineral Reserve at the time of reporting. A Pre-Feasibility Study is at a lower confidence level than a Feasibility Study.

  2.5.3

FEASIBILITY STUDY

A Feasibility Study is a comprehensive technical and economic study of the selected development option for a mineral project that includes appropriately detailed assessments of applicable Modifying Factors together with any other relevant operational factors and detailed financial analysis that are necessary to demonstrate, at the time of reporting, that extraction is reasonably justified (economically mineable). The results of the study may reasonably serve as the basis for a final decision by a proponent or financial institution to proceed with, or finance, the development of the project. The confidence level of the study will be higher than that of a Pre-Feasibility Study.

The term proponent captures issuers who may finance a project without using traditional financial institutions. In these cases, the technical and economic confidence of the Feasibility Study is equivalent to that required by a financial institution.

  2.5.4

MINERAL RESERVES

Mineral Reserves are sub-divided in order of increasing confidence into Probable Mineral Reserves and Proven Mineral Reserves. A Probable Mineral Reserve has a lower level of confidence than a Proven Mineral Reserve.

A Mineral Reserve is the economically mineable part of a Measured and/or Indicated Mineral Resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at Pre-Feasibility or Feasibility level as appropriate that include application of Modifying Factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

The reference point at which Mineral Reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.

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The public disclosure of a Mineral Reserve must be demonstrated by a Pre-Feasibility Study or Feasibility Study.

Mineral Reserves are those parts of Mineral Resources which, after the application of all mining factors, result in an estimated tonnage and grade which, in the opinion of the Qualified Person(s) making the estimates, is the basis of an economically viable project after taking account of all relevant Modifying Factors. Mineral Reserves are inclusive of diluting material that will be mined in conjunction with the Mineral Reserves and delivered to the treatment plant or equivalent facility. The term ‘Mineral Reserve’ need not necessarily signify that extraction facilities are in place or operative or that all governmental approvals have been received. It does signify that there are reasonable expectations of such approvals.

‘Reference point’ refers to the mining or process point at which the Qualified Person prepares a Mineral Reserve. For example, most metal deposits disclose Mineral Reserves with a “mill feed” reference point. In these cases, reserves are reported as mined ore delivered to the plant and do not include reductions attributed to anticipated plant losses. In contrast, coal reserves have traditionally been reported as tonnes of “clean coal”. In this coal example, reserves are reported as a “saleable product” reference point and include reductions for plant yield (recovery). The Qualified Person must clearly state the ‘reference point’ used in the Mineral Reserve estimate.

Modifying Factors are considerations used to convert Mineral Resources to Mineral Reserves. These include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

    2.5.4.1

PROBABLE MINERAL RESERVE

A Probable Mineral Reserve is the economically mineable part of an Indicated, and in some circumstances, a Measured Mineral Resource. The confidence in the Modifying Factors applying to a Probable Mineral Reserve is lower than that applying to a Proven Mineral Reserve.

The Qualified Person(s) may elect, to convert Measured Mineral Resources to Probable Mineral Reserves if the confidence in the Modifying Factors is lower than that applied to a Proven Mineral Reserve. Probable Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study.

    2.5.4.2

PROVEN MINERAL RESERVE

A Proven Mineral Reserve is the economically mineable part of a Measured Mineral Resource. A Proven Mineral Reserve implies a high degree of confidence in the Modifying Factors.

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Application of the Proven Mineral Reserve category implies that the Qualified Person has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report. The term should be restricted to that part of the deposit where production planning is taking place and for which any variation in the estimate would not significantly affect the potential economic viability of the deposit. Proven Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study. Within the CIM Definition standards the term Proved Mineral Reserve is an equivalent term to a Proven Mineral Reserve.

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  3

RELIANCE ON TECHNICAL EXPERTS – MINERAL RESOURCE ESTIMATES

This technical report has been prepared by the Authors for Newmarket Gold and is based, in part, as specifically set forth below in Table 3-1 on the review, analysis, interpretation and conclusions derived from information which has been provided or made available to the Authors by Newmarket Gold, augmented by direct field examination and discussion with former employees, current employees of Newmarket Gold and consultants who have previously worked for past operators or are currently working for Newmarket Gold. The Authors have reviewed such information and determined it to be adequate for the purposes of this Technical Report. The Authors do not disclaim any responsibility for this information.

It is the view of the Authors that the data collection, storage, and analysis methods utilized in estimating and compiling Mineral Resource estimates at Stawell Gold Mines are of sufficient quality to ensure the information is reliable and suitable for the compilation of this technical report. The Authors are not aware of any critical data that has been omitted so as to be detrimental to the objectives of this technical report. There was sufficient data provided to enable credible interpretations to be made in respect of the data. The principal Authors, Justine Tracey and Wayne Chapman believe that no information that might influence the conclusion of this technical report has been withheld from the study.

Newmarket Gold used the assistance of internal employees to assist with the generation of this technical report. Below is a summary of those roles and the areas they were responsible for within this technical report.

Area of Contribution Site Expert Sections
Metallurgy and Recovery Peter Wemyss
Metallurgy Manager
17
Environmental Studies David Coe
Environmental Manager
3 & 20
Geological Exploration Sarah Heard Senior
Exporation Geologist
9

TABLE 3-1 SITE EXPERTS WHO CONTRIBUTED TO THE TECHNICAL REPORT

The Author of Section 20 of this technical report is also reliant on external consultants for expert advice and opinions. Where external advice has been used in Section 20 the appropriate parties have been referenced.

  3.1

HISTORICAL INFORMATION

Information relating to historical exploration, production and Mineral Resources and Reserves, mining and metallurgy has in part been sourced from summary documentation prepared by past operators and Newmarket Gold, from previously filed NI 43-101 Technical Reports and corporate filings and press releases available on the System for Electronic Document Analysis and Retrieval (SEDAR) website:www.SEDAR.com and from other public sources. Where required the source of this information has been noted in this technical report.

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Interpretations and conclusions contained herein reflect the detail and accuracy of historical exploration data available for review. Given the nature of mineral exploration, and with more detailed modern exploration work and new exploration and mining technology, more precise methods of analysis and advances in understanding of local and regional geology and mineral deposit models over time, the interpretations and conclusions contained herein are likely to change and may be found to be in error or be obsolete. As part of Newmarket Gold’s ongoing process to improve Mineral Resource estimates, all mining information is reconciled against the models to ensure accuracy; this assists in improving the accuracy of the models. A qualified person has not done sufficient work to classifying historical estimates as current Mineral Resource or Mineral Reserves and Newmarket Gold is not treating any historical estimates as current Mineral Resources or Mineral Reserves .

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  4

PROPERTY DESCRIPTION & LOCATION


  4.1

LOCATION

Stawell Gold Mines is located in the Australian State of Victoria, 250km northwest of Melbourne and 2km from the Township of Stawell (see Figure 4-1). Stawell is a rural township of approximately 6,500 people and is within the Northern Grampians Shire.

  4.2

PROPERTY DESCRIPTION

Stawell Gold Mines principal approval is its Mining Lease MIN5260, (see Figure 4-2) issued by the Victorian State Government under the Sustainable Development Act. This MIN5260 lease (centroid coordinates of 142.80° E and 37.06° S, GDA94) encompasses both the Magdala and Wonga Mines and is located both under and around the Township of Stawell with an area of approximately 1,000.58 Ha. The mining lease is comprised of private and crown land including designated crown land reserves. Designated crown land reserves require particular consideration in accordance with the Sustainable Development Act and the National Parks (Box Ironbark and Other Parks) Act 2002.

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FIGURE 4-1 LOCATION MAP OF THE MIN5260 LEASE. THE GRID IS LATITUDE AND LONGITUDE (GDA94)

Stawell Gold Mines refers to the underground workings on Mining Lease MIN5260. Big Hill locally refers to the area approximately 1km north of the current underground mine and milling operations at Stawell Gold Mines. With respect to this technical report, Stawell Gold Mines refers to the underground workings as highlighted in yellow and the Big Hill surface project as highlighted in pink (see Figure 4-2 below).

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FIGURE 4-2 AERIAL VIEW OF THE MIN5260 LEASE INDICATING UNDERGROUND AND PIT LOCATIONS

The surface project site is mainly Crown land under the control of DELWP, with four exceptions, namely:

  CA 11B is a Public Purpose Reserve – Committee of Management is Northern Grampians Shire Council;
  CA 10A & 11A comprise a Reservoir Reserve – Committee of Management is GWM-Water;
  CA 10F is freehold land owned by GWM-Water; and
  Private land lot 11 is owned by Stawell Gold Mines.

Within the Crown land, apart from the mine ventilation shaft, above-ground mine development and some memorials and a picnic facility, four small areas of land are developed for specific uses, namely:

  CA 10B a former Forest Commission Reserve, containing a DELWP Fire Tower.
CA 10C, a former Municipal Purposes Reserve leased to Shire of Stawell is now leased to Vencorp and houses the organization’s radio communication facility.
CA 10D a former Police & Emergency Services Reserve, is leased to the Victorian Police. It houses the State Mobile Radio Network (Telstra).
CA 10G is leased by Optus Communications and contains a mobile telephone tower and three buildings. It houses the mobile telephone facilities for Optus, Telstra and Vodafone.

  4.3

LEGISLATION AND PERMIT

Stawell Gold Mines principal approval, MIN5260, issued by the Victorian State Government under the Mineral Resources (Sustainable Development) Act 1990 (Vic) (the Sustainable Development Act) is the applicable “right to mine” title over the land described in Section 4.1. Newmarket Gold is the 100% owner of the title. MIN5260 is current to 2020, when it will require to be renewed. This approval was first issued on the May 31, 1985 as ML1219 and has been amended on at least six occasions as a result of approved Work Plan variations.

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Attached to this title are mining license conditions upon which a Work Plan and all associated Work Plan variations are filed with the regulatory authority, the license is regularly reviewed and updated by the State according to current legislations and guidelines. Stawell Gold Mines mining license was last updated in 2011. The license covers the areas of:

  Work plan requirements
  Land use
  Insurances
  Public Safety
  Community engagement
  Reporting
  Environmental impact management
  Rehabilitation and bonds
  Exploration
  Hazardous material management

Mining License 5520 was granted to Stawell Gold Mines in March 2010 to cover the northern extension of the Magdala lodes outside MIN5260. No mining activity was undertaken on this license and exploration drilling found the prospect to be uneconomic. The license was relinquished in March 2013 and the area is now covered as Exploration License EL5474.

Apart from the primary mining legislation the Sustainable Development Act, operations on MIN5260 are subject to the additional following legislation and regulations for which all appropriate permits, (MIN5260) and approvals have been obtained.

Acts:

  Extractive Industry Development Act 1995 (Vic)
  Environment Protection Act 1970
  Mines Act 1958
  Planning and Environment Act 1987
  Environmental Protection and Biodiversity Conservation Act 1999
  National Environment Protection Council (Vic) Acts 1995
  Flora and Fauna Guarantee Act 1988
  Catchment and Land Protection Act 1994
  Archaeological and Aboriginal Relics Preservation Act 1972
  Heritage Act 1995
  Forest Act 1958
  Dangerous Goods Act 1985
  Mines Safety and Inspection Act 1994
  Drugs, Poisons and Controlled Substances Act 1981
  Health Act 1958
  Water Act 1989

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  Crown Land (Reserves) Act 1978
  Radiation Act 2005
  Sustainability Victoria Act 2005
  Country Fire Authority Act 1958
  Conservation, Forests and Lands Act 1987
  Wildlife Act 1975

Regulations:

  Dangerous Good (Explosives) Regulations 2011
  Dangerous Good (Storage and Handling) Regulations 2000
  Dangerous Goods (HCDG) Regulations 2005
  Occupational Health and Safety Regulations 2007
  Forest Fire Regulations 1992
  Biodiversity Conservation Regulation 2000
  Drugs, Poisons and Controlled Substances (Commonwealth Standard) Regulations 2001
  Mineral Resources (Infringements) Regulations 1991
  Environmental Protection (Vehicle Emissions) Regulations 2003

Regular reviews of legislation and regulation requirements have been completed and Stawell Gold Mines maintains all required statutory approvals to continue with mining operations.

  4.4

ROYALTY AND ENCUMBERANCES

There are two royalties associated with MIN5260. The first royalty of $2.00 per ounce is payable to Mineral Ventures of Australia (MVA). The royalty agreement was signed in February 2004 and is in place until the earlier of 15 years of production or 2.5 million ounces of gold produced. Furthermore, this royalty agreement extends to Victorian tenements held by Leviathan Resources Ltd (now a wholly-owned subsidiary of Newmarket Gold), which included MIN5260. The second royalty of 1% of Net Smelter Returns and is payable to Alamos Gold Inc. The royalty agreement was signed in January 2015 and is in place from 1st January 2016 and in perpetuity thereafter. Furthermore, this royalty extends to the following tenements: EL 5474, EL 3008 and EL 4279.

  4.5

ENVIRONMENTAL LIABILITIES

Stawell Gold Mines is operating under a Work Plan submitted as required under Section 3 of the general license conditions of MIN5260. A key component of this Work Plan is an Environmental Management Plan (EMP) the most recent of which was approved by the Department of Primary Industries Victoria in February 2013 in conjunction with the 2013 Processing of Mt Micke Stockpile Work Plan Variation. A requirement of the general license conditions of a mining license is to maintain the EMP and update it accordingly as work plan variations are presented.

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Rehabilitation for the project is ongoing: Stawell Gold Mines entered into a cooperative research project with both Curtin and Melbourne Universities in 2000 to conduct rehabilitation trials to prepare the rehabilitation program for eventual closure of the operations. These trials were extensive and have been undertaken over many years with works ongoing with O’Kane Consultancy regarding tailings storage facility capping design. Other than the existing rehabilitation bond of A$4,803,000 lodged with the Department of State Development and Business Innovation, the project is not subject to any other environmental liabilities.

  4.6

STAWELL GOLD MINES LOCAL SURVEY GRID REFERENCE

All survey data on MIN5260 is collected and stored using modified AMG co-ordinates, based on Australian Map Grid AGD 66 (Zone 54). The convention is to drop the first digit of the northing, so 5,896,000N becomes 896,000N. The easting value is unchanged.

The mine RL is calculated as -300m RL AHD and displayed as a negative number below surface. The RL origin is at 303.60 AHD station located adjacent to the mine on Big Hill, Stawell.

The principal local grid in use within the Mine Lease is the Stawell Gold Mines grid (also referred to as the “45 degree grid”) as shown in Figure 4-3. This grid is orientated 45° west of AMG north and has its origin at 5890137.479N and 659498.820E. It is convention to divide the northing by 20 and refer to this as the section northing line, (i.e. northing 6200 becomes the 310 section line). Additional local grids are used as required for presentation of geological information as required.

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FIGURE 4-3 LOCAL MINE GRID USED AT STAWELL GOLD MINES

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  5

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE & PHYSIOGRAPHY


  5.1

ACCESSIBILITY

Stawell Gold Mines is easily accessible from Melbourne via the Western Highway. Access closer to the mine site is provided through a network of sealed bitumen government roads. Roads within the mine site are unsealed and regularly maintained. The main Melbourne to Adelaide rail line passes through the Township of Stawell, which is also serviced by a local sealed airfield.

  5.2

CLIMATE

The Township of Stawell is located within the southern part of the Wimmera where the climate is described as semi-arid, allowing for exploration and mining activities all year round. Stawell weather for the past 20 years has recorded an annual daily average temperature of 19.7°C, a daily high mean maximum of 28.1°C in February and low daily maximum mean of 12°C occurring in July. Mean annual rainfall is 562.1 millimeters with 80.4 days per year on average recorded as having rain.

  5.3

LOCAL RESOURCE

Stawell Gold Mines has been in operation for over 30 years, developing a highly experienced workforce. Many contractors, also having a long association with the mine, are available in the Township of Stawell and surrounding regions. Due to Stawell Gold Mines’ close location to the Township of Stawell many facilities are available. Within the Township area is a police station, hospital, schools and shops. Main electricity and water are also accessible.

  5.4

INFRASTRUCTURE

Stawell Gold Mine’s facilities are extensive and representative of a modern gold mining operation. Surface facilities include the gold processing plant, offices, core shed, laboratory and workshops. Larger infrastructure onsite includes a tailings dam, covering 96 Ha and receiving 100% of gold tailings from the processing plant. Three freshwater dams occur throughout the mine lease. The mine purchases electric power from Origin Energy Australia.

Water supply is from harvested rainfall runoff, mine dewatering, recycling of process water from the tailings facility, and by way of a 1ML/day raw water right entitlement and urban customer access to potable supply from Lake Bellfield located in the Grampians Mountains. The capacity of the site water storages is approximately 690ML. Potable water is preferentially used in the processing operations as it improves gold recovery particularly for the sulphide portion of the orebody.

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  5.5

PHYSIOGRAPHY

The area surrounding Stawell is flat to gently undulating farmland with the Grampians Mountain Range and National Park 20 kilometers to the southwest. Close to the center of Stawell is Big Hill, the Township’s highest point 303.6 meters above mean sea level. Stawell Gold Mines is situated on the southern slope of Big Hill. Parts of the area adjacent to the mine are covered by iron bark forest.

  5.6

MINING PERSONNEL

The Stawell Gold Mines is located adjacent to the Stawell Township and as a result, the mining industry within the region is well understood and supported by the surrounding centers. Mining activities have a direct impact on the manufacturing, service and hospitality sectors of the local economies.

Newmarket Gold is involved with safety, management, mining engineering, geology and exploration, survey, stores control, processing and maintenance, environmental and permitting, and administration, mining and haulage functions.

A number of other contracting groups are engaged for maintenance services, labor hire services, road repairs, drilling activities and other typical contracted activities.

The Stawell Gold Mines has a workforce of 155 people, including 135 employees and 20 contractors. Newmarket Gold preferentially employs locally.

Mining and processing operations run 24 hours a day, each day of the year, primarily based on two 12 hour shifts and working a range of rosters.

  5.7

PROCESSING FACILITIES

The Stawell gold plant was commissioned in 1984, undergoing a number of upgrades over the years. Current mill capacity is 110tph on underground hard rock and up to 150tph on softer oxide ore types. The treatment facility employs gravity, flotation and fine grind and CIL processes. Infrastructure on site includes a crushing system using a primary jaw crusher and then secondary cone crushers within a closed circuit, before ore is milled in a 1.3MW single-stage, rubber lined, ball mill. The ball mill grinds the ore to 80% passing 120um. The milling circuit incorporates 2 Gravity concentrators (1 Knelson, 1 Falcon concentrator), extracting coarse gold from the slurry prior to flotation/regrind treatment. The 3-stage flotation circuit separates sulphide minerals from the gangue material for fine grinding in a stirred mill. Grinding to 12um further liberates gold and increases recovery.

The leach circuit consists of 12 leach/adsorption tanks, with loaded carbon recovered from the CIL circuit eluted in one of two pressure Zadra elution columns to remove gold as an auriferous caustic-cyanide solution from which the gold is recovered by electro winning. The stripped carbon is reactivated in a horizontal kiln and returned to the CIL circuit for reuse.

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Gravity gold recovered is daily discharged into a holding tank, and then intensively leached every 4 to 6 days. Pregnant solution is electrowon onto steel wool.

The electrowon gold (both CIL and Gravity) is pressure cleaned off the stainless steel cathodes then dried and smelted separately in a gas-fired furnace into doré bullion. Bars are stamped for identification and dispatched via security service to the Perth Mint (Western Australia).

Tailings from the processing plant are pumped to Tailings Storage Facility No 2 located approximately 1.8 km SE of the processing plant. This dam is has a currently approved work plan to allow lifting the wall from the current 251.3m RL a further 1.7 meters. At current milling rates this will allow a further 3 years of tailings deposition before further permitting is required.

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  6

HISTORY

Stawell is a historic goldfield that produced 2.7 million ounces of gold between 1853 and 1926 from both alluvial and hard rock sources. There was little mining activity in the Stawell area from 1926 to 11th March 1976 when Western Mining Company Resources Ltd. was granted an exploration license over the Stawell Goldfield.

In 1981, Stawell Gold Mines was reopened by a WMC/Central Norseman Gold joint venture with commencement of the Magdala decline. By 1984, the operation had expanded with the construction of a processing facility and subsequent commencement of an open cut operation at the Wonga Mine (2 kilometers south of Magdala). A number of historical tailing dumps were retreated during this period. Towards the end of mining of the Wonga open cut (1987), the Davis open cut operation was commenced. The Davis open cut exploited the oxide material on up dip projection of the Magdala deposit. The Wonga open cut operated from 1984 to 1987 and produced 778,847 tonnes recovering 69,159 ounces of gold. The Davis open cut operated from 1987 to 1989 and produced 154,525 tonnes for 8,992 recovered ounces of gold.

In December 1992, the operation was acquired in a 50/50 joint venture by MPI and Pittston. At this stage, the Magdala decline was at 410m RL, while the Wonga decline was at 180-200m RL. With the acquisition, there was a clear direction to increase expenditure on resource definition drilling and near mine exploration. The joint venture continued until 2004 during which time there was a record of continued exploration success with discovery of additional mineralized deposits that were subsequently mined.

In February 2004, MPI acquired Pittston’s 50% share of the project. Exploration continued in the Golden Gift area during 2004 with the commencement of the Golden Gift south surface exploration program. In November 2004, a de-merger of the MPI gold business came into effect, and Leviathan Resources Ltd. was floated in December 2004. The resource drilling into the Golden Gift Deposit initially identified seven areas of mineralization offset from each other due to late faulting. Conversion of these areas of mineralization into ore blocks wasn’t universal but was successful in the majority of cases. The further drilling of the fault blocks also identified other mineralized surfaces previously unknown due to the faulted nature of the Golden Gift. From the increased geological understanding of the Golden Gift Deposit, it was clear in the mine planning process that two declines were required, the GG5 and GG3 declines, to access the ore zones for continuity of supply.

In January 2007, Perseverance acquired Leviathan Resources Ltd. Perseverance was acquired by Northgate on February 18, 2008. Northgate was acquired by AuRico in October, 2011. Crocodile Gold completed their acquisition of Stawell Gold Mines from AuRico on May 4, 2012. A merger in July 2015 between Newmarket Gold Inc. and Crocodile Gold created Newmarket Gold the current owner of the Mine. Production of both the Magdala and Golden Gift Orebodies has remained continuous with workings having reached to a depth of -1200mRL and -1600mRL respectively. Lower mine closure was exercised by mid-2013 following the interception of the Wildcat Porphyry and subsequent exploration drilling was conducted only to determine orebody offsets beyond economic viability for the continuation of deep mine operations.

Since mid-2013 the mining and exploration methodology at the Stawell Gold Mines has changed considerably as a consequence of the retreat from mining of lower levels. The ‘lower levels’ primarily covers the Golden Gift orebodies which are located below the South Fault; from -900mRl to the lowest development level at -1646mRL. The closure of the lower levels of the mine and an adjusted cost model has resulted in lower operating costs. This has allowed for the consideration of lower grade material for mining. As a consequence, all undepleted material in the upper levels (above the South Fault) have required a review for mining potential and unrecognized mineralized extensions with this allowing for a continued operating profile of around 38koz pa from both lower grade underground sources and processing low grade surface stocks.

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In 1999 a proposal was submitted to the Victorian State Government to establish an open pit mining operation over the area known as Big Hill. This proposal was subsequently rejected due a number of ministerial concerns and an additional application addressing these concerns was presented in 2014. Ministerial assessment raised additional concerns that Stawell Gold Mines are currently addressing.

  6.1

HISTORICAL AND MODERN PRODUCTION

Stawell is a historic goldfield having produced approximately 2.7Moz of gold between 1853 and 1926 from both alluvial and hard rock sources. Since the commencement of mining in the modern period, 1984, until December 2015, over 2.2 million ounces have been produced from the Stawell orebody. A summary of annual gold production from underground is shown in Figure 6-1. Ore treated and head grades from underground are shown in Figure 6-2.


FIGURE 6-1 ANNUAL GOLD PRODUCTION SINCE 1984

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FIGURE 6-2 ANNUAL PROCESSING PLANT PRODUCTION SINCE 1984

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  7

GEOLOGICAL SETTING AND MINERALIZATION

For additional details on the geological setting of the Stawell Project beyond those required by the scope of this technical report, the reader is referred to Fredericksen, D., Miller, G., Dincer, T. (2008).

  7.1

REGIONAL GEOLOGY

The Stawell Goldfield is located in the western Stawell Zone of the Lachlan Fold Belt (Figure 7-1). The Stawell Zone is a belt of predominantly deformed meta-sedimentary rocks representing the lower parts of the Cambro-Ordovician Lachlan Fold Belt stratigraphy bound to the west by the Moyston Fault and to the east by the Coongee Break (Vandenberg et al. 2000).

Interpretations from the Victorian Geological Survey present a thin skinned tectonics model where the Moyston Fault is an east dipping basal detachment which has juxtaposed higher metamorphic grade rocks of the Stawell Zone against lower grade Cambrian rocks of the Delamarian Glenelg Zone. The west dipping Stawell Fault, Coongee Break and other parallel west dipping faults represent back thrusts from the Moyston Fault. These back thrusts have progressively emplaced deeper stratigraphy against shallower stratigraphy with a generally west over east sense. An apparent anomaly in this sequence is the presence of deeper magnetic stratigraphy in the Stawell-Wildwood corridor. Vandenberg et al. 2002 interprets that the Pleasant Creek Fault, to the west of the Stawell Fault, actually dips east and has an east over west sense - similar to the Moyston Fault. The Stawell-Wildwood corridor therefore represents a significant structural high in an up-thrown block of deeper stratigraphy between the Coongee Break and Pleasant Creek Fault.

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FIGURE 7-1 IMAGE SHOWING LACHLAN FOLD BELT, LOCATING STAWELL ON THE WESTERN BOUNDARY

  7.2

LOCAL GEOLOGY AND PROPERTY

There are three separate orebodies defined at Stawell; the Magdala (west flank and east flank), Golden Gift and Wonga. All have differing characteristics but the same local geology is relevant to the genesis of them all.

  7.2.1

STRATIGRAPHY AT STAWELL GOLD MINES

The stratigraphy at Stawell is divided into three principal units: Magdala Basalt; Albion Formation; Leviathan Formation, see Figure 7-4 (Squire and Wilson, 2005). Intruded into this sequence are the Stawell Granite and a number of felsic and mafic intrusions. Squire and Wilson (2005) interpret that the rock unit previously termed the Magdala Volcanogenics (Watchorn and Wilson, 1989) is an alteration facies that locally occurs adjacent to the Magdala Basalt.

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    7.2.1.1

MAGDALA BASALT

The Cambrian aged Magdala Basalt is composed of subaqueous low-K tholeiitic lavas that exhibit an aphyric to sparsely plagioclase-phyric texture (Watchorn and Wilson, 1989; Squire and Wilson, 2005). The Basalt body comprises flows ranging from 0.5m to 50m thick, pillows basalts with pillows ranging in size from 0.1 to 2m in size (Watchorn and Wilson, 1989) and monomictic basalt breccias of varying proportions (Pritchard, 2001; Squire and Wilson, 2005).

The basalts are interpreted to form part of the Victorian Cambrian greenstone sequences, which are the oldest known rocks in the Palaeozoic Lachlan Orogen and have inferred ages of 516-514Ma (Squire and Wilson, 2005). The basalts near Stawell occur as dome-like units in the footwall and hangingwall of major faults (Miller and Wilson, 2002). The Magdala Basalt, which closely resemble typical back-arc basin basalts (Kaufman, 2003; Crawford, 1988) has been interpreted to represent the medial to distal facies on the flank of a large basalt edifice upward of 500m thick (Squire and Wilson, 2005) and has similar magmatic affinities to the known basalt bodies north of Stawell, Wildwood and Kewell Basalts (Kaufman, 2003; Jupp, 2003).

    7.2.1.2

LEVIATHAN FORMATION

Overlying the Magdala Basalt is a 200-300m thick sequence of non-fossiliferous turbidites (Squire and Wilson, 2005). The turbidite sequence has been subdivided into two different lithologies: the Albion Formation and the Leviathan Formation. The differences between the two lithologies was first recognized but not explored by Gane (1998). He recognized the sediments on the western side of the Magdala Basalt graded from predominantly mud-rich to more sand-rich away from the basalt.

The Albion Formation is believed to have been deposited penecontemporaneously with the Magdala Basalt, and is found on the east flank, interbedded with the Magdala basalt and on the west flank (Squire and Wilson, 2005), it is the lowest clastic sequence found proximal to the Magdala Basalt. The unit varies in thickness with the top of the unit defined by a 20-100m sequence of black mudstone. Within the Albion Formation there are a number of facies which along with black mudstone include calcareous sandstone, siliceous siltstone and sulphidic black mudstone. Squire and Wilson (2005) suggested that the sediments were deposited predominantly due to suspension settling in a sediment-starved sedimentary basin. There were short-lived periods of oxygen-rich conditions shortly after volcanism recognized by the presence of siliceous siltstone but the dominance of black mudstone within the Albion formation indicates the basin of deposition was predominantly anoxic (Squire and Wilson, 2005). The provenance for the Albion Formation sediments has been identified from detrital compositions to be a low-grade metamorphic terrain (Cas, 1983).

The Leviathan Formation overlies the Albion Formation and is dominated by fine-to medium-grained quartz-rich sandstones (Squire and Wilson, 2005). The contact between the two formations is gradational and conformable. Although the Leviathan Formation was deposited in a higher-energy environment than the underlying Albion Formation, the detrital compositions indicate little change in the provenance between the two formations (Squire and Wilson, 2005).

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The Leviathan and Albion Formations are not segregated by the mine or exploration geologists at Stawell Gold Mines and are referred to by the local name of ‘Mine Schist’.

    7.2.1.3

MAGDALA FACIES

The Magdala Facies, termed Magdala Volcanogenics at Stawell Gold Mines, distinguished by its dark green color, is a result of intense chloritic±stilpnomelane alteration of mudstone and or shales located at the base of the Albion Formation and immediately above the Magdala Basalt. The Magdala Facies are the primary and most important host rock for the sulphide replacement style of gold mineralization at Stawell. Major mineralization sulphides include arsenopyrite, pyrite and pyrrhotite, the latter two commonly occurring along cleavage planes and concentrated within shear zones (Robinson, 2005). Magdala Facies is also occurs on the east flank, but differs in that BIF can clearly be identified as protolith and could provide the iron source for chloritic±stilpnomelane alteration. Sulphide replacement of magnetite bands is identified as bands of pyrrhotite.

    7.2.1.4

FELSIC INTRUSIONS

Quartz±feldspar-phyric felsic intrusions crosscut the turbidite sequence. The quartz±feldspar-phyric felsic intrusions vary in thickness from 50cm to 12m wide and showed chilled margins. They are predominantly composed of quartz and plagioclase with phenocrysts up to 3mm in size. The feldspar phenocrysts have euhedral shapes and display multiple twinning while the quartz phenocrysts had a cloud-like appearance and were rimmed with fibrous quartz (Gedge, 1997). The groundmass is composed of ~80% quartz in anhedral grains and display undulose extinction (Gedge, 1997).

The felsic intrusions tend to follow northwest-trending shear zones (Wilson et al., 1992) and the emplacement of the quartz±feldspar-phyric felsic intrusions post-dates the main Magdala mineralization event. The intrusions have been dated at 413±3Ma (Arne et al., 1998).

    7.2.1.5

STAWELL GRANITE

The Stawell Granite was emplaced during the early Devonian, 401±4Ma (Arne et al., 1998), and is located about 2km south of the Magdala Deposit (Xu et al., 1994) and adjacent to the Wonga Deposit. The pluton is approximately 20km wide and 13km long and intrudes the sandstone and shale units of the turbidite sequence. The pluton is an asymmetrically zoned, medium grained intrusion which contains diorites, granodiorites and magnetite-rich felsic granites (Wilson et al., 1992). There is a 0.5km to 1.0km contact aureole surrounding the Stawell Granite (Xu et al., 1994).

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There was a change in the regional stress field within the Lachlan Orogen during the Late Silurian, which is expressed at Stawell as a change from east-west shortening to sinistral wrenching along pre-existing faults (Miller and Wilson, 2004a). Reactivation of the D4 shears by sinistral wrenching is termed D5 (Mapani and Wilson, 1994). The sinistral wrenching was followed by another change in the regional stress field with the shortening direction changing to northwest-southeast. A set of major faults oblique to the earlier structural trends associated with this change in shortening direction are termed ‘early South Fault’ structures (Miller and Wilson, 2004a). The last major deformation event was associated with a final change to a northeast-southwest shortening. Faults associated with this event dip northwest and have a dip-slip sense of movement (Miller and Wilson, 2004a).

FIGURE 7-2 D1 TO D5 DUCTILE AND BRITTLE EVOLUTION OF THE STAWELL SYSTEM. FROM MILLER ET AL. 2006

Figure 7-2 above shows stereonets represent hangingwall transport direction calculated at pole to fault with the circle center of each arrow representing a single fault pole (Miller & Wilson 2004a). These transport directions are the inferred maximum resolved shear stress along a fault for an applied stress tensor. A change in the hangingwall transport direction for similarly oriented faults represents a change in stress tensor.

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FIGURE 7-3 EVOLUTION OF THE STAWELL SYSTEM FROM 420 TO 380 MA (MODIFIED FROM MILLER & WILSON 2004A)

Figure 7-3 above shows stereonets representing hangingwall transport direction calculated at pole to fault (Miller & Wilson 2004a). Map symbol are the same as those in Figure 7-5. From Miller et al. 2006.

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  7.2.3

STAWELL MINE GEOLOGICAL ARCHITECTURE

The dominant feature at Stawell is the 1.2km wide doubly plunging northwest striking Magdala Basalt dome. The Magdala Basalt is made up of a series of basalt noses, interpreted to be flow sheets (Squire and Wilson, 2005), which dip to the southwest and plunge to the northwest. Areas of sedimentation are present between the basalt noses and are locally termed ‘waterloos’. The Magdala Basalt has been drilled and identified to a depth of 1.7km and interpreted from existing drill information and from geophysical modelling to extend along strike at least 5km.

Surrounding the Magdala Basalt dome is the turbidite sequences of the Albion and Leviathan Formations (Mine Schist) which young to the west. The contact between the Mine Schist and Magdala Basalt on the western side is marked by the alteration package of the Magdala Volcanogenics. The Magdala Volcanogenics is weakly developed on the eastern surface of the Magdala Basalt.

This Magdala geology has been faulted and offset by later brittle deformation, the most notable of these offsets is the South Fault which has a northeast over southwest sense of transport (see Figure 7-4, Figure 7-5 and Figure 7-6).

Above the South Fault is the Magdala orebody which contains limited offsets due to late faulting. The Basalt surface in the Magdala orebody which contains limited offsets due to late faulting. The Basalt surface in the Magdala orebody dips to the west and strikes towards 340˚. Beneath the South Fault is the Golden Gift orebody, which is heavily offset by late faulting creating isolated ore blocks. Unlike the Magdala orebody the basalt in the Golden Gift dips to the east and strikes towards 315˚. The late faulting as well as creating isolated ore blocks also complicates the ore geometry within the each block.

To the south of the Magdala Basalt is the Stawell Granite, which structurally is situated below the South Fault (see Figure 7-6).

Located close to an embayment in the Stawell Granite are a series of brittle structures. One of the structures (the Hangingwall structure) strikes towards 350˚ and dips between 25˚ and 50˚ towards the east, and the other structural set, (the Link structures), generally trend toward 240˚ and dip between 40˚ and 70˚ to the southeast (Xu et al., 1994). Crosscutting these late brittle structures are a series of late felsic intrusive. This fault system hosts the Wonga orebody.

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FIGURE 7-4 MINE GEOLOGY CROSS- SECTION HIGHLIGHTING ARCHITECTURE OF THE MAGDALA AND GOLDEN GIFT OREBODIES

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FIGURE 7-5 PLAN VIEW GEOLOGICAL INTERPRETATION OF THE STAWELL STRUCTURAL AND STRATIGRAPHIC ARCHITECTURE AT 1000MRL

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FIGURE 7-6 STAWELL GOLD MINES WEST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF THE MINERALIZED ORE BLOCKS

NOTE : THE GEOLOGICAL AND S PATIAL RELATIONSHIP BETWEEN THE MAGDALA, GOLDEN GIFT AND WONGA DEPOSITS CAN CLEARLY BE SEEN

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FIGURE 7-7 STAWELL GOLD MINES EAST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF THE MINERALIZED ORE BLOCKS

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  7.3

MINERALIZATION

There are three mineralization styles at Stawell, being Magdala (separated into west and east flanks), Golden Gift and Wonga. The Magdala and Golden Gift ore types are hosted within the Magdala Volcanogenic. Within the Magdala Deposit there are three main ore types; Central Lode, Basalt Contact Lodes, and Magdala Stockwork Lodes. The east flank mineralization introduces a new ore type: the Hampshire Lode.

Central Lode mineralization (see Figure 7-8) was a significant production source from Magdala early in the mine’s history. It is a quartz-rich shear lode ranging from 0.5m to 10m in width and generally dips 55 - 65° to the west with a total strike length of 4km and a down dip extend of one kilometer. The overall structure is mineralized economic shoots that vary from 20m to 30m in strike up to 200 –; 350 meters in strike. Free gold in the quartz is associated with pyrite, arsenopyrite and recrystallized pyrrhotite. Average mined grade for Central Lode is 4.0 –; 7.0 g/t Au.

FIGURE 7-8 EXAMPLE OF CENTRAL LODE MINERALIZATION

Basalt Contact Lodes (see Figure 7-9) are located parallel to the Magdala Basalt and in ‘waterloo’ or re-entrant positions. They are typically 2m wide and are represented by arrays of quartz sulphide tension veins immediately adjacent to the Volcanogenic Basalt contacts. Sulphides include pyrrhotite, arsenopyrite and pyrite and occur as alteration selvages on tension vein margins. The main alteration mineral is stilpnomelane, resulting in its dark color. The mineralization is isolated to the Magdala Volcanogenic package with none present in the adjacent Magdala Basalt. Ore shoot lengths range between 50m and 450m. The average mined grade for Basalt Contact Lodes is 4.0 – 9.0 g/t Au.

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FIGURE 7-9 EXAMPLE OF BASALT CONTACT MINERALIZATION

The Magdala Stockwork Lodes are situated above major basalt noses and can be described as a hybrid between Central and Basalt Contact Lodes. They consist of large quartz tension vein arrays with arsenopyrite and pyrrhotite dominant sulphide mineralization. The strike extent is limited to 40m to 50m and limited vertically to between 30m and 50m. Average mined grade for Magdala Stockwork Lodes is 4.0 – 7.0 g/t Au.

Unlike the Magdala Deposit there is only one identifiable ore type in the Golden Gift and this is termed the Golden Gift Stockworks. Though there is only one discernible ore type in the Golden Gift, the Golden Gift Stockworks contain a spectrum of all Magdala styles. Typical widths range from 8-12m up to 30m and the strike extents of shoots range between 150m and 400m. Areas of highest gold grades and largest widths are situated above major basalt noses which are present in most orebodies. Quartz content is generally below 25%. Mineralization includes abundant recrystallized pyrrhotite and coarse grained arsenopyrite, pyrite and visible gold. Average mined grade is 4.0 – 10.0 g/t Au.

The Hampshire Lode is a sub-vertical lode hosted within the Stawell Facies on the east flank of the Magdala basalt. The protolith is predominately BIF which has been variably altered with chlorite ± stilpnomelane. This differs to the west flank where the alteration is so intense, protolith cannot be clearly identified. Where the BIF is sulphized tends to correlate with gold mineralization. The sulphides (pyrrhotite, arsenopyrite, pyrite) are the same as those associated with gold mineralization in the basalt contact lodes of the west flank. Visible gold is often seen in quartz veins, chlorite veins or within the strongly chloritized groundmass. The Hampshire Lode does not sit on the immediate contact of the east basalt flank, but in sediment about 10m to the northeast and appears to have an average width of 5m. There has been no mining of this ore type to date.

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The Wonga Deposit is hosted within the locally termed Wonga Schist that is part of the Leviathan Formation along two main fault systems. The Wonga Schist has undergone contact metamorphism during the emplacement of the Stawell Granite and undergone three ductile deformation events similar to other areas of the Stawell region. The two fault systems controlling the mineralization are the hangingwall structure which, strikes towards 350° and dips between 25° and 50° towards the east, and the Link structures, which generally trend toward 240° and dip between 40° and 70° to the southeast. The mineralization is represented by arsenopyrite disseminations to quartz veins within these structures. The main ore minerals present are anhedral fine grained pyrrhotite and arsenopyrite. The higher grade ore zones often show andalusite sericite alteration with rutile and ilmenite associations. Production grades from 4.0 – 6.0 g/t Au were common for the Wonga ore.

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  8

DEPOSIT TYPES

For details on the deposit types and mineralization of the Stawell deposits the reader is referred to the Technical Report on Stawell Gold Mine (Fredericksen, Miller and Dincer 2008).

  8.1

DEPOSIT TYPES AND MINERALIZATION

Victorian mineralization episodes have been dated to occur during the Devonian and Silurian periods with no gold mineralization occurring prior to 440Ma (Miller and Wilson, 2002). A description of the mineralization episodes in Western Victoria is described below.

The largest and most significant mineralization event in the western Lachlan Orogen occurred at ca. 440Ma (Foster et al., 1998). It occurred contemporaneously in the Stawell and in the Bendigo-Ballarat Zones with the mineralization occurring during late D4 deformation in the Stawell Zone and occurring during late D1 deformation in the Bendigo-Ballarat Zone. The mineralization is hosted in D4 brittle structures associated with east-over-west movement at Stawell while in the Bendigo-Ballarat Zone the mineralization occurs in saddle reefs in the hinges of D1 folds and in reverse faults created via D1 fold lock-up (Miller and Wilson, 2002; Schaubs and Wilson, 2002). This mineralization event produced the largest endowments of gold within the western Lachlan Orogen (Miller and Wilson, 2002).

The next episode of gold mineralization occurred at about 426-420Ma (Foster et al., 1998) and is associated with fault reactivation throughout western Victoria (Miller and Wilson, 2002). This episode of gold mineralization produced significantly smaller endowments than the 440 Ma event (Miller and Wilson, 2002). The late Silurian mineralization is associated with the D5 sinistral wrenching at Stawell and has been recognized at the Percydale fields in the Stawell Zone and at Tarnagulla in the Bendigo-Ballarat Zone (Miller and Wilson, 2002).

The final episode of mineralization recognized in western Victoria is the Wonga mineralization at Stawell (Miller and Wilson, 2004a). The mineralization at Wonga overprints the quartz- and felsic-rich intrusions and is overprinted by the Stawell Granite contact areole. Watchorn and Wilson (1989) suggested that this mineralization is temporally and spatially related to the granites emplacement. Miller and Wilson (2004b) advocate the mineralization event at Wonga formed at ca. 400Ma (Foster et al., 1998). The Wonga mineralization occurred during a late-stage magmatic event within a long-lived orogenic system at shallow crustal levels (Miller and Wilson, 2004b). This mineralization occurred in a series of brittle structures dependent on pre-existing weakness which are related to a fluid over-pressure event after the lockup of major structures (Miller et al., 2004).

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  8.2

ORE TYPES BY AREA


  8.2.1

FEDERAL ALBION AND FEDERAL ALBION SOUTH

The Federal Albion (Federal Albion and Federal Albion South) shares the same geological characteristics as the other Magdala orebodies above the South Fault: a west-dipping to sub-vertical ore lode adjacent to a basalt footwall contact; a Mine Schist hanging-wall; the westerly dipping, quartz-rich Central Lode shear structure; and a weakly Mineralized Volcanogenic package between Central Lode and the Footwall Ore Zone. A cross section showing the relationship between the various geological components is shown in Figure 8-1.

Central Lode is typified by wide (1-10m) laminated to buck quartz seam and associated FW stocks in the Magdala “Volcanogenic” and its hanging-wall is commonly defined by the Magdala “Mine Schist”. The Central Lode is often partially stoped out by later stage felsic porphyry dykes running sub parallel to the Central Lode shear, most dominantly on the hangingwall. Grade in the FAS Central lode is of a low but consistent tenor between 2-3 g/t Au with minor high-grade hits and outliers.

FIGURE 8-1 SCHEMATIC CROSS SECTION SHOWING LOCATION OF FEDERAL ALBION SOUTH AREA

The Extended Lode, on the western side of the Extended Basalt, is typified as a 1.0m to 2m wide, basalt contact mineralization, hosted within strongly chlorite altered volcanogenic or iron-rich sediments. Sulphide 54mineralization consists of coarse arsenopyrite up to 20mm with recrystallized pyrrhotite and minor pyrite.

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Quartz tension veins up to 0.75m wide are a common feature in the basalt contact zone. Another feature common in the immediate contact zone with the basalt is abundant silicification of the host rock in bands up to 1.0m wide. In places Central Lode has sheared out part or all of the Extended Lode and forms the hanging wall.

Figure 8-2 shows a typical cross section through the Federal Albion South area with current interpreted geology. This geological framework is the same for the Federal Albion area, which is the same as Magdala mineralization to the north.

FIGURE 8-2 TYPICAL CROSS-SECTION THROUGH FEDERAL ALBION AREA SHOWING CURRENT INTERPRETED GEOLOGICAL SETTING

  8.2.2

BELOW SCOTCHMAN 250

The Below Scotchmans 250 resource model includes the northern region of the upper levels in the underground mine environment. The Below Scotchmans 250 area is in the (Magdala) basalt flank in a structurally complex area where there is horizontal offset of the basalt and the formation of flat lodes over the basalt noses, Figure 8-3. The Central lode is bound to the top by the Scotchmans fault, and a narrow hangingwall lode is defined.

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The Central Lode shear zone generally has a well-defined hanging wall to shear along the Mine Schist/ Volcanogenic contact but may also occur within volcanogenic package. The footwall structure is generally marked by shear surface also, but at times gradational into stockworked volcanogenics. The distribution of grade with the Central Lode is not uniform.

The Hangingwall lode is geologically similar to the Central lode, with a high component of quartz, however, it is exclusively located in the hangingwall of the Central lode, bound by Mine Schist or pelite and is typically much narrower in the order of 2m -3m. In the model area the hangingwall lode is interpreted below the Flat 3 and above Flat 1, although does not appear to be present above Flat 3.

The Magdala Lode mineralization comprises dark green chloritic volcanogenics, host to quartz sulphide tension veining. The footwall is defined by the contact with Magdala Basalt, and the hangingwall defined by the reduction of chlorite and sulphide. The basalt contact Magdala lode wraps around a number of basalt noses and waterloos and has variable grade.

The Flat lodes 1 and 2 were previously mined by a hand held mining process and are effectively almost depleted. The lodes are typified by massive quartz reef with internal pyrite, arsenopyrite and pyrrhotite. The flats and are almost flat lying (25 -35°), most likely to be a flat fault formed above the basalt noses, that do not extend into the basalt.

Flat 3 is the highest flat defined by the 2015 drilling and appears to tear and offset the Central Lode. Higher gold grade is concentrated at the intersection of the Flat and the Central lode.

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FIGURE 8-3 GENERAL CROSS SECTION OF THE BELOW SCOTCHMANS AREA

  8.2.3

UPPER SOUTH FAULT 2

Geologically, USF2 shares the same characteristics as the other Magdala orebodies, comprising of three distinct mineralization types. The mineralization types, which occur in the Magdala System include the quartz rich shear of Central Lode, the chlorite and sulphide rich Basalt Contact Lodes and the Stockwork Lodes, Figure 8-4.

The Central Lode shear zone generally has a well-defined hangingwall to shear along the Mine Schist/ Volcanogenic contact. The footwall structure is generally marked by shear surface also, but at times gradational into stockwork volcanogenics. The distribution of grade with the Central Lode is not uniform. But in this model has some relationship with the internal basalt contacts. Up to two basalt dykes have intruded the Central Lode, one along or very near the Mine Schist contact, the other in the middle of the Central Lode.

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Grade in the U2 Central lode is of a low but consistent tenor between 2-3 g/t Au with minor high-grade hits and outliers.

The Dukes Basalt Lode mineralization comprises dark green chloritic volcanogenics, host to quartz sulphide tension veining. The hangingwall to the Basalt Contact Lode structure is generally defined by a quartz pyrrhotite shear structure, while the footwall is defined by the contact with the Basalt. The basalt contact is often bounded by a 30-50cm wide quartz sulphide vein containing coarse arsenopyrite and recrystallised pyrrhotite correlating with gold content.

The Stockworks is made up of the Footwall Volcanogenics between the Central Lode and the Dukes Lode and is made up of quartz-sulphide tension vein arrays and contains 10+% massive pyrite and pyrrhotite with trace arsenopyrite. This zone is constrained to the Southern end of the model and has been mined over the 2015 period.

FIGURE 8-4 GENERAL CROSS SECTION OF THE USF2 AREA

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  8.2.4

MAGDALA S6000

Magdala S6000 (North Mag) includes the northern region of the upper levels in the underground mine environment. The North Mag area is a part of the Magdala orebody, with geology dipping to the west. The North Mag area contains the Dukes and Extended Lodes basalt contact and Central Lode as well as the internal Stockworks Lode.

Ore zone geology is the same framework as described in Section 8.3.1 for the Federal Albion area, which is part of the same Magdala mineralization to the south.

  8.2.5

MARINERS

The underground Mariners area represents the down-dip extension/fault off-set of the surface Mariners target which was modelled as part of the Big Hill and Upper Levels model in April 2014.

Geologically, the Mariners area is a fault zone formed by the diverging Scotchmans Fault and Scotchmans Fault Splay which form a 'wedge'. This fault zone is known as the Scotchmans Fault Zone (SFZ) and contains highly deformed slices of basalt and pelites with high carbon alteration. The bounding faults have a high graphite content, which has enabled both faults to capture remobilized gold. Within the SFZ are secondary faults dipping moderately to the west, terminated by the bounding faults, these are considered to be the main structural hosts for the gold mineralization.

Mineralization is hosted within secondary shears and is associated with quartz breccia and fine-medium grained sulphides. The Curiosity, Opportunity, New, Mariners and Spirit Lodes are bound between the Scotchmans Fault at depth and the Scotchmans Splay (see Figure 8-5). Once considered an offset, upper component of the Central Lode system, the Mariners is now considered its own mineralization system within the Scotchmans fault zone. The schist within the zone have a much higher carbon content and are more intensely deformed than Albion Formation typically seen in the footwall of the Central Lode.

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FIGURE 8-5 TYPICAL CROSS-SECTIONS THROUGH MARINERS AREA SHOWING CURRENT INTERPRETED GEOLOGICAL SETTING

Figure 8-5 shows a typical cross section through the Mariners area with current interpreted geology.

  8.2.6

BIG HILL AND UPPER LEVELS

Immediately below the Big Hill surface Mineral Resource is the Upper Levels underground Mineral Resource and both are the up dip extension of the Magdala System. It contains Basalt Contact mineralization, Central Lode mineralization and Stockwork mineralization, all typically seen in the Magdala System and previously mined from an underground perspective.

Big Hill can be broken into 4 different geological mineralization domains, Mariners, Allens, Iron Duke and the Magdala Flank. All except Mariners and Allens are separated by faults. The Allens and Iron Duke are stockwork zones, which in their entirety are covered by the Big Hill Surface Mineral Resource. Mariners and the Magdala Flanks mineralization comprise the Upper levels underground Mineral Resource.

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Mariners Lode has a consistent width and orientation, averaging around 14m thick and dipping to the northwest (mine grid). It constitutes a shear package, with massive quartz veining. Offset extensions of the Mariners shear have also been modelled down dip (see Figure 8-6). The original Upper levels model (1998) identified one offset extension; however, two further offset domains were identified in the 2012 model update. Underground drilling of Mariners Lode in early 2012, developed an increased understanding of the structural framework, identifying increased faulting complexity and offset lodes, which now connect the underground Mariners and Upper levels Mineral Resource models with the Big Hill Surface model.

The main Mariners Lode is offset by Fault 3, which strikes at 145° (mine grid) and dipping at around 35-40º to the northeast. It has an apparent reverse displacement of between 5m and 25m, decreasing northwards. This section of the Lode (Mariners L1) is then truncated and offset at depth by the Cross Course Fault. In this model update, the Cross Course Fault has been remodelled as two fault surfaces, which bounds another lower lode offset (Mariners L2). A third, less horizontally extensive offset (Mariner L3) was identified and was modelled to be bounded by the Cross Course Fault lower surface and the Scotchmans Fault. The Mariners L3 has a greater offset of 15m-25m and plunges further to the north than the above two offset lodes, which is likely due to a greater influence from the Scotchman’s faulting. In reality these offset sections are likely to be more complexly faulted than modeled.

FIGURE 8-6 SECTION THROUGH MARINERS AND ALLENS LODES STRUCTURAL COMPLEXITY AND OFFSETTING

The Allens stockwork zone is located below the Mariners Lode Figure 8-6. Exposures in the Allens open pit show lithological layering (S0 and later S2/3 foliation), as well as massive veining to be very steeply dipping to vertical. Diamond core shows the same. Two massive veins exposed in the Allens Adit strike approximately N-S mine grid and dip vertically. Some stockwork style veining is seen near the top of the zone (immediately beneath the Mariners structure), but much of the geology consists of well bedded, foliated volcanogenic sediments which are unpredictably mineralized. Much of the gold occurs in these sediments. Because of the highly oxidized state of these rocks the form of sulphide mineralization is largely obliterated.

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The relationship between the Mariners and Allens zones is problematic. Little overlap exists beneath the main flatter section of Mariners and the Allens zone, which dies out rapidly northwards beneath Mariners. No clear evidence is seen of the timing relationships between these structures, the Allens mineralization appears to hang as a pendant of dilational veining beneath a change in orientation of the Mariners shear structure.

FIGURE 8-7 SECTION 5840 THROUGH SOUGH PIT SHOWING IRON DUKE AND MAGDALA LODES

The Iron Duke zone Figure 8-7 is a wedge of geology occurring between the Scotsman’s Fault and the Lower Cross Course Fault. It is the up-dip extension of the volcanogenic package/shear zone which has been offset from the main Magdala shear system by reverse movement on the Lower Cross Course Fault. It is truncated above by the Scotchman’s Fault Zone, which again displaces the orebody westward. Interpretation of the geology in this zone is hampered by the strongly oxidized nature of the rocks. It has been interpreted as a series of constrained stockwork style ore zones (same as Allens), where the envelopes have been used to constrain a tonnage. In reality the margins will have gradational boundaries.

For wireframing purposes the Iron Dukes domain has been considered in the same context as Allens and was treated as broad envelope around the stockwork zones constraining bulk tonnage. The previous model wireframed the structures at a higher gold grade which effectively reported a higher grade for lower tonnes.

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A lower domain gold threshold was used for this model update (0.35 g/t Au); in line with the model economic cut off (0.44 g/t Au) to ensure all minable tonnes were economically represented for mining analysis.

“Magdala Flanks” is the term used to describe the package of volcanogenic rocks lying to the west of and continuing up dip from the nose of, the Basalt Antiform. The western margin of this zone is the contact between the Volcanogenics and the Mine Schist, which often plays host to the Hangingwall Shear. The eastern margin is generally the contact with the basalts, but above the basalt noses, the eastern margin is a transitional boundary to siliceous eastern schists, usually marked by a clear grade boundary. The Hangingwall Shear was most recently air-leg mined from the – and –145 levels (and from sub-levels up to –109), between 269 and 278 N sections in the early 2000’s. Significant old stoping on this shear was encountered up dip of those recent air-leg workings. The geological interpretation of the Magdala Flanks mineralization is shown in Figure 8-8.

FIGURE 8-8 SECTION THROUGH THE MAGDALA MINERALIZATION SHOWING THE INTERPRETED GEOLOGY BELOW

  8.2.1

AURORA B

The Aurora B represents the first Mineral Resource on the east flank of the Magdala Basalt, 500m from the west flank where mining currently takes place. Gold mineralization is hosted in the Hampshire Lode about 10m off the contact of the East Basalt nose (Figure 8-9) within the sediments of the lower Albion Fm. The 63Hampshire Lode averages about 5m thick. Grade varies; averaging 3-4 g/t Au with several intercepts greater than 25 g/t Au. Stronger grades appear to be within shallowly north plunging shoots such as those seen on the west flank. Currently, Mineral Resource has only been defined in the Hampshire Lode, but there is the potential for both waterloo mineralization and Wonga Type mineralization within the Aurora B area.

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FIGURE 8-9 TYPICAL CROSS-SECTION THROUGH AURORA B SHOWING CURRENT INTERPRETED GEOLOGICAL SETTING

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  9

EXPLORATION

Exploration during 2015 has focused on extensions of both the Magdala and Wonga Orebodies. Significant exploration progress and successes have been made over the life of the Stawell project including the discovery of the current Mineral Resources and Mineral Reserves. In the period from January 1, 2015 to December 31, 2015 exploration has taken place on the west flank of Magdala, the east flank of Magdala and on surface at the Brummigans Prospect to the immediate east of the Wonga open pit.

Exploration on the Magdala west flank (Figure 9-1) has concentrated on areas close to current infrastructure and includes the Federal Albion South, 250L Below Scotchmans and Upper South Fault 2 areas, which have resulted in conversion into Mineral Resource and reserve. Other areas include the 486L Mid-North Magdala, Golden Gift Offset and Upper South Fault 5 programs.

Exploration on the east flank (Figure 9-2) resulted in the discovery of the Aurora B mineralization. Drilling also occurred on the north extension of Aurora A.

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FIGURE 9-1 MINE GEOLOGY LONGITUDINAL SECTION OUTLINING NEAR MINE EXPLORATION ON WEST FLANK FROM JANUARY 2015 TO DECEMBER 2015

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FIGURE 9-2 MINE GEOLOGY LONGITUDINAL SECTION OUTLINING NEAR MINE EXPLORATION ON EAST FLANK FROM JANUARY 2015 TO DECEMBER 2015

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FIGURE 9-3 PROJECT LOCATION PLAN

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  9.1

CURRENT EXPLORATION

Significant exploration progress and successes have been made over the life of the Stawell Project, including the discovery of the current Mineral Resources and Mineral Reserves. In the period from January 2015 to December 2015 exploration success of note is the conversion of the Federal Albion South, 250L Below Scotchmans and Upper South Fault 2 areas to Mineral Reserves (Figure 9-1).

Near mine exploration during the period concentrated on adding to Mineral Resource in the underground environs and testing extensions to the mineralization for potential conversion in the near to mid-term, this includes the 468L Mid-North Magdala, Golden Gift Offset, Upper South Fault 5, Aurora A and Aurora B programs. Surface exploration focused on the area immediately east of the Wonga open pit known as the Brummigans Prospect.

Exploration work in 2016 is budgeted at A$2.5M and will be primarily directed toward larger remnant areas on the west flank of Magdala (Federal Albion South Up dip Extension, USF2, 486L Mid-North Magdala, Mariners 109) with the view to add these to Mineral Resource, Table 9-2. Work on the east flank will continue on Aurora B with the aim of increasing the Mineral Resource along the east flank.

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FIGURE 9-4 MINE GEOLOGY LONGITUDINAL SECTION OUTLINGING PROPOSED NEAR MINE EXPLORATION FOR 2016, WEST FLANK

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FIGURE 9-5 MINE GEOLOGY LONGITUDINAL SECTION OUTLINGING PROPOSED NEAR MINE EXPLORATION FOR 2016, EAST FLANK

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  9.1.1

EXPENDITURE

Total geological expenditure (both near mine and regional) for the period 1st January 2015 to the 31st December 2015 was $1.09M. A breakdown of expenditure by quarter is included in Table 9-1.

  Q1, 2015 Q2, 2015 Q3, 2015 Q4, 2015 2015 TOTAL
Expenditure $346,400 $236,900 $416,400 $93,400 $1,093,100

TABLE 9-1 EXPENDITURE ($AUD) BY YEAR FOR THE PERIOD JANUARY 2015 TO DECEMBER 2015

  9.2

MAGDALA WEST FLANK

Over the last 30+ years of operation at Stawell Gold Mines, mineralization proximal to the western flank of the Magdala Basalt has been the dominant source of ore feed. Since modern operations began in 1984, ~2.3Moz of gold has been extracted. In the past 12 months, exploration on the west flank has focused on the larger areas remaining between depleted mining areas.

  9.2.1

FEDERAL ALBION SOUTH (FAS)

The Federal-Albion South target is located at the very southern part of the Magdala Mineralized System and includes the Central Lode and Extended Basalt Contact Lode. During the previous reporting period (April 2012 – December 2014) part of the area was converted to inferred Mineral Resource. Work completed in 2015, comprising 12 holes totaling 1,634.5 meters, extended this area a further 140m x 100m to the South. As a result Mineral Resource and Reserve conversion was possible (see Section 14.3.6.1) .

The area remains open to the immediate South but is limited due to the South Fault and Transverse Fault; both have post-mineralization timing and truncate the Magdala ore body. The area is also open up dip another 115m to the base of the transitional zone.

  9.2.2

UPPER SOUTH FAULT 2 (USF2)

The USF2 target comprises three sub-vertical Mineralized structures, the Central Lode, Stockworks and Dukes Basalt Contact Lode mineralization, which are bounded by the Upper South Fault above and Lower South Fault below. The Upper South Blocks are a fault offset of the Magdala Orebody by movement along the Upper and Lower South faults during the Early South Fault event, which were then re-activated during the South Fault event.

USF2 is bound along strike to the north by the USF2 mined area and to the south by the convergence of the bounding faults. A target area of 100m strike length remained; drilling during 2015 has scoped this gap with four holes drilled totaling 482.2m. Drilling resulted in successful conversion to Mineral Resource and Reserve (see Section 14.3.6.3) .

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  9.2.3

250L BELOW SCOTCHMANS

The 250L Below Scotchmans drilling program targeted the Central Lode in the upper Magdala System. In the Upper Magdala, the Central Lode becomes more complex as it reaches the basalt ‘nose’, the ‘butressing’ effects of the basalt causes the Central Lode to split with part of the lode continuing on the same dip and the other part flattening over the crest of the basalt to create a ‘Flat Lode’. There is also minor mineralization in the Hangingwall Lode (a thin, quartz lode on the hangingwall side of the Central Lode) and the Magdala Basalt contact lode.

Drilling targeted a 200m strike length, where a gap in the drilling database existed. A program of 21 holes totaling 1,202.5m was successful in converting part of the area to inferred and indicated Mineral Resource (see Section 14.3.6.2) .

  9.2.4

486L MID NORTH MAGDALA

The 486L Mid-North Magdala target is located above the Magdala Basalt Nose and below the Scotchmans Fault. Gold mineralization is hosted in the Central Lode, which in this position changes orientation to flatten within the sediment over the top of the Magdala Basalt Nose. Flat lying splay faults off the Scotchmans fault cause minor displacement of the Central Lode with top to the southeast movement. Dilation of the Central Lode tends to occur over the top of the basalt nose and also where the splay faults intersect the Central Lode; these dilation zones have been found to be associated with higher grades in other areas of the Magdala Orebody.

During 2015, a 9 drill hole program totaling 1,536 meters was drilled to investigate an area of 200m x 50m. Results demonstrate the Central Lode mineralization to be continuous, with the northernmost hole returning strong gold mineralization. The area remains untested down plunge to the north.

  9.2.5

GOLDEN GIFT OFFSET

The Golden Gift Offset is located to the south of the Federal Albion South area along the basalt contact. Strong basalt contact mineralization was identified in hole SD452 (3.0m @ 6.5 g/t Au, drilled in 1990) and is interpreted to be the continuation of the strong basalt contact mineralized shoots typical of the Golden Gift, but in the hangingwall of the South Fault. A program of 4 diamond drill holes for 1,237.9m did not identify the strong basalt contact mineralization, results matched more closely to those of the nearby Federal Albion South. Any further drilling in this area will be part of the Federal Albion South programs.

  9.2.6

UPPER SOUTH FAULT 5 (USF5)

The USF5 target is located in the same structural zone as the USF1 and USF2, but beneath the Federal Albion South area. Drilling of the Federal Albion South in 2014 and 2015 has given greater control on the geometry of the South Fault in this position, indicating a ‘kink’ in the South Fault surface. A ‘kink’ can be indicative of a splay fault which could create a block of basalt contact mineralization and Central Lode to become offset. In addition, hole SD610 (drilled 2000) returned 4.0m @ 5.9 g/t Au from the basalt contact in the footwall which shows a clear offset to the basalt contact above the South Fault. A drill program of 5 holes for 961.7 meters was targeted around SD610, the program confirmed the presence of the USF5 block, but failed to demonstrate any continuation in mineralization along the basalt contact or any evidence of the Central Lode. No further work is recommended for USF5.

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  9.3

MAGDALA EAST FLANK

The east flank of Magdala includes the whole ~3km strike length of the eastern flank of the Magdala Basalt. There is no record of gold production from the east flank and little exploration has taken place with priority given to the already accessible west flank (located 0.5 -1km to the west). During 2015, 260m of development was advanced towards the east flank to create a drill platform to expedite exploration (see Figure 9-3). The East flank has been divided into two areas with the South Fault as the boundary: Aurora B in the hangingwall and Aurora A in the footwall.

  9.3.1

AURORA B

The Aurora B is located along the east flank, in the hangingwall of the South Fault. The target is located between -350mRL and -600mRL. The area was identified as a target as a result of a revision of the 2010/2011 Aurora A drilling (formerly known as the Wonga Gift). The Aurora A sits in the footwall of the South fault; the basalt contact mineralized surface has been interpreted to continue north to the South Fault where it has been offset. From work completed during the discovery of the Golden Gift in 1999 the direction and magnitude of movement along the South Fault is known. The interpreted zone of Aurora A mineralization was then traced to its offset location in the hanging wall of the South Fault, only to find drill hole, SD577 (drilled 1998), with an intercept of 8.05m @ 3.2 g/t Au.

Phase 1 drilling was conducted Q2, 2015 targeting both directions along strike from the SD577 intercept. A program of two diamond drill holes for 1,175.3 meters were completed with both holes returning gold mineralization in a zone of altered sediment about 10m from the basalt contact, termed the ‘Hampshire Lode’. Diamond drill hole MD6339 returned 8.3m (Estimated True Width) @ 7.06 g/t Au including 0.4m @ 52.9 g/t Au from the Hampshire Lode. Phase 2 drilling was then conducted in Q3, 2015 with 4 diamond holes totaling 2,068.9m targeting around the Phase 1 holes. All holes intersected alteration associated with the Hampshire Lode with diamond drill hole MD6347 returning 5.0m @ 8.03 g/t Au including 0.45m @ 51.8 g/t Au.

Phase 3 drilling was conducted Q4, 2015 with 13 diamond drill holes totaling 4,767.7m. Drill spacing was approximately 50m x 50m concentrating on the 150m x 150m prospective area delineated by Phase 1 and 2 drilling. Drilling has resulted in successful conversion to an inferred resource and has been added to Mineral Inventory (Figure 9-5). Drilling also confirms the continued presence of the Hampshire Lode as well as mineralization within a waterloo in the basalt, but with the added complexity of flat faulting slightly offsetting the lodes as seen throughout the Magdala ore body.

The geological understanding of the east flank has increased considerably with the three phases of diamond drilling. Of note is that the east flank of a particular basalt nose appears to be most prospective. This will be the focus for future exploration.

  9.3.2

AURORA A

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The Aurora A has formerly been reported under the name ‘Wonga Gift’; this includes the 2010/2011 surface diamond program in which gold mineralization was confirmed on the east flank of the Magdala Basalt. During 2015, a single underground diamond drill hole of 1,301.4 meters was completed to test the northern extension of the Aurora A mineralization. The northern extension was targeted as this is closer to underground infrastructure and could potentially provide future access to the 2010/2011 drill area. Diamond drill hole MD6358 successfully intersected the eastern basalt contact in the footwall of the South Fault with altered sediment (predominantly chlorite) on the contact. The result of the program shows that the basalt contact continues another 500m north along strike from the 2010/2011 drill area and this, combined with 2015 Aurora B results, is demonstrative of the whole 1.5km strike length to the South Fault being prospective.

  9.4

WONGA PIT (BRUMMIGANS)

A program of air core drilling was completed to the immediate east of the Wonga Open Pit. The program had three aims: To test the Mt Micke stockpile, to test for in situ mineralization under the Mt Micke Stockpile and to test the Brummigans structure where it is projected to occur at the surface. The Brummigans structure has been identified at depth in earlier diamond drill holes and is interpreted as a repeat of the Wonga main Hangingwall lode to the immediate east of the Wonga Open pit. Like the Hangingwall lode, the structure strikes approximately north-south and dips at 30-40° to the east; this orientation is interpreted to mean the Brummigans structure will project to the surface immediately east of the current Wonga Open Pit.

A program of 127 air core holes totaling 4,327 meters was complete on a 20 x 15m spacing to test the top 30 vertical meters of bedrock. The results confirm the presence of the Brummigans structure but gold mineralization is not as strong as indicated by deeper diamond drill holes. The structure also remains open to the south.

  9.5

REGIONAL EXPLORATION

Regional exploration focuses on the northern extension of the Moornambool Metamorphic Complex (MMC) looking for repeats of the Magdala system. At Stawell, the Palaeozoic basement outcrops, but 10km further north it is obscured by the much younger Cenozoic age Murray Basin Sediments (clays and flowing sands), which become deeper further north. Identification of basalt bodies in the corridor is initially by aeromagnetic data and regional gravity data where they show up as coincident gravity and magnetic anomalies. Ground geophysics is then conducted to better refine drill targets before drilling is then used to test the basement geology. Relying solely on geophysics has created difficulties, as areas of potential mineralization along the extensive geophysical anomalies have not been tested. To address this, a large Ionic Leach (ALS Laboratory trademark) soil sampling program was conducted in 2012-2013 providing the geochemical information to complement the existing geophysical dataset in order to further define robust drill targets.

No regional exploration has occurred during the reporting period as the exploration focus has been on extending the mine life at SGM through near mine exploration. SGM has been exploring the north for over 20 years, a summary of the main prospects and work completed is included in Table 9-2.

Locations for all regional exploration prospects are shown in Figure 9-6.

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Prospect Status Mineralization
Model
Size Cover
Depth
Geophysics Geochemistry Drilling
Wildwood Confirmed Cambrian basalt and mineralization Magdala type Basalt
contact hosted
1,300m x 400m 10-20m
Murray
Basin seds

CSAMT/AMT

TEM

IP

Detailed Gravity

MIMDAS IP & MT

35 BLEG Soil Samples

1,331 Ionic Leach soil samples

AC

RC

DD

212 holes(11,271m)

269 holes (20,002m)

99 holes (23,412m)

Stawell
Fault North
Confirmed structure and mineralization Shear hosted quartz
vein
7,500m x 750m 2-15m
Murray
Basin seds

AC

RC

DD

9 holes(550m)

14 holes(1,145m)

5 holes(1,051.5m)

Glenorchy Confirmed sheared
intebedded schist/basalt and mineralization
Shear hosted quartz
vein
9,000m x 2,000m 35-70m
Murray
Basin seds

AMT (single line)

Detailed Gravity

AC

RC

DD

270 holes(19,276m)

25 holes(2,728m)

4 holes(1,088m)

Browns Confirmed intebedded schist/basalt and
mineralization
Shear hosted quartz
vein
7,000m x 700m Nil Detailed Gravity

16 Auger samples

963 Ionic Leach soil samples

AC

RC

DD

365 holes(26,675m)

13 holes(889m)

16 holes(5,794m)

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Germania Old workings (quartz vein),
coincident magnetic and
gravy anomaly
Magdala type Basalt
contact hosted,
quartz shear hosted
1,400m x 900m 0-7m
Murray
Basin seds

Detailed Gravity

Passive MT

1,007 Ionic Leach Soil
samples

AC

DD

25 holes(1,252m)

5 holes(555.4m)

Commercial
Road

Confirmed Cambrian
basalt, no mineralization
as yet
Not determined at
present
2,000m x 400m Nil

Detailed Gravity

Passive MT

DD 4 holes(851m)
Ashens

Confirmed sheared
intebedded schist/basalt and mineralization

Shear hosted quartz
vein, Wonga type
10,000 m x 3,000m 60m
Murray
Basin seds

TEM

Detailed Gravity

266 BLEG soil samples

AC

DD

185 holes (17,991m)

2 holes (894m)

Lubeck Confirmed intebedded
schist/basalt, no
mineralization as yet
Shear hosted quartz
vein
3,000m x 600m 50m
Murray
Basin seds

Detailed Gravity

MIMDAS MT

AC 41 holes (3,643m)
Wal Wal Confirmed Cambrian
basalt and mineralization
Magdala type Basalt
contact hosted,
quartz shear hosted
4,000m x 1,000m 35m-45m
Murray
Basin seds

Detailed Gravity

MIMDAS MT

493 Ionic Leach soil samples

AC

DD

289 holes (17,477.5m)

3 holes (970.8m)

Holts &
Bismark
Confirmed Cambrian
basalt, no mineralization
as yet
Magdala type Basalt
contact hosted,
quartz
2,000m x 700m 50m
Murray
Basin seds
27 BLEG soil samples AC 21 holes (1,537m)

TABLE 9-2 SUMMARY OF PAST EXPLORATION ON REGIONAL PROSPECTS WITHIN SGM HELD TENEMENTS

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FIGURE 9-6 REGIONAL TENEMENTS ELS 3008, 5443 AND 5474 SUPERIMPOSED ON AEROMAGNETIC TMI IMAGE OUTLINING REGIONAL EXPLORATION PROSPECTS

  9.6

ONGOING EXPLORATION PROGRAM

Ongoing exploration for 2016 on the west flank will continue to investigate underground areas close to current infrastructure, which can be quickly added to Mineral Resources. Exploration on the east flank will concentrate on the Aurora B Prospect with the aim of expanding the Mineral Resource base. More regional surface exploration will continue to be conducted based on results from the Ionic Leach soil sampling program conducted in 2012-2013, drill targeting the highest prioritized gold anomalies.

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These targets and proposed exploration expenditure are summarized in Figure 9-4, Figure 9-5, Figure 9-7 and Table 9-3; all areas have had various levels of exploration over recent years.

Prospect

Diamond Drilling ($)

Federal Albion South

$274,000

U2

$97,000

486L Mid-North Magdala

$91,000

475L Federal Albion

$170,000

Mariners 109

$130,000

Aurora B

$553,000

Moray

$240,000

Brummigans Structure

$240,000

Germania Prospect

$360,000

Wal Wal Prospect

$327,000

Totals

$2,482,000

TABLE 9-3 FORECAST EXPLORATION DRILLING ($A) FOR 2016

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FIGURE 9-7 REGIONAL TENEMENTS ELS 3008, 5443 AND 5474 SUPERIMPOSED ON AEROMAGNETIC TMI IMAGE OUTLINING ONGOING REGIONAL EXPLORATION FOR 2016

Beyond 2016, exploration will continue to concentrate on near mine prospects with the objective of increasing Mineral Resource definition, dependent upon successful results returned in 2016. Regional exploration will continue to investigate potential mineralization below the Murray Basin sediments to the north of Stawell, expanding the Ionic Leach soil sampling area and drill testing gold soil anomalies identified in the 2012-2013 Ionic Leach soil programs.

Preliminary proposed exploration expenditure is summarized in Table 9-4.

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  2017 2018
Prospect  Geochemistry
($)
 Geophysics
($)
 
 Drilling
($)
 Geochemistry
($)

  Geophysics
($)
 
 Drilling
($)
Federal Albion South 0.27
USF2     0.1      
486L Mid-North     0.1     0.2
475L Federal Albion 0.17
Mariners 109     0.13     0.23
Aurora B     0.55     0.5
Moray     0.24     0.5
Brummigans     0.24      
Germania     0.36   0.01 0.25
Wal Wal     0.33        0.04   0.25
Regional Scoping              0.05 0.05  
Sub Total              0.09 0.06 1.93
Yearly Total 2.49 2.08
TOTAL 4.57

TABLE 9-4 PROPOSED EXPLORATION EXPENDITURE (A$1,000,000) FOR THE PERIOD 2017-2018

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  9.6.1

FEDERAL ALBION SOUTH

Successful drilling during 2015 has led to further Mineral Resources being defined and Reserves established in the Federal Albion South area. Mineralization on both the Central Lode and Extended Basalt Lode remains open further south, but is constrained by the South and Transverse Faults. The area remains open up dip 115m to the base of the transitional zone. A program consisting of 11 holes totaling 1,825m of diamond drilling has been designed to scope the up dip area and, if successful, add to the Federal Albion South Mineral Resource.

  9.6.2

UPPER SOUTH FAULT 2 (USF2)

Drilling planned for 2016 will target the 100m northern strike extent from the 2015 drilling area, where no defined Mineral Resource currently exists. The proposed drilling, consisting of 6 holes totaling 745m will focus on the lower offset block, which has the greatest vertical extent at 20m. Targeted mineralized zones will include the Central Lode, Stockworks Lode and Basalt Contact Lode. If successful, the program could potentially add to the USF2 Mineral Resource.

  9.6.3

486L MID-NORTH MAGDALA

Drilling completed in 2015 on the 486L Mid-North Magdala confirmed the presence of strong mineralization associated with the Central Lode between the Magdala Basalt Nose and the Scotchmans Fault. The northernmost hole returned strong results and show mineralization to be highly prospective down plunge. A platform to drill this target is currently unavailable due to void conditions. Work is underway to investigate access to the area and if established will allow for drilling of the down plunge extension. A program of 4 holes totaling 700 meters could scope the 230m down plunge strike extension beyond the northernmost hole.

  9.6.4

475L FEDERAL ALBION

A review of remnant underground areas has led to the identification of an area 150m x 50m in the mid part of the Magdala system, untouched by previous mining activities. Previous drilling indicates a continuation of the Central Lode in the area. A program consisting of 1,100m will scope the area and, if successful will allow for conversion to Mineral Resources.

  9.6.5

MARINERS 109 DOWN PLUNGE EXTENSION

Drilling during 2014 in the Mariners down plunge Extension has identified strong mineralization within a quartz vein above the Scotchmans Splay, which sits just above and outside the Mariners Underground area. Mineralization above the Scotchmans Splay is related to the Big Hill mineralization as it sits directly down plunge and shows similar characteristics. The mineralized quartz vein intersected in 2014 occurs within the down plunge extension of the Big Hill Mariners 109 Lode. A program consisting of 800m of diamond drilling will scope a 300m x 80m area within the unweathered extension of the Mariners 109 Lode. Successful results will require a further 1,400 diamond drill metres for conversion to Mineral Resources.

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  9.6.6

AURORA B

The Hampshire Lode and ‘waterloo’ structures within the Aurora B remain open along strike to both the south and the north and at depth. Exploration in 2016 will concentrate on theses extensions with a total of 3,130m of drilling planned. A program of 4 diamond holes for 1,640m will target the 170m strike length x 100m dip extent to the immediate north of the 2015 phase 3 drilling area. A program of 3 holes for 1,490m will target the 100m strike length x 100m dip extent to the immediate south of the phase 3 drilling area. Successful results could allow for continued conversion to Mineral Resources.

  9.6.7

MORAY

The Moray is located within a large ‘waterloo’ within the middle upper portion of the Magdala Basalt, between the west and east flank of the Magdala Basalt. Basalt contact mineralization occurs on the west flank of the Moray Basalt Nose. A 250m strike length will be targeted from the Aurora B exploration development and represents the portion of the Moray ‘waterloo’, which could be accessed from current mining infrastructure. A program of 8 holes totaling 1,470 meters has been planned.

  9.6.8

BRUMMIGANS STRUCTURE

The Brummigans structure was modelled in the 2008 Wonga Pit Mineral Resource Model as a repeat of the Wonga Main Hangingwall Lode to the immediate east of the Wonga open pit. The modelling was conceptual in nature and no resource estimation was complete. During a structural review of the Wonga Mine and surrounds in the first quarter of 2014, the Brummigans structure was highlighted as a potential target. Diamond drill data from the 2012 on the southeastern area program was re-evaluated and found to support the presence of a continual mineralized structure in the approximate Brummigans position. A surface program of 1,200m will build on the 2015 surface Air Core program, investigating the down dip extension to a vertical depth of 60m.

  9.6.9

GERMANIA

Proposed exploration will build on work completed in 2013, targeting the West Germania Ionic Leach gold soil anomalies, which appear to be associated with the old West Germania Mine. A surface drill program of 1,800 diamond drill meters is proposed to test for the presence of basalt and associated basalt contact lodes to a vertical depth of at least 300m. There is also the potential for mineralized quartz shear lodes, which was the source of ore from the West Germania Mine. Diamond drilling has been chosen as the proposed drilling method due to the vital structural information it returns.

  9.6.10

WAL WAL

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Proposed exploration will build on work completed in 2012, drill testing the Ionic Leach soil anomalies to the west of the Wal Wal basalt. Soil anomalies are interpreted to be sourced by either Magdala-type basalt contact mineralization or Stawell Fault type quartz shear mineralization, which is supported by the existing geophysical data. A surface program of 1,600 diamond drill meters has been designed to test three of the five soil anomalies. Diamond drilling has been chosen as the proposed drilling method due to the vital structural information it returns.

  9.7

EXPLORATION MINERAL TENURE

Stawell Gold Mines comprises a package of tenements covering the Northern MMC covering an area of approximately 720km2 (see Figure 9-8 and Table 9-5). No tenements are under Farm-in or Joint Venture agreements.

Under the Sustainable Development Act, ongoing relinquishment of tenements is required at regular intervals, these being 25% of the original tenement size at the end of year 2 and 35% of the original tenement size at the end of year 4. Amendments to the Sustainable Development Act which came into effect on the 1st Feb 2012 require an additional 20% of the original tenement size at the end of year 7 and 10% of the original tenement size at the end of year 10 (leaving 10% of the original tenement area). Exploration licenses that are more than 10 years old may be renewed for up to an additional 2 years. Further renewal may be given for a period not exceeding 2 years but only in exceptional circumstances and where it can be demonstrated that there is likelihood of identifying a Mineral Resource in the term. Following this term, no further renewals are allowed. Tenements that are greater than 10 years old will have the 7 year and 10 year relinquishments applied (for a total of 30%) at their next renewal. An outline of current Stawell Gold Mines tenements can be found in Table 9-5, note that EL3008 is now the only tenement older than 10 years.

EL3008 is considered a “strategic” exploration tenement under amendments to the Sustainable Development Act and any renewal of the tenement will not be subject to additional relinquishment requirements for a combined total of 5 years from the next renewal date after the introduction of the Amendments (June 2013). Beyond the 5 year term, further renewals for a period up to 2 years may be given only in exceptional circumstances and where it can be demonstrated that there is likelihood of identifying a Mineral Resource in that term. Following this term, no further renewals are allowed.

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FIGURE 9-8 STAWELL GOLD MINES REGIONAL VICTORIA TENEMENT AREAS

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Name Number Area (km 2 ) Annual
Expenditure
commitment
Grant Date Expiry /
Renewal Date
Comments
Wildwood EL3008 363 $120,000 16/12/1988 20/06/2013 Under Renewal
Application
Glenorchy ELA6156 24 TBC Under Application
Barrabool EL5443 319 $78,800 26/11/2013 25/11/2018 Stawell Gold Mines
Managed
North Magdala EL5474 11 $16,650 24/01/2013 23/01/2017 Stawell Gold Mines
Managed
Stawell Gold Mines MIN5260 4 $899,730 31/05/1985 30/05/2020 Stawell Gold Mines
Managed

TABLE 9-5 REGIONAL TENEMENT INFORMATION

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  10

DRILLING


  10.1

STAWELL GOLD MINES MINERAL RESOURCE DEFINITION PROCESS

The Mineral Resource definition process at Stawell Gold Mines is an ongoing activity. Current Mineral Resources extend from surface to -1025mRL (effectively 1025m below surface). Geological information is collected by a variety of methods with the objective of improving the confidence of the Mineral Resource estimates prior to and during the mining process, including grade control drilling for development and stope definition.

Given the continuous and ongoing nature of the Mineral Resource process, the data utilized varies as the mining operations develop towards the Mineral Resource area (Table 10-1). A summary of the main drilling methodologies employed and data types utilized is as follows with specific details on the sampling and assaying methodologies given below.

  Surface RC Drilling

o

Used for definition of near surface resources where diamond drilling is not required for detailed structural definition.

  o

Used as a method for pre collaring deeper diamond drillholes

  o

Drilling completed using 5 ¼ ” Face Sampling Hammers

  o

Samples collected from cyclone discharge

  o

Hole depths vary but are generally less than 200m

  o

Drilling is conducted dry or with sufficient air to ensure collected samples are dry

o

Sample tipped into three tier splitter from crate to ensure equal quantities available to all vanes

  o

Splitter cleaned by shaking/banging/brush/air compressor as necessary between samples

  o

Cyclone cleaned at regular intervals by banging/checked by hand/arm

o

Around 99.5% of drilling was conducted dry. Big Hill surface Mineral Resource has been very effectively de- watered by the underground mining operation. In a handful of holes, samples were damp after a rod change. Where a sample could not be effectively riffle split it was spear sampled using a PVC spear.

o

Sample volumes were not routinely recorded pre -1999. An initial program in which samples were weighed showed little variation in recovered volume by depth or geology and it was decided that routine weighing was unnecessary.

o

In some project areas casing advancing technologies have been utilized to ensure drilling through fill produces reliable samples

o

Undertaken by contract drilling personnel under the supervision of Stawell Gold Mines’ Geology Team


  Surface Diamond Drilling

o Primarily used in initial exploration programs, near surface resource definition and to provide structural and geological information in near surface RC drilling programs

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  o Drilling by wireline methods.
  o Hole sizes PQ3, HQ3, NQ3, HQ2, NQ2, BQ2
  o Hole depths vary from <100m to >2,000m
  o Directional drilling utilized for specific tasks
  o Core orientation devices often utilized to aid in structural interpretation
o Undertaken by contract drilling personnel under the supervision of Stawell Gold Mines’ Geology Team

  Underground Diamond drilling

  o Used at all stages of the geological process, exploration, resource definition and grade control
  o Drilling by conventional and wireline methods
  o Hole sizes HQ3, HQ2, NQ2, BQ2, LTK60, LTK48 (not used post 1997)
  o Hole depths vary from <50m to 1200m
  o Directional drilling utilized for specific tasks
o Undertaken by contract drilling personnel under the supervision of Stawell Gold Mines’ Geology Team

  Open hole percussion sampling “sludge sampling”

  o Used after development of ore drives for final stope definition
  o Hole sizes 89mm open hole
  o Hole depths vary from 5m – 25m
o Samples of cutting of variable length are collected primarily for geological logging of the chips to identify major faults and geological contacts
  o Undertaken by Stawell Gold Mines production blasthole rigs

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Activity     Resource  
Area Target Type Criteria Classification Geological Data available
    Conceptual geological model   Geophysics
    Geophysical anomaly   Mapping
  Conceptual Geochemical anomaly   Wide spaced Exploration Drilling
Exploration Targets       Conceptual geological models
    Geological model confirmed by drilling   Wide Spaced Exploration drilling
  Confirmed Mineralization confirmed by drilling   Assay information
  Targets       Drill logs
    Geological continuity established Pre resource, Broad spaced Grid Drilling
    Ore grade intersections established Resource Detailed cross sectional interpretations
    Preliminary geological model   Assay information
  Scoped Drill spacing 160 m X 120 m   Drill logs
  targets       Geological models
    Geological continuity confirmed   Regular Grid Drilling
    Ore grade intersections continuous   Detailed cross sectional and 3D interpretations
    Geological interpretation modeled   Assay information
    Geostatistical model established   Drill logs
    Drill spacing 80m X 60m   Geological models
Resource       Inferred Geostatistical model
Definition       Resource QA/QC analysis
    Geological continuity confirmed   Regular Grid Drilling
    Ore grade intersections confirmed   Detailed 3D modeling and interpretations
    Geological interpretation modeled   Assay information
    Geostatistical model   Drill logs
    Economic analysis   Geological models
    Drill spacing 40m X 40m to 30m X 30m Indicated Geostatistical model
        Resource QA/QC analysis

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Activity     Resource  
Area Target Type Criteria Classification Geological Data available
    Confident Geological continuity   Close Grid Drilling
    Ore grade intersections confirmed   Detailed 3D modeling and interpretations
    Geological interpretation modeled   Assay information
    Geostatistical model   Drill logs
    Economic analysis   Geological models
Grade   Drill spacing 20m X 20m Indicated Geostatistical model
Control       Resource QA/QC analysis
    Confident Geological continuity   Close Grid Drilling
    Ore grade intersections confirmed   Detailed 3D modeling and interpretations
    Geological interpretation modeled   Assay information
    Geostatistical model   Drill logs
    Economic analysis   Geological models
    Level Development above and below   Geostatistical model
    Drill spacing 20m X 20m to 15m X 15m   QA/QC analysis
    Open-hole sludge drilling Measured Development face mapping sheets and ore runs
        Resource Open-hole sludge drill geological data where required

TABLE 10-1 STAWELL GOLD MINES GEOLOGICAL PROCESSES AND APPROXIMATE DRILL SPACINGS

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  10.2

DRILLING PROCESS

A flow sheet of the diamond drill process from design to implementation is shown in Figure 10-1.

The diamond drill contract personnel provide a daily record of drilling activities for all drill rigs. Data from the daily record sheet is entered daily to a site database for tracking of drilling production and to enable tracking of drilling progress interrogation at a later date.

Geological personnel track the drillhole path and maintain in control of the daily activities of all drill rigs including which drillers were responsible for various sections of the hole should there be issues with core presentation or down hole depths that require clarification. A regime of regular rig audits and inspections are also used to assist with maintaining the high level of core presentation and sample quality. These drill records are kept indefinitely, enabling a review of drillhole information many years after completion of drilling.

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FIGURE 10-1 SGM DIAMOND DRILLING PROCESS FLOWSHEET

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  10.3

DRILL SPACING

Drill spacing varies for exploration, resource definition and grade control programs within the ranges as indicated in Table 10-1. The appropriate Mineral Resource classifications indicated in this Table are a guide only and each Mineral Resource area is classified based on a number of criteria as discussed in Section 14 of this report. Each of the Mineral Resource areas will have drilling at various hole spacing depending upon the stage of resource development and mining activities.

Initially the exploration drilling is carried out on broad spaced targets, and if the continuity of the structure is apparent, as in the case of known basalt bodies, the targets will generally be tested on centers 160m along strike and 120m down dip (160m x 120m).

When exploration is successful in locating appropriately mineralized environments this spacing is closed down to approximately 80m along strike x 60m up and down dip. If updated geological interpretations completed at this drill spacing are able to demonstrate geological continuity and define sufficient grade to complete and define a Mineral Resource it is possible to classify the defined Mineral Resources as Inferred Mineral Resources.

Ongoing drilling will be completed once appropriate drill platforms can be established to enable the drill spacing to be reduced to 30m x 40m centers. At this spacing, if the geological and grade continuity is well constrained, a Mineral Resource could be classified as Indicated Mineral Resource. Generally at this stage, final mine design and scheduling is possible and capital development infrastructure can be designed and commenced to access the area for mining including the design of appropriate platforms to complete ongoing resource definition and grade control diamond drilling.

Grade control diamond drilling targets a drill spacing of at least 15m to 20m, along strike, by 15m to 20m, up and down dip, on the mineralized structures. This drilling is a component of the mine production process and is required to identify any small scale changes in geometry which will affect mining shapes. Where the geometries are complicated by faulting or other geological features then the spacing can locally be closed to 10m x 10m. Following this work, detailed stope and development design is completed and ore development is designed and implemented under survey control. As a general rule, only after completion of development or sufficiently close-spaced diamond drilling will a Mineral Resource be classified as a Measured Mineral Resource.

As development is implemented every face or development round (3.8m to 4.0m spacing) is visited by Stawell Gold Mines geological personnel to map the location of the major contacts and structures exposed by the development. This information is critical in ensuring development is in the correct location and also to provide the detailed geological information required for final stope extraction and stope tonnes and grade determination. Sludge sampling programs are completed only where there is a requirement to gain additional geological information beyond that already available. Drilling is completed in fans of holes drilled up from the development locations (a typical fan is shown in Figure 10-2). These fans are only completed as required but may be as close as 10m along strike and on rare occasions 5m along strike when structural complexity is high or ore geometries vary.

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FIGURE 10-2 TYPICAL SLUDGE DRILLING FANS COMPLETED FROM AN ORE DEVELOPMENT DRIVE

The drive outline is shown as well as the geological mapping collected from each face location. The uphole drill fans are colored by logged geology: green = potentially mineralized volcanogenics, yellow = basalt.

  10.4

DRILLHOLE ORIENTATION


  10.4.1

UNDERGROUND

Where possible, drilling is oriented perpendicular to the structures being tested. The nature of the mineralization at Stawell and the availability of suitable drilling platforms in the underground environment will always result in compromises in the ability to obtain near perpendicular tests of the mineralization. An example of the orientations of the drillholes through the FAS and USF2 Mineral Resource area is shown in Figure 10-3 and Figure 10-4. Similar drillhole orientations relative to the strike and dip of the structure exist in many of the Mineral Resource areas.

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FIGURE 10-3 LONGITUDINAL SECTION VIEW (TOP), CROSS SECTION VIEW (BOTTOM LEFT) AND PLAN VIEW (BOTTOM RIGHT) OF THE FAS MINERALIZED DOMAIN AND ALL UNDERGROUND DIAMOND DRILLHOLES USED TO CONSTRAIN THE MOST RECENT FAS MINERAL RESOURCE ESTIMATE

(NOTE: EXTENDED BASALT LODE = GREEN; CENTRAL LODE = REC; STOCKWORKS = PURPLE).

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FIGURE 10-4 PLAN VIEW (TOP LEFT), CROSS SECTION VIEW (TOP RIGHT) AND LONGITUDINAL SECTION VIEW (BOTTOM) OF THE USF2 MINERALIZED DOMAINS SHOWING ALL UNDERGROUND DIAMOND DRILLHOLES USED TO CONSTRAIN THE MOST RECENT MINERAL RESOURCE ESTIMATE

(NOTE: DUKES BASALT LODE = GREEN; CENTRAL LODE = RED; STOCKWORKS = LIGHT GREEN).

  10.4.2

SURFACE

The nature of the mineralization around Big Hill near surface and the availability of suitable surface drilling platforms due to community constraints will always result in compromises in the ability to obtain near perpendicular tests of the mineralization. Overall the Big Hill Mineral Resource area has been drilled at regular 20m along strike intervals with the drill holes oriented perpendicular to the strike and dip of the main mineralization system with an up and down dip spacing of 25m, see Figure 10-5 and Figure 10-6.

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FIGURE 10-5 PLAN VIEW (TOP) AND CROSS SECTION VIEWS (A-B AND C-D)

(NOTE: POTENTIAL PIT = GREY, SURFACE TOPOGRAPHY INCLUDING PREVIOUSLY MINED DAVIS PIT = RED, DAVIS LOD E = LIGHT GREEN, IRON DUKE LODE = DARK GREEN, ALLENS LODE = LIGHT BLUE AND MARINERS LODE = DARK BLUE

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FIGURE 10-6 PLAN SHOWING BIG HILL PIT USED FOR RESOURCE REPORTING, DRILL COVERAGE AND DRILL ORIENTATIONS

Noting recent 2012 RC infill drilling in red and 2013 in blue. Used in both the Mineral Resource estimate and void modelling

  10.5

COLLAR SURVEY CONTROL

All survey control for the underground drilling programs is established by Stawell Gold Mines survey personnel. Survey control points are maintained in the underground decline by Stawell Gold Mines survey personnel and these locations provide the control for all mark out and pick-up surveying that is conducted in the underground environment. On conclusion of drilling and drillhole grouting, diamond drilling personnel will insert a wooden wedge labeled with the drillhole ID into the collar of the hole. This provides permanent

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FIGURE 10-7 PLOT OF DRILLHOLE SURVEY METHOD

(SE = EASTMAN SINGLE SHOT IN PURPLE, SD = ELECTRONIC SURVEY INSTRUMENT IN BLUE, SM = MULTISHOT SURVEY IN ORANGE) BY TIME. POST 2001, THE STANDARD SURVEY INSTRUMENT USED HAS BEEN AN ELECTRONIC SINGLE SHOT DOWNHOLE SURVEY TOOL.

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FIGURE 10-8 MAGNETIC DECLINATION CORRECTION AS CURRENTLY APPLIED TO STAWELL GOLD MINES DRILLHOLE DATA

  10.7

DOWNHOLE SURVEY QUALITY CONTROL

Several quality control and quality assurance processes are in place to ensure that appropriate survey (downhole and collar) information is stored to the database. Along with the database manager’s checklist, the project geologist reviews the single shot and multishot data and makes a final decision on the expected trajectory of the hole azimuth. This is illustrated in Figure 10-9.

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For longer drillhole traces, the survey information is plotted to provide a graphical review of the information method utilized where adjustments to the survey information can be made using the overall trend of the drillhole trace (see Figure 10-9).

FIGURE 10-9 EXAMPLE OF THE CHECK PLOTS USED TO CORRECT DOWNHOLE SURVEY AZIMUTH INFORMATION

(GREEN = DESIGN, RED = SINGLE SHOT SURVEY, ORANGE= MULTISHOT 3M SURVEY, BLUE = GE OLOGIST SMOOTHED AZIMUTH)

Where clear discrepancies have been identified with the validity of the survey information and adjusted surveys entered, the original surveys are given a lower priority in the database system. A record of survey methods and or adjustments are maintained in the main acQuire database as part of the audit trail.

Stawell Gold Mines personnel utilize a survey camera test bed with known azimuth and dip to routinely check the accuracy of the downhole survey cameras. This test bed) is located on the surface well away from any potential magnetic sources and is utilized by contract drilling personnel to routinely check camera performance and determine if equipment requires servicing or re-calibration.

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  10.8

DIAMOND DRILL CORE PROCESSING

A detailed flow sheet of core processing activities is shown in Figure 10-1. Prior to drillhole number MD2678 and SD607, the core photography was taken on film and stored as prints.

FIGURE 10- 10 AN EXAMPLE OF THE DIAMOND DRILLCORE PHOTOGRAPHS STORED DIGITALLY FOR ALL DIAMOND DRILLCORE

  10.9

LOGGING

All diamond drill core is logged by the site geological teams using a standardized logging methodology. The data is captured electronically at the point of collection using either a barcode logging “Datcol” software system or acQuire logging system. The “Datcol” system was developed on site in the mid 1990’s and has remained the standard process since that time where the key tables for lithology, alteration, structure, and geotechnical information are populated during the logging process. During 2009, the acQuire logging system was developed to replace the “Datcol” system. The acQuire logging system utilizes the same standard key tables for the lithology, alteration, structure, and geotechnical information, which are populated during the logging process.

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  10.10

CORE RECOVERY

During the logging process, any lost core is estimated and logged as lost core with a specific start and end interval.

A review of database for recently drilled holes indicates exceptionally good core recovery throughout the deposit, particularly adjacent to the major mineralized zones. Where core is lost, it is usually associated with significant faulting. Lost core is identified in the logging as “LOST” and as such there are very few if any assay intervals utilized in the Mineral Resource estimate where core recovery is less than 100%.

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  11

SAMPLE PREPARATION, ANALYSIS AND SECURITY


  11.1

REVERSE CIRCULATION DRILLING SAMPLING

All Stawell Gold Mines RC sampling was carried out using the following protocol:

 

Generally the entire hole was sampled from the collar unless it was recognized as recent fill or material associated with the construction of the drill pads

 

Samples were collected at 1.0m sampling interval, bit pulled back and flushed between intervals

 

Samples discharged into tightly fitting plastic sample bag from cyclone

 

Sample transferred to a rectangular plastic tub, the same size as the splitter

 

Sample tipped into three tier splitter from plastic tub to ensure equal quantities available to all vanes

 

1/12 sub split (nominally 3kg) collected in calico sample bag, tied and placed in lots of five into plastic bags

 

Residue sample collected in original sample bag and transported to bag farm for storage

 

Splitter cleaned by shaking/banging/brush/air compressor as necessary between samples

 

Cyclone cleaned at regular intervals (completion of each hole minimum) by banging/checked by hand/arm

 

Every 20th sample re-split to give a second 3kg sample, field splits dispatched along with first splits

 

Samples dispatched to laboratory for analysis by fire assay.


  11.2

DIAMOND DRILLING SAMPLING

During the logging process, the geologist will mark up the intervals of core required for sampling. Not all diamond core is sampled. Thorough sampling process has identified the key lithological and structural units that will host mineralization and the selection of units for sampling follows the protocols shown below.

  All Magdala Facies, also known as ‘Magdala Volcanogenics’, are sampled for assay
  A minimum of 2.0m into the hangingwall and/or footwall is sampled
  Fault zones and zones of sulphide are sampled at the geologist’s discretion
  Magdala Basalt and Albion Formation units are sampled at the discretion of the logging geologist

Not all diamond core is cut in half prior to sampling. Sampling of diamond drill core follows one of two methods as detailed below.

  1.

Exploration and Resource Definition – HQ or NQ drill programs


  o

Core is logged and geological derived intervals are marked up for sampling

  o

Sample intervals are matched to geological boundaries (structural or lithological) and fall within the range of 0.10m to 2.0m. The average sample interval is approximately 1.0m

  o

For Resource definition drilling programs, 1 in 5 holes is cut with a diamond saw prior to sampling and one half of the core is sent for assay. The remaining half core is retained as a record within the core library. All other drillholes are sampled as whole core, which is sent for sample preparation and assay as per the flow sheet shown in Figure 10-1. All drillholes deemed to be for the purpose of exploration are ½ core sampled and the entire remaining core retained in storage on site at Stawell Gold Mines.

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  2.

Grade control diamond drill programs – NQ or LTK60


  o

Core is logged and geological derived intervals are marked up for sampling

  o

Sample intervals are matched to geological boundaries (structural or lithological) and fall within the range of 0.10m to 2.0m. The average sample interval is approximately 1.0m

  o

For grade control drilling programs drillholes are sampled as whole core with the entire sample sent for sample preparation and assay as per the flow sheet shown in Figure 10-1.

Detailed operating procedures for sampling of diamond drill core are used at Stawell Gold Mines to ensure uniformity of process and prevent errors.

  11.2.1 DIAMOND DRILL CORE SAMPLES

Exploration physicals for January 2015 to December 2015 totaled 10,757.1m of drilling comprising of drillholes into the Golden Gift Offset, Upper South Fault 5, Aurora A and Aurora B targets. The 2015 exploration drill area is shown in Table 11-1.

Resource definition physicals for January 2015 to December 2015 totaled 3,737 comprising 34 drillholes into the SM250, USF2, Mid Magdala and Federal Albion South. Resource definition drilling of FAS and USF2 and SM250 resulted in convertible Mineral Resource and presentation of material into the mine plan.

Grade control physicals for January 2015 to December 2015 totaled 589m comprising 13 drillholes into the SM250. This grade control spacing defined a complex area of the model where the Flat 3 intersected the vertical Central lode.

    2015
EXPLORATION
DRILLING
HOLES 27
meters 10,757
Targets AURORA A, AURORA B, GGO,
RESOURCE DEFINITION
DRILLING
HOLES 34
meters 3,737
Targets SM250 ,USF2, MID MAG, FAS
GRADE CONTROL
DRILLING
HOLES 13
meters 589
Targets SM250

TABLE 11-1 DRILL STATISTICS FOR STAWELL GOLD MINES UNDERGROUND DURING APRIL 2012 -DECEMBER 2015

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The sample interval statistics from January 2015 to December 2015 are shown in Table 11-2. The same sample methodologies are utilized for 2016 and it is anticipated similar sample lengths will be achieved. The average sample intervals demonstrated by the 2015 data are consistent, and are indicative of the complete data set utilized to estimate Mineral Resources at Stawell.

2015 
TOTAL  2,516
MIN 0.2
MAX 1.5
AVERAGE 1.07

TABLE 11-2 DIAMOND DRILL CORE SAMPLE INTERVAL STATISTICS FOR SAMPLES TAKEN DURING 2015

  11.2.2

RELIABILITY OF SAMPLES

It is the opinion of the Authors that the drilling and sampling methods employed by Stawell Gold Mines are of a high standard and provide representative tests of the mineralization suitable for the estimation of Mineral Resources. Standard drill spacings adopted by Stawell Gold Mines are appropriate for the various stages of Mineral Resource classification and whilst other factors also contribute to decisions regarding classification of the Mineral Resources, the drill spacings discussed in this section enable appropriate geological interpretation and Mineral Resource classification decisions to be made.

  11.3

ASSAY LABORATORIES

During the life of the Stawell Gold Mines, a number of laboratories have been utilized for routine assaying of diamond drillcore and RC samples. The details of the laboratories and the periods for which assaying has been conducted are as follows:

  11.3.1

STAWELL GOLD MINE LABORATORY

The Stawell Gold Mine Laboratory is an onsite laboratory which is operated by the mine Staff and is not classified as an independent laboratory. Utilised intermittently prior to 1995 for assaying of diamond drill core and RC samples.

  Non accredited company assay laboratory.
  Assay method was 10g Aqua Regia with pre digest roasting and AAS finish. Samples are roasted before digest to burn off the sulphides, to ensure all gold is released.
  Assaying of diamond drill core and RC samples was discontinued in 1995 and the laboratory sample preparation and assay methods updated to industry standard practice.
  Now utilize a 25g Aqua Regia method with pre-digest roasting and AAS finish.

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  Utilized for Metallurgical assaying, underground face sample and other geological grade control sampling.
  Undertake routine sample preparation of diamond drill core samples – post 2004.

  11.3.2

WMC BALLARAT ASSAY LABORATORY

WMC Ballarat Assay Laboratory utilised prior to 1995 for assaying of diamond drill core and RC samples and was not an independent laboratory at the time of being used.

 

Non accredited company assay laboratory.

 

Assay method was 10g Aqua Regia with pre digest roasting and AAS finish.

 

Assaying of diamond drill core and RC samples was discontinued in 1995 and the laboratory sample preparation and assay methods updated to industry standard practice.


  11.3.3

AMDEL LABORATORY

AMDEL Laboratories is an independent laboratory based in Adelaide, South Australia. The relationship between AMDEL and the Stawell Gold Mines was on a client/supplier arrangement with a contract in place.

 

ISO 9001 accredited

 

Utilized intermittently from 1995 through to present day

 

Primary supplier of assay services from 2004 to mid-2007

 

Ongoing utilization for check assays

 

Stawell Gold Mines reduced reliance on AMDEL Laboratories in mid-2007 as a result of very slow turnaround of assays results


  11.3.4

AMINYA LABORATORY

AMINYA Laboratories was an independent laboratory based in Ballarat. The relationship between AMINYA and the Stawell Gold Mines was on a client/supplier arrangement with a contract in place.

  Not accredited
  Primary supplier of assay services for diamond drill core and RC samples period from 1995 to 2004
  Discontinued in 2004

  11.3.5

INTERTEK GENALYSIS

Intertek Genalysis (formerly Genalysis Laboratory Services) is an independent laboratory based in Perth. The relationship between Intertek and the Stawell Gold Mines was on a client/supplier arrangement with a contract in place.

  Genalysis is a NATA accredited laboratory to ISO 17025.
  Provider of assay services during the period from 2004 to 2006.

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  11.3.6

ALS LABORATORY GROUP

ALS Laboratory Group is an independent laboratory based in Orange NSW. The relationship between ALS and the Stawell Gold Mines is on a client/supplier arrangement with a contract in place.

  ALS Laboratory Group is accredited to ISO 9001 and ISO 17025
  Primary provider of assay services to Stawell Gold Mines from August 2007 to present day

  11.4

SAMPLE PREPARATION

The sample preparation protocol for diamond drill core is shown in the flow sheet given in Figure 11-1. This sample preparation flow sheet was developed in 1995 and has been in operation for all Stawell Gold Mines diamond core and RC samples since that time.

During the period from 1995 to 2004, all sample preparation was conducted by the assay laboratory facilities. In 2004, it was decided by site personnel to complete this task on site at the Stawell Gold Mines laboratory facility. The sample preparation follows the same process utilizing modern sample preparation equipment.

 

Daily primary crusher and pulverizer size fraction calibrations are reported in the weekly Stawell Gold Mines site laboratory report to ensure that the size fractions are meeting the set standards. The size fraction calibration for the crusher is that 75% of material must past through a 2mm screen (LABSOP- 060 Boyd Crusher Size Fraction Analysis). For the pulverizer, the size fraction calibration is that 90% of material must pass through a 75µm screen (LABSOP-061 LM5 Pulverizer Size Fraction Analysis).

 

A quartz flush is inserted, at a 1:5 ratio, at the crushing stage for all diamond drill core and RC chip drill samples. If visible gold is identified at the logging stage then a quartz flush is inserted after every sample within that mineralized zone.


  11.4.1

SPLITTING USING A VIBRATING FEED CONE SPLITTER


  Pulverizing to 95% passing 75um using Labtechnics LM5 pulverizing mills.
  A quartz flush is inserted after every sample at the pulverizing stage for all diamond and RC drill core.

By retaining responsibility for this work through the existing site based facility, Stawell Gold Mines has flexibility in sending the pulps only to a variety of assaying laboratories and also retain the coarse rejects on site for ongoing metallurgical test work programs.

Stawell Gold Mines internal Laboratory process was independently reviewed by ALS Global Laboratory’s February 2016. No serious concerns compromising the quality of data were identified. Outcomes of the audit will be actioned during 2016 to maintain a high quality of sample preparation.

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FIGURE 11-1 STAWELL GOLD MINES DRILL CORE SAMPLE PREPARATION, ASSAY AND QA/QC FLOWSHEET

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  11.5

SAMPLE SECURITY

All drill core and RC samples are delivered directly to the mine site based core farm facility. This is on a shift by shift basis for underground drill core and on a daily basis for all others. Access to the mine site is restricted to authorized personnel only or as a visitor under full supervision by Stawell Gold Mines personnel. A single site access point exists, which is manned 24 hours per day and 7 days per week by security personnel.

Security of drill core and samples is managed by maintaining records throughout the complete process from drilling, core processing, logging, sampling, sample preparation and assaying through to return of results.

Key record keeping procedures utilized in managing sample and data security are described below:

 

Daily drilling records are entered into the database which provides records of drill core produced.

 

Core is photographed within 24 hours of being delivered to the core processing facility.

 

The Stawell Gold Mines sample processing facility is located on the mine lease within a security fenced area. All core stored here is only able to be accessed by Stawell Gold Mines personnel.

 

At the conclusion of logging, a sample requisition sheet generates listing sample numbers, assay standard insertion and assay requirements. This is loaded directly to the acQuire database, enabling tracking of samples after this process.

 

Stawell Gold Mines personnel are trained in appropriate procedures for logging and sampling of the diamond drill core and generating an analytical request sheet outlining sample identification and assay requirements.

 

The production of carefully labeled sample pulps for dispatch by registered posts.

The pulps are dispatched from the Stawell Gold Mines prep laboratory to the assay laboratories using registered post or courier services. Consignments travelling by registered post or courier services are required to be signed off by each leg of the postage route on arrival and can be tracked online. The assay laboratories are also required to send a statement informing Stawell Gold Mines that the pulps have arrived and that the samples, as detailed on the analytical request sheet, can be accounted for. It is the opinion of the Authors that the sample security is adequate.

  11.6

ASSAY METHODS

A summary of the laboratory methods utilized by the various laboratories is given in Figure 11-1. All assaying for gold that is utilized in the Mineral Resource estimates have been completed by fire assay method (30 – 50g charge weights) with an AAS finish.

For samples reporting below LLD, a value of 0.5xLLD is utilized as standard in Mineral Resource estimation.

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  11.7

DATABASE STORAGE AND INTEGRITY

All Stawell Gold Mines drilling data is stored within the “acQuire” Database Management System (the acQuire database) which operates in an SQL server framework, and data security is established by having various levels of user access rights. Stawell Gold Mines maintain a security access system where loading and manipulation of data is only conducted by one of two data managers. All geological personnel have access to the acQuire database for read-only purposes.

Analysis results are received from laboratory in fixed digital format. The load routine imports assay data matching against sample identification created during the logging procedure.

In March 2010, the acQuire database was updated to the CorpAssay ADM (acQuire Data Model) and QA/QC functionality “Assay Pending” has been implemented on analysis results reported since then. Assay results are first loaded to the database with an initial priority indicating that they have not passed through QA/QC. When subsequently reviewed for QA/QC, the priority is changed as the result is accepted or rejected. Only data with a priority of 1 is visible in the MineSight drillhole views ensuring QA/QC of drill data in the model has been approved.

Data validation occurs during upload of data to the acQuire database using the acQuire DBMS. Checks include:

 

All alphanumeric codes (e.g. lithology) are valid and not duplicated

 

All numeric fields are within acceptable limits and not duplicated

 

Sample from-to depths cannot be greater than the maximum hole depth

 

Checks are performed for overlapping samples

Analysis results are received from laboratory in fixed digital format. The load routine imports assay data matching against sample identification created during logging procedure.

Alpha analysis codes are stored as logged and/or reported e.g. NS (Not Sampled), IS (Insufficient Sample), <0.01. The database MetaAssayExport table records equivalent values which are substituted by client software (e.g. MineSight). The convention for defined values is a numeric value half detection limit for results at LLD, and for all other codes, -1 is substituted.

After data compilation is complete, it is critically reviewed by geologists with on-going scrutiny using logs, section/plan plotting and 3D modelling.

  11.8

QUALITY ASSURANCE/QUALITY CONTROL

The general flow sheet for the sample preparation and assaying including the QA/QC samples submitted to ensure this compliance is shown in Figure 10-1.

  Exploration diamond drill core is routinely half core sampled.
  Mine diamond drill core is mostly full core sampled, with approximately 1 in every 5 holes being half core sampled. Samples are crushed and pulverized.

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Routine assaying of diamond drill core samples has been undertaken utilizing fire assay methodologies (30 gram charges) with an AAS finish.

 

Screen fire assay is also completed on samples in which visible Au was observed during the logging process.

 

Assays are reported to an LLD of 0.01ppm. For sample reporting below LLD, a value of 0.5 x LLD is used in resource estimation.

 

Sample QA/QC procedure incorporates routine check assays including repeats (re-assay or repeat assay), duplicate (second sample taken after pulverization), splits (half coarse sample split, at ratio of approximately 1:10 samples), and standards.

 

The Stawell Gold Mines QA/QC process was independently reviewed by Quantitative Group consultants in July 2011 and February 2012. Quantitative Group found the system in place to be a robust and appropriate process for ensuring quality assay returns (Stewart July 2011 and Stewart Feb 2012). The Authors agree with the conclusions reached by Quantitative Group.


  11.8.1

QA/QC CHECKS AND ACTIONS

A range of checks and resulting actions are in place to monitor the QA/QC of the Stawell Gold Mines data set as set out in the QA/QC flow sheet shown in Figure 11-1. When monitoring these checks, the following guidelines are followed.

    11.8.1.1

STANDARDS

A range of standards (Table 11-3) are regularly inserted at the sampling stage (1:20 ratio) to monitor assay analysis accuracy. Figure 11-2 shows the assay standard performance for the FAS May 2015 Mineral Resource Model update.

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StandardID Start/End Date
Used
Gating Values Comment
Lower Limit

(ppm)
Recommended
Value
(ppm)
Upper
Limit

(ppm)
SGM LowA June 2001 - May
2010
3.18 3.34 3.5 Discontinued
SGM LowB June 2001 - May
2005
3.24 3.54 3.64 Discontinued - no sample left
SGM HighA June 2001 - Oct
2008
3.885 4.2 4.515 Discontinued - performance issues
SGM HighB June 2001 - Feb
2009
4.18 4.54 4.9 Discontinued - no sample left
SGM High Feb 2006 - May
2010
8.77 9.36 9.95 Discontinued
SGM1 Nov 2008 / current 3.49 3.63 3.77  
SGM2 Nov 2008 / current 6.77 7.15 7.53  
SGM3 Aug 2009 / May
2011
2.06 2.21 2.36 Discontinued
OR2Pd Apr 2011 / Jan
2012
0.058 0.89 0.943 Discontinued - no sample left,
replaced with matrix matched standards
OR54Pa Apr 2011 / Jan
2012
2.68 2.9 3.12 Discontinued - no sample left,
replaced with matrix matched standards
OR15h Apr 2011 / Mar
2014
0.97 1.02 1.068 Discontinued - no sample left,
replaced with matrix matched standards
OR10c Apr 2011 / current 6.27 6.6 6.92  
OR62d Apr 2011 / current 9.84 10.36 11.16  
OR12a Apr 2011 / current 11.31 11.79 12.27  
OR17c Jan 2012 / current 2.87 3.04 3.21  
OR15g Jan 2012 / May
2012
0.481 0.527 0.573 Discontinued - no sample left,
replaced with matrix matched standards
OR15f May 2012 / Mar
2014
0.301 0.334 0.366 Discontinued - no sample left,
replaced with matrix matched standards
OR200 Mar 2014 / current 0.316 0.34 0.365  
OR204 Mar 2014 / current 0.966 1.043 1.12  
OR904 Aug 2014/current 0.0407 0.045 0.0493 ppb standard

TABLE 11-3 RANGE OF STANDARDS USED AT STAWELL GOLD MINES (G/T AU)

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FIGURE 11-2 ASSAY STANDARD PERFORMANCE FOR FAS DATA (STANDARD DEVIATION TO THE EXPECTED VALUE) G/T AU

If the standard assay is outside ±3 standard deviations from the known value then ~50% of that batch is automatically repeated, similarly if two standards report outside ±2 standard deviations then ~50% of that batch will be repeated. Entry is made into the onsite QA/QC diary to indicate which sample identifications have been repeated.

The performance over time for each standard used for the resource model update is individually graphed and reviewed. This is to understand if non-compliant standards were well-spread across all standard material in use and if it is likely a result of the tightness of the confidence intervals. The spread of the results for each standard are reviewed to identify any potential respectively.

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FIGURE 11-3 OR200 STANDARD PERFORMANCE FOR FAS MAY 2015 RESOURCE MODEL UPDATE

FIGURE 11-4 OR204 STANDARD PERFORMANCE FOR FAS MAY 2015 RESOURCE MODEL UPDATE

    11.8.1.2

BLANKS

Regular insertion of basalt blanks are inserted at the sampling stage (1:20 ratio), to monitor sample contamination or sample homogeneity. If the blanks are not compliant, checks are initiated to determine the cause of the non-compliance and what remedial action is required. Basalt blank material is non mineralized basalt chip material purchased by site for batch plant process.

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Figure 11-5 shows the May 2015 FAS model update analysis of the basalt blank assay results. This figure shows two basalt blanks failing. Given the Stawell Gold Mines laboratory uses of a quartz flush between each sample at the pulverizing stage, failed blanks are rarely returned. Investigation into the two failed blanks identified no apparent cause for the elevation in assay and the blank were accepted based on an adjacent compliant standards.

FIGURE 11-5 BLANK STANDARD PERFORMANCE FOR FAS MAY 2015 RESOURCE MODEL UPDATE

    11.8.1.3

LABORATORY SPLITS

Laboratory (lab) splits are taken during the sample preparation crush stage to monitor sample preparation and homogeneity. If splits are outside of 10% of the initial assay, investigations into the quality of the sample preparation for the batch are conducted.

Figure 11-6 shows the monitored precision for the sample crush stage within the sample preparation process given for assay preparation from the May 2015 FAS model update drillholes. The graph presented in Figure 11-6 is an indication of the homogeneity of the sample after the crush stage and shows whether error is introduced during the splitting of this crushed material prior to pulverization. This analysis shows more than 90% of the sample split assays have an average relative difference of less than 20% to the original assay value indicating an appropriate level of precision through this stage in the sample preparation.

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Figure 11-7 is a Log/Log graphical representation of the same data, which is also used to analyze the lab split outliers. This graph indicates that the split variation more frequently occurs in the lower grade range (yellow highlight) which is of lesser interest. The higher grade splits are all relatively constant and within acceptable tolerance.

FIGURE 11-6 PRECISION OF THE CRUSHER FOR THE MAY 2015 FAS RESOURCE MODEL UPDATE

FIGURE 11-7 SAMPLE SPLITS (CRUSH SIZE FRACTION SPLIT) COMPARISON FOR MAY 2015 FAS RESOURCE MODEL UPDATE

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  11.8.2

LABORATORY DUPLICATE ASSAYS

Lab splits and duplicates are taken to monitor sample precision. Three sets of duplicate data are collected for Stawell Gold Mines samples.

 

The first duplicate data set is collected at the split stage of the sample preparation, before pulverization; every tenth sample is split with the split portion becoming the B sample (1:10 ratio). This duplicate is referred to as the lab-split at Stawell Gold Mines and monitors the precision of the crush and split stages of the sample preparation.

 

The second duplicate data set is collected at the fire assay stage; randomly, within a 1:20 ratio, a second scoop of the pulp is taken and processed in the same batch to the original sample. This duplicate is referred to as the duplicate at Stawell Gold Mines and monitors the precision of the pulverization sample preparation stage.

 

The third duplicate data set is collected at the fire assay stage; randomly a second scoop of the pulp is taken and processed in a different batch to the original sample. This duplicate is referred to as the repeat at Stawell Gold Mines and monitors the precision of the pulverization sample preparation stage as well as across batch repeatability. This type of duplicate can also be requested if there is a non- compliant standard within the batch.

Where duplicates and repeats are outside of 10% of the original sample, additional duplicate repeats are requested to determine if it is a laboratory issue or associated with coarse gold within the sample.

Figure 11-8 shows the monitored lab duplicate results from the May 2014 USF1 Resource Model update drillholes. The graph presented in Figure 11-8 is an indication of the homogeneity of the sample after the pulverization stage. This analysis shows more than 95% of the duplicate assays have a relative difference of less than 10% to the original assay value indicating an appropriate level of precision through this stage in the sample preparation.

Figure 11-9 is a Log/Log graphical representation of the same data, which is also used to analyze the lab duplicate outliers. This graph indicates that the duplicate variation more frequently occurs in the lower grade range (yellow highlight) which is of lesser interest. The higher grade duplicates are all relatively constant and within acceptable tolerance.

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FIGURE 11-8 PRECISION OF THE LAB DUPLICATE DATA FOR MAY 2015 FAS RESOURCE MODEL UPDATE

FIGURE 11-9 ASSAY DUPLICATE COMPARISON OF THE LAB DUPLICATE DATA FROM MAY 2015 FAS RESOURCE MODEL UPDATE

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FIGURE 11-10 LOG-LOG GRAPHICAL REPRESENTATION OF THE ASSAY REPEATS FROM MAY 2015 FAS RESOURCE MODEL UPDATE

Figure 11-10 is a Log/Log graphical representation of assay repeats which is used to analyze the lab assay repeatability. Repeated assays are compared to initial assays and as duplicate assays are a measure of laboratory precision any large discrepancy is followed up and discussed with the internal and external laboratory managers.

Monitoring of these checks is done within two days of the sample batch return and actioned generally no later than seven days after the return date. Any actions taken during the monitoring process are recorded in the Stawell Gold Mines QA/QC diary, which was set up in September 2007.

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  11.8.3

STAWELL GOLD MINES STANDARD REFERENCE MATERIAL

Nine standard reference materials are currently being used by Stawell Gold Mines, and as per standard procedure are inserted at frequencies of 1:20-1:25 samples to monitor accuracy. All of the current standards, except for OR62d and OR904 are matrix matched. Standard OR904 was introduced into the system in 2015 as a low grade oxide standard for a surface RC drill program. All standards are commercially made by Ore Research and Exploration Pty Ltd (ORE) of Melbourne and certification certificates are available in Stawell Gold Mines records. The details of the standards and when they were introduced to the system are shown in Table 11-3.

  11.8.4

QA/QC FOR THIS TECHNICAL REPORT

For this technical report, an analysis of the QA/QC data returned for the current mined Mineral Resource estimates have been compiled (Table 11-4).

This analysis encompasses all QA/QC data returned to Stawell Gold Mines for the current model updates and serves to demonstrate that a responsible and ongoing approach to managing assay data quality is maintained at Stawell Gold Mines and that assaying information is of a good quality for Mineral Resource estimation.

As a result of QA/QC monitoring processes carried out during the current reporting period, 998 samples were re-assayed from an initial 2,720 assays, for holes reported during this period. The proportion of re-assay is variable for each model update and given time period. Full details of QA/QC results can be found in individual reports by area (see Table 11-5).

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FED ALB           AURORA

 

STH USF2 SM 250 FED ALB MARINERS BIG HILL B

 

Au Au Au Au Au Au Au

Population #

32 167 509 47 192 69 98

# Outside 2σ

1 9 74 9 26 4 7

% Outside 2σ

3% 5% 15% 19% 14% 6% 7%

# Outside 3σ

0 2 22 6 5 1 1

% Outside 3σ

0% 1% 4% 13% 3% 1% 1%

% positive

47% 35% 45% 45% 66% 46% 48%

% negative

41% 57% 38% 51% 32% 52% 43%

% at zero

3 10 15 2 3 1 9%

TOTAL BIAS

-3.4% -0.5% 0.3% -0.8% 1.4% -0.1% 0.8%

TABLE 11-4 QA/QC ASSAY RESULTS

 

      FED   BIG AURORA    

 

   FAS  USF2 SM 250 ALB MARINERS HILL B   TOTAL

Initial assays

98 43 181 297 503 1072 526   2720

Repeated assays

53 21 86 163 397 38 240   998

Repeated assays_Accepted

53 21 75 142 339 38 190 858

% Repeated Assays

54.1% 48.8% 47.5% 54.9% 78.9% 3.5% 45.6%   36.7%

% Accepted Repeated Assays

100.0% 100.0% 87.2% 87.1% 85.4% 3.5% 36.1% 86.0%

TABLE 11-5 REPEATED ASSAY FOR ALL CONTRIBUTING MINERAL RESOURCE AREAS

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  11.8.5

OPINIONS ON SAMPLING

It is the opinion of the Authors that the drilling and sampling methodologies employed by Stawell Gold Mines are of a high standard and provided representative tests of the mineralization for the estimation of Mineral Resources. Standard drill spacing’s adopted by Stawell Gold Mines are appropriate for the stages of Mineral Resource development and whilst other factors contribute to decisions regarding classification of the Mineral Resources the drill spacing’s discussed in this section enable appropriate geological interpretation and Mineral Resource classification decisions to be made.

  11.8.6

RECOMMENDATIONS

The results from the QAQC analysis of drill samples has indicated a good level of confidence in assay grades for use in the resource model. The following recommendations for improvements in the current procedures:

  Standards continue to be inserted on site and independently of the assaying laboratories. The process of matrix matched standards needs to be continued.
  Implementation of outcomes from 2016 internal lab independent audit.
  SGM Laboratory recommences round robin sampling with off-site laboratories.
  Quarterly SGM laboratory audit and process review.
  Continued monthly reporting and meetings with the external independent laboratory.
  The procedure for monitoring the acceptance gates of standards needs to continue to be reviewed regularly with increased sample support to validate the standard reference material precision of grade.

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  12

DATA VERIFICATION

Newmarket Gold utilize specialized industry computer software to manage its drillhole and assay database and employ dedicated personnel to manage the database and apply appropriate QA/QC procedures to maintain the integrity of the data. Data is assessed for errors with respect to standards and blanks prior to loading into the acQuire™ database software. Data is then spatially assessed in a commercially available mining software package (Surpac™) for any other questionable results.

To confirm compliance, the earlier 2012 Stawell Technical Report involved completing various database checks, which did not identify any reportable errors, which would have raised any concerns about the integrity of the data. During the preparation of this technical report, which has included search and lookup of assay results, generation of plans and sections and estimation of Mineral Resources, the Authors did not encounter any difficulties with the database; hence the Authors believe the data/database has been verified to a sufficient level to permit its use and confidence in its reliability.

In addition to the quality control and data verification procedures discussed in detail above, the Qualified Persons preparing the Mineral Resource estimates have further validated the data upon extraction from the database prior to Mineral Resource interpolation. This verification used MineSight™ as the primary tool to identify data problems. This allowed the omission of holes, if they were of questionable quality, for example due to low quality sample techniques or incomplete assaying. When coupled with the more mechanical check processes ensuring high quality is entering the database in the first place, these checks were effective in allowing the Qualified Persons to be confident that the data was geologically coherent and of appropriate quality for the purposes used in this technical report.

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13

MINERAL PROCESSING AND METALLURGICAL TESTING


  13.1

MINERAL PROCESSING

A complete description of the mineral processing techniques used at Stawell Gold Mines is outlined in Section 17-1 of this report.

  13.2

METALLURGICAL TESTWORK

An ongoing program of metallurgical test work is conducted at Stawell Gold Mines. The program utilizes diamond drill core to determine the expected plant recovery for all ore blocks at a stope scale within the immediate and long term mine plan. In addition to this, in areas where previous mining and treatment have occurred, previous actual plant performance is also taken into account.

Samples of the ore and estimated dilution are tested to determine the expected preg rob index and expected gold recovery through the Stawell Gold Mines processing circuit. Samples for metallurgical test work are selected from each ore lode; desirable samples are tested for each stope block. In this way expected rates of recovery can be determined for individual stope blocks, levels and ore lodes. As the metallurgical test work program is an ongoing process, the samples being tested to determine recovery rates can be said to be representative of the current and future production areas.

The results of the test work program (combined with previous performance where applicable) provide an expected plant recovery on a campaign basis. Stawell Gold Mines metallurgists are able to compare the actual versus predicted plant recoveries using the test work results. The preg rob index is a relative scale that indicates preg rob severity per ore source. Determining the preg rob index for each ore source allows for tailoring of processing to suit.

Test work results by area are summarized in Table 13-1 below. Examples of monthly recovery reconciliation by ore designation (Table 13-2) and test work results by area (Table 13-3) are given below.

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TABLE 13-1 STAWELL GOLD MINES RECOVERY IN PERCENT AND PREG-ROB INDEX BY ORE SOURCE

TABLE 13-2 ACTUAL VERSUS EXPECTED RECOVERY – ALL ORE TYPES MONTHLY RECOVERY RECONCILIATION

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TABLE 13-3 METALLURGICAL LEACH TEST WORK RESULTS FOR FEDERAL ALBION ORE SAMPLES

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  14

MINERAL RESOURCE ESTIMATES


  14.1

INTRODUCTION

The geographical locations of the various Mineral Resource and Mineral Reserve areas for Stawell Gold Mines are shown in Figure 14-1 and Figure 14-2.

This estimate is an update of the previously published Mineral Resources and Mineral Reserves in the 2014 Stawell Technical Report.

Over the period from 1st January 2015 to 31st December 2015, 18,892.8 meters of underground exploration, resource definition and stope definition diamond drilling was completed. This drilling has assisted in defining additional Mineral Resources, and when coupled with detailed mine design and mine planning work has resulted in a significant addition to the Mineral Reserve.

This estimate is a compilation of a number of separate models and estimates. Some of the Magdala upper levels areas have been estimated using manual 2D estimation techniques. These individual areas have been reviewed as part of the updated Mineral Resource and Mineral Reserve estimate and complies with the relevant Mineral Resource classifications as defined by NI 43-101.

Post 1998, the majority of the Mineral Resources and Mineral Reserves have been estimated using 3D geological block models. A number of block models have been created for separate geographical areas. The key Mineral Resource and Mineral Reserve areas and the date of the most recent updates are recorded and outlined in Table 14-1.

Table 14-1 summarizes the Resource Models that have been completed during the reporting period. Table 14-2 summarizes the Resource Models that were completed prior to the reporting period and contribute to the current reported Mineral Resource.

MODEL

RELEASE DATE REPORT DATE

Federal Albion South

May-15 FAS Resource Model Report May 2015

Below SM 250

Jul-15 Below SM 250 Resource Model Report July 2015

Upper South Fault 2

Dec-15 Upper South Fault 2 Resource Model Report December 2015

Aurora B

Mar- 16 Aurora B Resource Model Report March 2016

TABLE 14-1 SUMMARY OF MINERAL RESOURCE MODELS UPDATED IN 2015

MODEL

RELEASE
DATE
REPORT DATE

Federal Albion

Apr-12 Mid Magdala Resource Model Report June 2012

Mariners

Dec-13 Mariners Resource Model Report December 2013

Magdala S6000

Aug-14 Magdala S6000 Resource Model Report May 2014

Big Hill

Mar-14 Big Hill Resource Model Report April 2014

TABLE 14-2 SUMMARY OF RESOURCE MODELS COMPLETED PRIOR TO 2015 THAT CONTRIBUTE TO THE MINERAL RESOURCE

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The Big Hill Resource model (March 2014) has been reported as Big Hill Surface (within pit shell and evaluated at a cut-off grade of 0.44 g/t Au). The underground potential of this model includes all Mineral Resource outside of the resource pit shell, and has been reviewed at a 2.0 g/t Au cut-off. The underground component of this model is referred to as Upper Levels.

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FIGURE 14-1 WEST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF MINERAL RESOURCE AND MINERAL RESERVE AREAS AS OF 31 DECEMBER 2015

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FIGURE 14-2 EAST FLANK LONGITUDINAL PROJECTION SHOWING THE LOCATION OF MINERAL RESOURCE AND MINERAL RESERVE AREAS AS OF 31 DECEMBER 2015

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  14.1.1

STOPE RECONCILIATION

The resource calculation methodologies used at Stawell Gold Mines are supported by a significant period of mining and reconciliation information (see Figure 14-1 and Figure 14-2). Figure 14-3 illustrates final design versus actual mill reconciled stope performance since 2003.

FIGURE 14-3 STOPE RECONCILIATION

  14.2

MANUAL 2D MINERAL RESOURCE AND MINERAL RESERVE ESTIMATION METHODOLOGIES

Mineral Resources and Mineral Reserves calculated using manual 2D methodologies are confined to sections of the upper levels at the Magdala Deposit. Underground mining for 2015 had a high reliance on upper levels remnant material which was manually estimated at 67% of milled ounces from underground, with a total site contribution of 58%. This is illustrated in Figure 14-4 and in Table 14-3.

 

SITE UNDERGOUND

2015

TOTAL OZ TOTAL OZ

Reserve P&P

29% 33%

Remnant

58% 67%

Surface Oxide

13%  

TABLE 14-3 SUMMARY OF ORE SOURCE CONTRIBUTION FOR 2015

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FIGURE 14-4 CLASSIFICATION OF MINED MATERIAL FOR 2015

Prior to 1996, manual 2D methodologies were the primary estimation methodology for all Mineral Resources and Mineral Reserves estimated for the Stawell Deposits. Progressively 3D block modelling methodologies were introduced and the remaining areas estimated using the 2D methods are remnant parcels for which detailed 3D models have not yet been created. There is an expectation that prior to mining of these areas additional geological information will be collected and processed to produce 3D block models upon which final economic evaluation and mine design will be completed.

The validity of estimation methodology is supported by many years of mining experience at Stawell Gold Mines where with the addition of appropriate dilution and grade cutting have produced reliable estimates of tonnes grade and contained metal.

The Mineral Resource estimation methodology is based on compiled 1:500 scale longitudinal projections for individual mineralized structures showing the locations of all diamond drillholes or sample intersections.

 

Each mineralized intercept is firstly interpreted on cross section and/or level plan or in 3D to determine the true width of the mineralized structure and to ensure the intercept is plotted on the longitudinal projection for the appropriate structure.

 

Individual assays within the mineralized structure are top cut to 17.0 g/t Au for Magdala and 15.0 g/t Au for Wonga mineralization respectively. The assays are accumulated by length weighting of the individual assay grades.

 

Magdala mineralization is diluted to a minimum width of 3.0m by adding dilution at 1.0 g/t Au.

 

Additional interpretation is completed to confirm the strike and dip of the mineralized structure being modelled to calculate a strike/dip correction factor to be applied to the individual polygon being estimated.

 

Polygon areas are measured for each block to be estimated.

 

The tonnage is calculated by multiplying Area x Mean Width x SG x Strike/Dip Factor.


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  The grade assigned to the block is estimated based on the diluted length weighted intersections deemed to be informing each polygon.

  14.3

COMPUTER 3D MINERAL RESOURCE AND MINERAL RESERVE ESTIMATION METHODOLOGIES


  14.3.1

INTRODUCTION

Post 1997, 3D block modelling methodologies have progressively become the standard method for estimating Mineral Resources and have provided the basis for detailed mine design and the estimation of Mineral Reserves.

This work is carried out in the MineSight software suite which is an industry standard geology and mine planning software package. Detailed Mineral Resource documents for each individual Mineral Resource area are referenced in Section 27. The description below describes the general process that is common to each model.

  14.3.2

DATA TYPES

The estimation of contained gold has been based on assays sourced from surface and underground drilling as detailed in Section 10. The data available as at December 2015 consisted of diamond core samples derived from earlier exploration and mining definition campaigns as well as reverse circulation drill chip samples. Sludge drilling and channel sample results were included in the database but excluded from compositing and subsequent estimation. Also excluded were quality suspect drillholes which are listed in each resource report.

All data was provided in local grid co-ordinates.

  14.3.3

GEOLOGICAL INTERPRETATION


    14.3.3.1

FEDERAL ALBION AND FEDERAL ALBION SOUTH

The mineralized zones of the Magdala Federal Albion all exhibit visually distinctive mineralization styles.

The Central Lode presents as a strongly sheared volcanogenic unit with high quartz content that frequently contain arsenopyrite-filled stylolites. Silica alteration is prevalent, as is moderate to strong sulphide mineralization (predominantly sheared pyrrhotite and cubic pyrite).

The Extended Lode is a Basalt Contact mineralization typified by coarse arsenopyrite up to 20mm in size along with recrystallized pyrrhotite and minor pyrite. Large quartz tension veining up to 750mm wide is often associated with coarse arsenopyrite and other sulphides that form along stylolites and is a common feature of the Basalt Contact mineralization. In the immediate contact zone with the basalt it was common to see silicification of the host rock and bands up to 4m wide of mineralized siliceous sediment. Assay values are weaker in these zones by comparison due to the lack of chlorite.

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A hangingwall weak volcanogenic unit is found on the contact zone of the Mine Schist and the Central Lode, which is less chloritic, has higher graphitic alteration than the host rock, and shows a moderate S2 or S3 fabric more typical of Mine Schist. This domain consists of recrystallized pyrrhotite and sedimentary pyrrhotite found around weak to moderate chlorite altered host rock. Quartz shear zones consisting of puggy graphite and weak chlorite alteration are also present. The hangingwall weak volcanogenic unit was often geologically logged as pelite to distinguish it from the mineralized zone and the Mine Schist proper. A footwall weak volcanogenic unit separates the Central Lode mineralization from the Extended Lode.

A felsic porphyry runs oblique to the Central Lode from 5220N to the South Fault and in places stopes out areas of the Central Lode. This porphyry is post mineralization and does not host any gold.

    14.3.3.2

BELOW SCOTCHMANS 250

The geological interpretation of the Below Scotchmans 250 area eventuated in having six ore domains, basalt waste zones and a halo of low grade volcanogenic (which was estimated).

The Central Lode is defined by a hangingwall proximal to the Mine Schist contact, and a footwall contact defined by a reduction in grade, RQD% and quartz content.

The Magdala Basalt Contact Lode is defined by a footwall basalt contact, a sulphide/chlorite/quartz rich zone adjacent. The hangingwall is defined by a reduction in sulphides, chlorite, and grade and often a reduction in shear intensity. The Magdala Basalt Contact Lode and the basalt are defined and estimated in the 2012 Federal Albion Resource model.

An upper basalt has been modelled higher than the Magdala Basalt of which there is a small zone of Basalt Contact Lode identified which has been modelled as Magdala Contact Lode, but in reality is the contact lode of the next basalt nose up (Federal Basalt).

The Low-Grade Volcanic domain is a zone of weakly mineralized, less chloritic material surrounding the mineralized domains extending out from the basalt to the Mine Schist.

    14.3.3.3

UPPER SOUTH FAULT 2

The geological interpretation of the USF2 area eventuated in having three ore domains and three waste zones. The waste zones are the Basalt and Central Lode Basalt and Mine Schist geological units (grade is set to 0.005 g/t Au). The geological framework is bound by the Upper and Lower South Faults with the internal flatter My Fault and Your Fault laterally offsetting the geology top to the west. The Central Lode has internal basalt stoping or controlling the Central Lode mineralized material. Both the Upper South Fault and the Lower South Fault have been intersected in the 2015 development from the southern end of the USF2. The flat My Fault has also been intersected in both the 752 and 772 levels at the southern end of the model area. The Lower USF2 block is the dominant block in terms of the vertical extent with the My Fault and Your Fault blocks being less vertically extensive. The lateral offset of the blocks is variable from north to south. The USF2 block is constrained by the pinching of the Upper and Lower South Faults to the South and the past USF2 mining to the North.

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The Central Lode is defined by a hangingwall proximal to the Mine Schist contact, and a footwall contact defined by a reduction in grade, RQD% and quartz content.

The Central Lode domain is comprised of the 101 and 102 subdomains, separated based on a curve in the orientation of the ore but with the same characteristics. The vertical extent of the Central Lode is constrained by the Upper and Lower South Faults, but also internally by the My and Your Faults which offset laterally the Central Lode. The 100 domain was used for statistical analysis and used in variography studies.

The Dukes Lode is defined by a footwall basalt contact, a sulphide/chlorite/quartz rich zone adjacent. The hangingwall is defined by a reduction in sulphides, chlorite, gold grade and often a reduction in shear intensity. The Dukes basalt contact has been faulted out in the southern extent of the geology model and has not been intercepted by any drilling.

The Stockworks Lode is a single domain and is defined by a broad zone of variable grade between the Central Lode and Dukes Lode in the southern extent of the modelled area. The Stockworks Zone contains the greatest grade in the southern end of the domain and reduces in grade to the north. The contact between the Central Lode and the Stockworks can be difficult to precisely define. The Stockworks Domain is now outside the area of interest and is essentially depleted.

The Weak Volcanic Domain is a zone of weakly mineralized, less chloritic material from between the Central and Dukes Lodes and excluding the Stockwork Lode. The Weak Volcanic is typically of low grade and less chloritic than the mineralized Central and Dukes Lodes.

    14.3.3.4

MAGDALA S6000

The mineralized zone of the Magdala S6000 is geologically identical to the Magdala Federal Albion area. This is the northern extent of the Magdala ore mineralization as described in Section 14.3.3.1. All ore zones exhibit visually distinctive mineralization styles.

The Central Lode shear zone generally has a well-defined hanging-wall shear along the Mine Schist/Volcanogenic contact but may also occur within Volcanogenic package. The footwall structure is generally marked by shear surface also, but at times gradational into stockworked Volcanogenics. The distribution of grade with the Central Lode is not uniform. The Central Lode is often partially replaced by later stage felsic porphyry dykes running sub parallel to the Central Lode shear, most dominantly on the hangingwall.

The Dukes Lode mineralization comprises dark green, chloritic, Volcanogenics, host to quartz sulphide tension veining. The hangingwall to the Dukes Lode structure is generally defined by a quartz pyrrhotite shear structure, while the footwall is defined by the contact with Dukes Basalt. The basalt contact is often bounded by a 30-50cm wide quartz sulphide vein containing coarse arsenopyrite. The Dukes Lode and Dukes Basalt are on the Hangingwall side of the Extended Basalt.

The Extended Lode mineralization comprises dark green, chloritic, Volcanogenics, host to quartz sulphide tension veining. The hangingwall to the Extended Lode structure is generally defined by a quartz pyrrhotite shear structure, while the footwall is defined by the contact with Extended Basalt. In places, the Central Lode has sheared out part or all of the Extended Lode and forms the hangingwall. The basalt contact is often bounded by a 30-50cm wide quartz sulphide vein containing coarse arsenopyrite.

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The Stockworks is made up of the footwall Volcanogenics between the Central Lode and the Basalt Contact Lodes and is made up of quartz-sulphide tension vein arrays and carbonate spotting is often visible. The Stockworks material is typical of a higher grade in the hangingwall.

    14.3.3.5

MARINERS

The mineralized zones of the Mariners area can be divided into five mineralized lodes, Curiosity Lode, Opportunity Lode, New Lode, Mariners Lode and Spirit Lode. All lodes strike at around 335° and dip around 50° towards the southwest. Each lode is separated by a low grade zone from 2m to 14m thick.

All shear zone lodes differ from the typical Central Lode structure due to having diffuse boundaries on both the hangingwall and footwall of the structures. The Mariners Lode also differs in that it is mostly brecciated quartz rather than a laminated quartz vein. The lode structures are generally identified as a quartz-sulphide rich shear zone with varying widths of 2-10m and have been modelled using elevated quartz-sulphide percentage as a guide, but has also been driven by grade boundaries.

The Spirit Lode is located to the west of the Mariners Lode. In comparison to the other four lodes, the Spirit lode tends to be the most weakly mineralized and with the thinnest width. It disappears along strike in the middle of Mariners and even though evidence of the hangingwall shear can be seen in drill core, there is little to no mineralization within the lode itself. The weaker grade is likely a result of the position of the lode on the Hangingwall Extreme of the Mariners area.

The Mariners Lode is located between the New and Spirit Lodes. As this domain had the most drill intercepts it was the easiest in which to identify the mineralising structure, which was also confirmed in old face and backs mapping in the area, which identified these secondary faults and mineralization associated with them. Once Mariners Lode had been established, the other shears and associated quartz breccia, fine grained sulphides and mineralization on either side of the Mariners Lode could be identified.

The New Lode is located between the Mariners and Opportunity Lodes. This is one of the stronger mineralized lodes, second to the Opportunity Lode.

The Opportunity Lode is located east of the New Lode and is the largest and most strongly mineralized of the five lodes. It is the widest of the modelled lodes and is likely to comprise more than one mineralized shear.

The Curiosity Lode is located east of the Opportunity Lode. The Curiosity Lode is a new addition to the previous model and was identified as mineralization not domained in the footwall of the Opportunity Lode. Recent drilling, particularly in holes MD6253, MD6254 and MD6255, has now provided enough evidence for the delineation of the Curiosity Lode. Drilling, however, is sparse and not all holes have been sampled in this position resulting in low confidence in the Curiosity Lode (classified as inferred).

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    14.3.3.6

BIG HILL AND UPPER LEVELS

The package of Magdala Volcanogenics, as a whole, plays host to the bulk of mineralization. Where continuous mineralized structures can be defined with a reasonable degree of confidence, these structures have been constrained within wire-frame solids. Where it has not been possible to confidently define the geological control to mineralization this material has been left unconstrained. The Magdala Flanks, Mariners, Iron Duke and Allens are geologically modelled in the surface component of the model (Big Hill). The Mariners and Magdala Flanks units are geologically modelled in the underground component of the resource model (Upper Levels).

Mariners is the least problematic geological unit to define. It has a consistent width and orientation, averaging around 14m thick and dipping to the northwest (mine grid). It constitutes a shear package, with massive quartz veining. Less consistent is the Mariners Spur which is modelled on the upper surface (sections 289/290N). This is only defined by 8 (3 are twinned) drillholes and the geology in this area will be more complex than modeled.

Offset extensions of the Mariners Shear have also been modelled down dip. The original model identified one offset extension; however, two further offset domains were identified in the Upper Levels 2012 model update. Underground drilling of Mariners Lower in early 2012 drove an increased understanding of the structural framework, identifying increased faulting complexity and offset lodes which now connect the underground Mariners Resource with the Upper Levels resource model.

The main Mariners Lode is offset by Fault 3, which strikes at 145º (Mine Grid) and dipping at around 35-40º to the northeast. It has an apparent reverse displacement of between 5m and 25m, decreasing northwards. This section of the Mariners Lode (Mariners L1) is then truncated and offset at depth by the Cross Course Fault. In this model update, the Cross Course Fault has been remodelled as two fault surfaces which bounds another lower lode offset (Mariners L2). A third, less horizontally extensive offset (Mariners L3) was identified and was modelled to be bounded by the Cross Course Fault lower surface and the Scotchmans Fault. The Mariners L3 has a greater offset 15m-25m and plunges further to the north than the above two offset lodes, which is likely due to a greater influence from the Scotchmans faulting. In reality these offset sections are likely to be more complexly faulted than modeled. Beneath the Cross Course Fault, the general definition of geology is hampered by a lack of diamond drill core. The principal reason for modelling a fault in this location is to explain the lack of semi-planar continuity of the otherwise predictable Mariners structure. Without a fault, it would be necessary to invoke a 30-40° change in the orientation of the shear. Note that the geological confidence of the interpretation around these faults is lower than the more planar section of Mariners, because of the greater degree of structural complexity and uncertainty in the location of the faults.

Interpretation of the Allens zone is somewhat problematic and the interpretation of the geology in this zone is of lower confidence than other sections of the orebody. Exposures in the Allens open pit show lithological layering (S0 and later S2/3 foliation), as well as massive veining to be very steeply dipping to vertical. Diamond core shows the same. Two massive veins exposed in the Allens Adit strike approximately N-S mine grid and dip vertically. Some Stockwork style veining is seen near the top of the zone (immediately beneath the Mariners structure), but much of the geology consists of well bedded, foliated volcanogenic sediments which are unpredictably mineralized. Much of the gold occurs in these sediments. Because of the highly oxidized state of these rocks the form of sulphide mineralization is largely obliterated.

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There is a poor correlation between quartz percentage and gold mineralization. Accordingly the wire-framed structures which have been interpreted in the Allens zone should not be regarded as having hard boundaries for ore/waste selection. Rather they have been used to constrain a tonnage that should reflect the tonnage of ore grade material adequately. They have dimensions that reflect the attitude of mappable geology, which in turn will probably exert the main influence on gold mineralization. However, the locations of the wire-frames are controlled as much by distribution of gold grades as by definable geological contacts/zones.

It is suspected that a degree of remobilization of gold within the weathering profile may exist, although no distinct enrichment horizons could be identified. The geology is of such variety that it is very difficult to separate any secondary effects from primary variation.

The relationship between the Mariners and Allens zones is also still problematic. Little overlap exists beneath the main flatter section of Mariners and the Allens zone, which dies out rapidly northwards beneath Mariners. No clear evidence is seen of the timing relationships between these structures, the Allens mineralization appears to hang as a pendant of dilational veining beneath a change in orientation of the Mariners shear structure.

The Iron Duke zone is a wedge of geology occurring between the Scotsman’s Fault and the Lower Cross Course Fault. It is the up-dip extension of the Volcanogenic package/shear zone which has been offset from the main Magdala shear system by reverse movement on the Lower Cross Course Fault. It is truncated above by the Scotchman’s Fault zone, which again displaces the orebody westward. Interpretation of the geology in this zone is hampered by the strongly oxidized nature of the rocks. It has been interpreted as a series of constrained stockwork style ore zones (same as Allens), where the envelopes have been used to constrain a tonnage. In reality the margins will have gradational boundaries.

For wireframing purposes the Iron Dukes domain has been considered in the same context as Allens and was treated as broad envelope around the stockwork zones constraining bulk tonnage. The previous model wireframed the structures at a higher grade which effectively reported a higher grade for lower tonnes. A lower domain threshold has been utilized for the current model at (0.35 g/t Au); in line with the model economic cut off (0.44 g/t Au) to ensure all minable tonnes was economically represented for mining analysis.

“Magdala Flanks” is the term used to describe the package of volcanogenic rocks lying to the west of and continuing up dip from the nose of the Basalt Antiform. The western margin of this zone is the contact between the Volcanogenics and the Mine Schist, which often plays host to the Hangingwall shear. The eastern margin is generally the contact with the basalts, but above the basalt noses the eastern margin is a transitional boundary to siliceous eastern schists, usually marked by a clear grade boundary.

    14.3.3.7

AURORA B

The Aurora B zone is the first Mineral Resource area estimated on the east flank of the Magdala Basalt. The Aurora B area is located in the hangingwall of the South Fault, between 232N and 245N, and from -350mRL to -600mRL. The area is terminated to the south and at depth by the South Fault, but depth extent increases further north due to the moderate northeast dip of the South Fault. The area remains open to the north and could continue for the estimated 3km strike length of the Magdala Basalt. The area is complicated by at least 2 Flat Faults with a top to the southeast movement; however, this only has a slight affect as the apparent eastwards dislocation is about 5-15m. The rocks within the Aurora B area show good competency.

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The Aurora B model has three ore domains: the Hampshire Lode, Waterloo Lode and Wonga Style Lode. At present, only one of which has sufficient drill hole information for a Mineral Resource estimation – the Hampshire Lode. The Hampshire Lode is an approximately 5m wide, subvertical lode, dipping 80° to the northeast, and located 10m to the northeast of the East Basalt Nose. The Hampshire Lode strikes parallel to the strike of the East Basalt Nose. There are indications that the dip of the Hampshire Lode shallows over the crest of the East Basalt Nose, but further investigation is required. An area roughly 150m x 150m has been scoped on 50m x 50m spacing for conversion to Inferred Mineral Resources.

The Hampshire Lode is mostly hosted within sulphidized BIF. Sulphides associated with gold mineralization include arsenopyrite, pyrrhotite (replacing magnetite bands) and pyrite (associated with carbonate alteration and brecciation). Free gold has been observed in quartz veins, chlorite veins and within the chloritized groundmass.

The Waterloo Lode is similar to the Waterloo Lodes on the west flank; in the Aurora B area it sits within sediments between the Moray Basalt Nose and the East Basalt Nose. It differs to the west flank waterloos in that BIF can clearly be identified and mineralization is similar to the Hampshire Lode, where the BIF is sulphidized and tends to be associated with gold mineralization. An area 300m x 60m has been modelled, but has insufficient drill hole information for a Mineral Resource estimation. With further drilling, there is the possibility to add this domain to Mineral Resources.

The Wonga Style Lode sits 30m-60m further east of the Hampshire Lode, but dips at a shallower angle of about 60° to the east. An area 190m x 300m has been modelled, but there is insufficient drill hole information for a Mineral Resource estimation. Like the Wonga Deposit, gold mineralization is associated with needle-form arsenopyrite hosted within pre-existing weaknesses in the sediment. In the Aurora B area this pre-existing weakness appears to be a fault in which a feldspar-porphyry dyke has intruded prior to the mineralization event. These similarities and relative timing leads this to be interpreted as being associated with the later Wonga Gold event rather than the earlier, main Magdala Gold event (~440Ma) with which the Hampshire and Waterloo Lodes are associated. With further drilling, there is the possibility to add this domain to Mineral Resources.

  14.3.4

GEOLOGICAL MODELLING

Geological modelling is carried out on individual Mineral Resource areas by the geological team at Stawell Gold Mines. The data available for this varies depending upon the stage of the development and understanding of the deposit area. The general process is as follows:

  Wireframe models of major geological units are interpreted and created using MineSight software.
  Less advanced deposit areas are first interpreted on paper while more advanced deposits are often interpreted directly within the software.
  All available geological information is utilized in the interpretation process including:

  o

Diamond drill core logs, core photographs, face mapping and photographs, and sludge drilling geology logging.

  o

The drillhole logging information available to the mine geologists including:

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  Lithology
  Alteration
  Quartz veining percentage and veining style
  Sulphide percentages, type and style
  Location and orientation of lithological contacts, shears and fault structures
  Core texture – indicating faulting, shearing etc.
  Core photographs

Geological modelling is an ongoing process and models are progressively updated to reflect the most up to date information. Model updates are generally made upon completion of infill drilling programs and completion of development levels, where this information results in a significant change in the amount and quality of data involved and/or changes in geological understanding.

The gross geological architecture of the mineralization systems is well understood and described in detail in Sections 7 and 8 of this technical report. Mineralization is hosted by relatively distinctive and predictable geological units that are modelled by the area geologists. The key units that are modelled in each area are:

 

Magdala Basalt – unmineralized. Geometry of the Magdala Basalt is required to estimate the quantity of dilution that will be incorporated into the mine designs.

 

Mine Schist – unmineralized. Also important for estimating the quantity of dilution that will be incorporated into the mine designs.

 

Weakly Mineralized Volcanogenic – important to estimate dilution in the mine designs.

 

Mineralized domains. For individual model areas these may be either in the Central Lode position, Basalt Contact positions and contain zones of Stockwork Lode mineralization.

 

Key fault structures are modelled as they can have a significant impact on the shape of the mineralized domains.


 

Structures have a significant impact on the outcomes of the geological modelling and improving the understanding of the location of these is a key component to producing reliable estimations of the in situ Mineral Resources.

 

Structures have a significant impact on the in situ Mineral Resources.


  Wireframes are, where possible, snapped to drillhole intervals.

The key control on Mineral Resource estimation is accurate definition of the constraining geological models. Estimation of grade within the domains, whilst still very important, is of secondary importance to the first order geological demining.

  14.3.5

WEATHERING MODEL

A model of weathering state was constructed using the following criteria:

  Base of total oxidation was defined as the point down hole at which weathering is not completely pervasive of the rock mass and where remnant fresh rock first appears.

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  Base of transition zone was defined as the point down hole at which oxidized material becomes an insignificant volume component (<1% by volume).

In diamond core, these boundaries were identified from logging and core photographs. In RC chips identification was principally made from color and mineralogy changes. Many of the holes were re-logged specifically for identification of these boundaries.

Both surfaces are highly variable in reality due to uneven fracture density and fluid flows, and holes may pass in and out of totally weathered, partially weathered and fresh material. The surfaces are therefore necessarily smoothed.

  14.3.6

MINERAL RESOURCE INTERPRETATION BY AREA

The process of block modelling at Stawell Gold Mines is relatively standard across all model areas and is drawn from the details for the individual model areas.

Stawell Gold Mines has retained the services of several key consulting groups over time to ensure Mineral Resource estimation processes have been maintained to a high standard. Quantitative Geosciences (QG) have provided ongoing coaching, training, mentoring, and Mineral Resource estimation services on an as required basis since the late 1990’s. This has ensured consistency of process over this time as key site personnel have changed and these consulting services have provided ongoing support to the resource estimation function.

Resource wireframes were used to code the drill intercepts contained within them by coding the raw drillholes for ‘GEOCD’. This flagging allows the selection of data within domains by codes for the purposes of sample analysis and compositing.

All resource interpretation wireframes have been used as hard boundaries.

    14.3.6.1

FEDERAL ALBION SOUTH

The Federal Albion South (FAS) area has become one of the major mining areas for the 2015 – 2016 mine plan, and as such grade control and resource extension drilling has been undertaken over the Mineral Resource to continue defining the extent and tenor of mineralization to the South. The initial FAS Mineral Resource was produced in February 2014 with a large area of the north of model overlapping with the April 2012 Federal Albion Model, with preliminary variography. A subsequent model update in November 2014 incorporated grade control drilling and undertook QA/QC statistical comparisons of the dataset to validate the dataset. This model produced an initial Mineral Reserve over the area and development access on 4 levels commenced (310; already accessed, 293, 273 and the 253) in early 2015. Another phase of drilling from the 310 exploration cuddy in 2015 defined the penthouse offset to the south fault and infilled the drill spacing to the south to approximately 30m spacing up to the -275mRL. The height of the drilling definition is constrained by the maximum angle of drilling from the 310 drill drive.

The geological framework was updated to include the Penthouse Fault and the Upper South Fault which are later, offsetting the mineralization below to the west (1m to 10m). The Penthouse Fault was intersected in the 310 development. The Penthouse mineralization is a small splay which is bound by the upper South Fault and the Penthouse Fault. Below this, is the U5 mineralization which remains open to the south for further definition. The location of the South fault was redefined to move further north based on the structural data of the Golden Gift offset diamond holes.

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During 2015, an additional 21 new diamond holes were drilled into the FAS project area. This drilling was a combination of grade control and resource definition drilling with Inferred and unclassified areas converted to Indicated status.

    14.3.6.2

BELOW SM250

The Below Scotchmans 250 (SM 250) mineral resource model covers and area from the 281N to 301N below the Scotchmans Fault down to -300mRL all within fresh rock. The below Scotchmans 250 model is the northern extent of the Magdala orebody, with geology dipping to the west. The model area contains the Central and Hangingwall and the Magdala Basalt Contact as well as the two flat structures (1 and 3) and internal splays.

The Below Scotchmans 250 Mineral Resource model includes the northern region of the upper levels in the underground mine environment. Recent drilling in the area has focussed on an unmodelled area between the -212 and -250RL above early mine workings. The Below Scotchmans 250 area is in the (Magdala) basalt flank in a structurally complex area where there is horizontal offset of the basalt and the formation of flat lodes over the basalt noses. The Central lode is bound to the top by the Scotchmans Fault, and a narrow hangingwall lode is defined.

During 2015, 24 new diamond holes were drilled into the Below Scotchmans 250 project area, allowing for definition of inferred and indicated mineral resources across multiple lodes. The model was compiled from previous data collected throughout the 30 years of modern mining of the Magdala Mine, including diamond drilling, sludge drilling, underground drive mapping and previous cross-sectional interpretation.

    14.3.6.3

UPPER SOUTH FAULT 2

The Upper South Fault 2 (USF2) mineral resource model covers the area from the 314N to the 331N on the 45° mine grid and is bounded by the Upper South Fault above and the Lower South Fault below. The geology has structural complexity due to offsetting flat faults which differentiates it from the Magdala geology. The northern area of the model is bound to the north by USF2 Mining on two lodes and from the southern area by the convergence of the Upper and Lower South Faults. Two stopes taken during access development during 2015 and infill diamond drilling has assisted in the understanding of the area during 2015.

Access development into the southern end of the USF2 fault block was initiated in early 2015. Development has included the collection of face samples, sludge samples and a diamond drilling program on both the 752 and 772 mine levels. Short ore drives on each level have been developed and are limited by the plunge of the ore between the Upper and Lower South Faults. Drilling of four diamond holes was undertaken which infilled previous drilling which was sparse to the south end of the model area. The original USF2 Mineral Resource model (2006) focused on the northern area of the USF2 mineral resource and was used to provide structural guidance for the model update.

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Geologically, the USF2 shares the same characteristics as the other Magdala orebodies, comprising of three distinct mineralization types. The mineralization types, which occur in the Magdala System include the quartz rich shear of Central Lode, the chlorite and sulphide rich Basalt Contact Lodes and the Stockwork Lodes.

    14.3.6.4

FEDERAL ALBION

The Mid Magdala mineral resource model (Federal Albion) covers the area from 255N to 295N, and is bounded by the surface or Scotchmans Fault at the top and the South Fault at depth. The geology of the area is generally well recognized as mining activity within the area has been occurring for over 100 years.

At the time of modelling there were thousands of surface and underground diamond drillholes intersecting the orebody at varying intercept spacings. The model consists of 6 ‘ore lodes’ – four basalt noses (or waterloos), the Central Lode, and a separated ‘Stockwork’ Lode. Weakly Mineralized Volcanogenic material is also modeled.

Given the long mining history of the Magdala orebody, a full ‘depletion’ model does not exist, however, many of the mined voids are modeled. As such, reporting from the model was limited to areas that were known to still be intact and totally unmined and as such additional mineral resources could be identified pending accurate void modelling exercises.

During the 2011/2012 period, an additional 23 drillholes were drilled into the Federal Albion area. These drillholes were the basis of the April 2012 model update. This was a combination of grade control and mineral resource definition drilling with Inferred Mineral Resource blocks taken straight to grade control status. There has been no further drilling or model update to the area since 2012.

    14.3.6.5

MAGDALA S6000

The north Magdala S6000 mineral resource model covers the area from 1510N to 2260N, and is bounded by the Scotchmans Fault at the top and the South Fault to a maximum depth of -800RL at depth. The geology of the area is generally well recognized as mining activity within the area has been occurring for over 100 years. Remnant mining activity has been undertaken throughout 2014 in the S600 mineral resource area which has necessitated the requirement for a complete mineral resource model to use for targeting and estimation of potential recoverable Mineral Reserve.

A total of 1,530 drillholes are within the north Magdala S600 area and were extracted from the acQuire database for use in the mineral resource evaluation. These holes were intersecting the lodes at varying spacing and angles. The model consists of four ore lodes - the Central Lode, Dukes Lode, Extended Lode and the Stockworks Lode. Weakly Mineralized Volcanogenic material is also modeled.

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Given the long mining history of the Magdala orebody a full ‘depletion’ model does not exist, however, many of the mined openings are modeled. As such, reporting from the model was limited to areas that were known to still be intact and totally un-mined.

The aim of the modelling exercise of the north Magdala S600 was to improve the previously incomplete geology shapes to identify areas that warrant more detailed assessment for remnant targeting. As such a detailed QA/QC analysis was not undertaken and no reportable resource was delivered.

    14.3.6.6

MARINERS

The underground Mariners area represents the down-dip extension/fault off-set of the surface Mariners target which was modelled as part of the Big Hill and Upper Levels model in April 2014. The Mariners underground area is not a Magdala contact or quartz lode style mineralized system but part of the Scotchmans Fault Zone (SFZ) where mineralization is hosted within shear zones and the whole zone is typified by high carbon content. The prospect has been drilled over several surface and underground campaigns dating back to 1986. The first mineral resource estimate and mineral reserve was conducted in December 2009 and followed by updates in July 2012 and October 2013. The Mariners Lode has had four mineral resource model updates, the last completed in December 2013. A further 6 diamond holes were drilled into the Mariners Lode but there was no mineral resource model update for this drilling as drill results were disappointing.

At the time of the latest model update in December 2013, there were 72 diamond drillholes coded as part of the geological modelling process. The most noticeable to this model update is the addition of the Curiosity Lode. The Curiosity Lode is similar to the other four lodes defined previously (Spirit, Mariners, New and Opportunity), like these, it strikes roughly 335° and dips at 50° to the west. The Curiosity Lode sits further in the footwall to the Opportunity Lode and has been delineated from the October 2013 Footwall Low Grade Domain.

    14.3.6.7

BIG HILL AND UPPER LEVELS

The original Big Hill Mineral Resource estimate was completed in 1998, with the next estimate completed in 2012, incorporating new drilling from 2008 and 2012 RC programs, an updated geological interpretation and mineral resource estimation and a void model review. The most recent estimate (March, 2014) is a small update to the estimate presented in November 2012, and continues to utilize the same resource parameters for reporting. The underground resource is reported above a 2.0 g/t Au cut-off (see Table 14-4). The March 2014 model update was a small update to the 2012 model, with inclusion of data from the 2013 geotechnical drill information.

Upper Levels is the up dip extension of the Magdala system, which has been historically mined from underground. It contains basalt contact mineralization, Central Lode mineralization and Stockwork Lode mineralization, all typically seen in the Magdala system. Above the South Fault is the Mariners mineralization, which extends up-dip into the Big Hill surface resource.

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    14.3.6.8

AURORA B

The Aurora B Resource Model is the first Resource Model released for the Eastern Flank Area of Stawell Gold Mine. The Aurora B is the above Scotchmans Fault mineralization on the Eastern Flank, located at the South end of the Basalt dome. This model included 18 new diamond holes, which were drilled through late-2015 and early-2016 and 3 previous diamond drill holes. Drill hole spacing is in the order of 50m x 50m.

This model run included a northerly plunge component, which is consistent with that observed in the Central mineralization shoots of the western flank. The Aurora B Mineral Resource only included material from the Hampshire Lode, but there is scope for material to be added from the Waterloo zone and Wonga style zone (outboard of Hampshire Lode) upon further drilling. The Waterloo and Wonga style zones were not included in this model as currently there is not the confidence for these to be considered Inferred Resource.

A high-level independent review of the Aurora B resource modelling parameters was undertaken by Mike Stewart from QG Group in March 2106. Consideration of material included into the Inferred Category of classification was reviewed and determined appropriate to the geological understanding of the Aurora B Eastern Flank mineralization.

To increase this resource, a structural review of all available drill core (~8,000m of oriented core) is required before drill testing the down plunge position of the mineralized shoots.

  14.3.7

BLOCK MODEL DIMENSIONS

The primary consideration of the 3D model was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralized zone within the respective models, as indicated in Table 14-1 and Table 14-2. All block models are rotated to align with the strike of the area being modeled. The details of the rotations applied are shown in Table 14-4.

Block model dimensions are varied based upon the density of the available drilling data and also on the overall geometry of the mineralized structures. The chosen block size represents approximately half the best data spacing in the northing direction and a choice in the vertical and easting dimension controlled by the need to appropriately represent the volume of the wireframes.

The three mineral resource models built in 2015 were constructed as sub block models, where previous models were constructed as partials models. Sub block models are of equal quality to partials models. The reason for the change to sub block model construction is to ensure ease of model translation between software packages, and better reporting functionality for subblock models. No dip rotation has been applied to the models (also allows for greater software compatibility) and as such, the model RL height has been reduced.

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  BLOCK SIZE          

 

  E x N x RL MODEL COMP     GRADE

 

METHOD (meters) ROTATION LENGTH LODE GEOCODE  CAP Au g/t

FEDERAL

OK 2.5 x 10 x 5 320NW 1 m      

ALBION

        Extended 300 23

SOUTH

        Central 100 25

 

        Stockworks 630 35

 

      Hangingwall Volc 710 4.5

 

        Footwall Volc 720 5

BELOW

OK 2.5 x 10 x 5 325NW 1 m      

SM 250

        Central 100 16

 

        Hangingwall 400 -

 

        Magdala 500 26

 

        Flat 1 650 13

 

        Flat 3 660 19

 

        Splays 670 -

 

        Low grade volc 700 4

USF2

OK 2.5 x 10 x 5 320NW 1 m      

 

        Central 100 -

 

        Dukes 300 -

 

        Stockworks 600 -

 

        Central Volc 720 -

AURORA B

OK 5x20x10 325NW 1 m      

 

        Hampshire 100 -

 

        Waterloo 220 -

 

        WongaStyle 30 -

 

OK 2.5 x 10 x 325NW/-        

FEDERAL

  10 25 1 m      

ALBION

        Extended 300 25

 

        Central 100 30

 

        Dukes 200 -

 

        Moonlight 400 20

 

        Magdala 500 25

 

        Stockworks 630 25

 

        Hangingwall Volc 710 12

 

        Footwall Volc 720 20

 

OK   3240NW/-        

MARINERS

  3 x 10 x 10 35 1 m      

 

        Spirit Lode 110 21

 

        Mariners Lode 120 21

 

        New Lode 130 21

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  BLOCK SIZE          

 

  E x N x RL MODEL COMP     GRADE

 

METHOD (meters) ROTATION LENGTH LODE GEOCODE  CAP Au g/t

 

        Opportunity Lode 140 25

 

        Curiosity Lode 150 -

 

        LowGrade    

 

        Hangingwall 300 -

 

        LowGrade Spirit 310 2

 

        LowGrade Mariners 320 2

 

        LowGrade New 330 2

 

        LowGrade    

 

        Opportunity 340 7

 

        LowGrade Footwall 350 7

 

OK 10 x 10 x          

MAGDALA

  2.5 0/55 1 m      

S6000

        Extended 300 31

 

        Central 100 28

 

        Dukes 200 32

 

        Stockworks 600 24

 

        Central Volc 710 6

 

        Dukes Volc 720 9

 

        Extended Volc 730 7

BIG HILL

OK 5x10x5 315/0 2 m 601 Brown 601 18

 

        602 Orange 602 20

 

        603 Blue 603 -

 

        604 Green 604 -

 

        605 Purple 605 -

 

        606 Grey 606 -

 

        607 Teal Links 607 -

 

        608 Cyan 608 -

 

        610 Yellow 610 13

 

        611 Pink Stocks 611 4

 

        505 LG Volc Davis 505 4

 

        107 Iron Duke 107 12

 

        108 Allens 108 -

 

        109 Mariners 109 18

 

        111 Mariners Splay 111 15

 

        501 LG Volc Allens 501 -

 

        502 LG Volc Iron Duke 502 -

 

        503 LG Volc Mariners 503 4

 

        504 LG Volc 3 504 -

TABLE 14-4 BLOCK MODEL DIMENSIONS AND MODEL SET UP

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  14.3.8

BLOCK MODEL CODING

Block models are coded with the key structural/mineralization domains. This process is completed using the MineSight software. The details of the coding also vary by Mineral Resource area and are determined based on the number and geometry of the mineralization domains. Given the block sizes utilized by Stawell Gold Mines relative to the individual structures, multiple domain codes and percentages are stored in each block.

  14.3.9

DRILLHOLE CODING

A diamond drillhole set is coded with the key structural/mineralization domains. Importantly, this coding is checked manually by the area geologists to ensure that they match the wireframes created. Where required, some manual adjustment is made and if suspect drillhole locations are noted these drillholes may be excluded from the estimate. These details are included in the documentation for each Mineral Resource area.

  14.3.10

COMPOSITING AND STATISTIC

Compositing of the raw drilling sample data is necessary to establish a single support for the data (length) and to avoid bias when calculating statistics and undertaking any estimation of the data into 3D volumes. A number of items are considered when selecting an appropriate composite length; they include the original support of the raw sample data, the assumed selectivity (and therefore the block size) of the model and the imposed spatial dimensions of the mineralized domains.

The drillhole files are composited downhole to fixed length composite intervals. The composites are matched to honor the geological domains as coded in the drillhole files.

Composite intervals appear in Table 14-6.

The choice of composite intervals is made on a model by model basis that will best reflect the block size, data available, geometry of domains and selective mining unit.

The effect of a small number of outlier composite grades or spatially isolated composites may have an undue effect on the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the composite data in 3D visualization for both mineralized and waste domains.

A number of high cuts were identified as necessary within both mineralized and waste domains. A statistical summary of the mineralized/waste domains, Table 14-5 and their corresponding high cuts Table 14-6 are outlined below.

The data populations within the majority of mineralized domains are positively skewed with moderate variability. The variability is reduced somewhat by high cutting of gold grades in those domains with relatively high coefficients of variation. Alternatively, an outlier restriction function has been applied which allows for a localization of the high grade with a range restriction for its influence, Table 14-7.

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Within waste domains, the high cut was applied with the aim of reducing the influence of singular ‘outlier’ high grades whilst allowing any genuine anomalous areas to be represented within the estimation.

 

    Minimum Maximum Mean Gold  

 

  Number of Gold Grade Gold Grade g/t Grade g/t Co-efficient

Model

Domain Composites g/t Au            Au Au of Variation

FEDERAL

100 1404 0.01 52.49 2.56 1.332

ALBION

           

SOUTH

300 1119 0.01 85.04 3.07 1.551

 

630 1163 0.01 113.67 3.65 1.816

 

710 3063 0.01 20.60 0.26 2.936

 

720 3914 0.01 22.50 0.45 2.351

BELOW

           

SM250

100 284 0.05 45.78 3.00 1.320

 

400 63 0.45 26.80 4.64 0.946

 

500 671 0.01 169.62 3.34 2.490

 

650 170 0.01 23.00 3.59 0.905

 

660 191 0.01 48.80 3.43 1.470

 

670 44 0.61 12.44 3.73 0.690

 

700 17860 0.01 5.46 0.80 1.500

USF2

100 295 0.01 20.29 2.66 1.270

 

300 65 0.02 26.50 5.15 1.150

 

630 153 0.01 30.50 3.16 1.550

 

720 835 0.01 49.10 0.87 4.160

AURORA B

100 172 0.01 34.34 2.59 1.762

 

220 169 0.01 6.13 0.27 1.798

 

30 33 0.01 7.25 1.57 1.271

FEDERAL

300 1417 0.01 97.78 2.49 1.976

ALBION

100 5578 0.01 141.63 2.80 1.950

 

200 9 0.01 4.60 1.01 1.903

 

400 911 0.01 25.96 1.29 2.173

 

500 1488 0.01 169.62 2.59 2.517

 

710 2504 0.01 29.90 0.65 3.771

 

720 10312 0.01 80.11 0.82 6.044

MARINERS

110 287 0.01 41.44 1.76 2.333

 

120 482 0.01 44.70 2.31 1.793

 

130 822 0.01 48.71 2.44 1.606

 

140 1473 0.01 107.22 2.74 1.795

 

150 276 0.01 12.60 1.69 1.068

 

300 396 0.01 2.12 0.11 2.280

 

310 495 0.01 12.35 0.19 3.206

 

320 957 0.01 5.21 0.24 1.742

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    Minimum Maximum Mean Gold  

 

  Number of Gold Grade Gold Grade g/t Grade g/t Co-efficient

Model

Domain Composites g/t Au            Au Au of Variation

 

330 623                0.01 3.61 0.27 1.311

 

340 776                0.01 28.06 0.98 2.380

 

350 834                0.01 30.60 0.52 2.998

MAGDALA

300 2714                0.01 183.00 5.54 1.797

S6000

100 5669                0.01 119.80 4.52 1.425

 

200 689                0.01 77.60 5.49 1.334

 

600 2699                0.01 128.70 2.76 2.192

 

710 3522                0.01 95.00 0.58 4.426

 

720 3992                0.01 19.29 0.57 2.219

 

730 8261                0.01 57.90 0.60 3.285

BIG HILL

601 1493                0.01 72.60 2.13 1.507

 

602 501                0.01 31.94 2.27 1.244

 

603 237                0.01 14.83 1.60 1.348

 

604 59                0.01 9.12 1.66 0.990

 

605 36                0.01 9.85 1.32 1.265

 

606 140                0.01 6.23 0.89 1.131

 

607 74                0.01 6.71 1.16 1.087

 

608 28                0.01 5.15 1.07 1.023

 

610 318                0.01 71.20 2.82 2.072

 

611 30                0.01 64.98 3.39 3.445

 

505 6681                0.01 13.99 0.18 2.599

 

107 405                0.01 24.55 1.23 1.519

 

108 541                0.01 12.68 1.16 1.248

 

109 526                0.01 129.60 2.91 2.114

 

111 74                0.01 28.48 2.42 1.944

 

501 342                0.01 2.17 0.12 1.620

 

502 331                0.01 2.51 0.17 1.547

 

503 585                0.01 11.52 0.26 2.473

 

504 18                0.01 1.31 0.24 1.452

TABLE 14-5 STATISTICAL SUMMARY BY AREA

      Minimum Maximum    Mean Co-
      Gold Gold Gold efficient
    Number of Grade g/t Grade g/t Grade g/t of
Model Domain Composites Au Au Au Variation
 FEDERAL 100 1404 0.01 20 2.50 1.143
 ALBION            
 SOUTH 300 1119 0.01 22 2.97 1.265
  630 1163 0.01 35 3.47 1.394
  710 6063 0.01 6 0.25 2.312
  720 3914 0.01 6 0.43 1.962
 BELOW SM 100 284 0.01 16 2.88 1.059

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      Minimum Maximum    Mean Co-
      Gold Gold Gold efficient
    Number of Grade g/t Grade g/t Grade g/t of
Model Domain Composites Au Au Au Variation
250            
   500 671 0.01 26 3.02 1.460
   650 170 0.01 13 3.51 0.828
   660 191 0.01 19 3.24 1.176
   700 - 0.01 4 0.80 1.500
FEDERAL  300 1417 0.01 25 2.37 1.570
ALBION  100 5578 0.01 30 2.68 1.490
   400 911 0.01 20 1.27 2.100
   500 1488 0.01 25 2.40 1.690
MARINERS  110 287 0.01 21 1.64 1.937
   120 482 0.01 21 2.19 1.499
   130 822 0.01 21 2.36 1.423
   140 1473 0.01 25 2.64 1.458
   310 495 0.01 2 0.17 1.529
   320 957 0.01 2 0.23 1.515
   330 623 0.01 2 0.27 1.227
   340 776 0.01 7 0.85 1.752
   350 834 0.01 7 0.47 2.162
MAGDALA  300 2714 0.01 31 5.10 1.350
S6000  100 5669 0.01 28 4.34 1.125
   200 689 0.01 32 5.36 1.220
   600 2699 0.01 24 2.53 1.512
   710 3522 0.01 6 0.48 1.920
   720 3992 0.01 9 0.56 2.035
   730 8261 0.01 7 0.51 2.187
BIG HILL  601 1493 0.01 18 2.05 1.100
   602 501 0.01 20 2.24 1.165
   610 318 0.01 13 2.38 1.046
   611 30 0.01 4 1.36 0.898
   505 6681 0.01 4 0.17 2.206
   107 405 0.01 12 1.20 1.285
   109 526 0.01 18 2.69 1.018
   111 74 0.01 15 2.15 1.609
   503 585 0.01 4 0.25 1.883

TABLE 14-6 STATISTICAL SUMMARY FOR HIGH CUT COMPOSITES, GOLD G/T AU

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  14.3.11

GEOSTATISTICAL PARAMETERS

For all of the Mineral Resource areas that have been modelled in 3D, gold grades (Au g/t) have been estimated by Ordinary Kriging. Over time, this methodology, when coupled with detailed and robust geological models, provided reliable estimates of in situ gold grade.

The key geostatistical parameters are modelled for each project separately (see Table 14-8). Variography studies have been conducted using MineSight MSDA analysis software. Individual variogram studies are conducted for each domain and modelled separately (see Table 14-8).

Key variogram parameters, nugget and sill, and variogram ranges are modelled on untransformed data. Whilst the estimated values for the nugget and ranges vary for each modelled area, they are generally relatively consistent.

The variograms studies have been performed to optimize the Kriging neighborhood as per the methodologies presented in Vann et al (2003). This enables quantitative evaluation of the results of the Kriging to be performed and, as well as enabling the search neighborhoods to be optimized, provides numerical outputs from the Kriging runs (slope of regression of the estimate). The slope of regression is used in part to aid in classification of the Mineral Resource.

Search rotations were applied directly from the variography search orientation. Search distance was given the same ratio as variography range but made much larger than the variography range to ensure that it does not limit the Kriging, and only influences the search directions (Figure 14-5 and Figure 14-6).

FIGURE 14-5 LONGSECTION OF BIG HILL DOMAIN 601 SHOWING COMPOSITES WHICH HAVE A PLUNGE

THE VARIOGRAPHY IS ORIENTATED DOWN THIS PLUNGE (RED ELLIPSE) IN AN ANISOTROPIC ORIENTATION. SEARCH IS GIVEN THE SAME ORIENTATION WITH A LARGER RATIO (GREEN ELLIPSOID).

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FIGURE 14-6 PLAN OF BIG HILL DOMAIN BIG 602 SHOWING COMPOSITES WHICH HAVE NO PREFERRED PLUNGE

THE VARIOGRAPHY IS ORIENTATED IN THE PLAN OF THE DOMAIN (RED ELLIPSE) IN A FLAT ISOTROPIC ORIENTATION. SEARCH IS GIVEN THE SAME ORIENTATION WIT H A LARGER RATIO (GREEN ELLIPSOID)

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AREA

DOMAIN x axis - 1st search y axis - 1st search z axis - 1st search x axis - 2nd search y axis - 2nd search z axis - 2nd search  min. samples 1st search  max. samples 1st search Search Type  max samples per Quadrant max. samples per hole min. samples 2nd search  max. samples 2nd search Search Type max samples per Quadrant max. samples per hole Outlier cut - off Max. 3D Search for Outliers

 

search distance sample number definition outliers

FEDERAL ALBION SOUTH

 CENTRAL

101 105 113 20 - - - 1 24 -  - 6 - - - - - - -

 CENTRAL

102 75 62.5 12.5 - - - 1 24 -  - 6 - - - - - - -

 CENTRAL

103 105 113 20 - - - 1 24 -  - 6 - - - - - - -

 CENTRAL

104 105 113 20 - - - 1 24 -  - 6 - - - - - - -

 EXTENDED

301 100 75 20 - - - 1 24 -  - 6 - - - - - - -

 EXTENDED

302 75 62.5 17.5 - - - 1 24 -  - 6 - - - - - - -

 FED ALB STK

630 62.5 62.5 12.5 - - - 1 20 -  - 6 - - - - - - -

 HW VOLC

710 100 75 25 - - - 1 16 -  - 4 - - - - - - -

 FW VOLC

720 100 75 25 - - - 1 16 -  - 4 - - - - - - -

SM 250

 

 CENTRAL

100 75 50 25 - - - 1 28 -  - 5 - - - - - - -

 HANGING WALL

400 75 50 25 - - - 1 16 -  - 4 - - - - - 12 10

 MAGDALA

500 75 60 25 - - - 1 20 -  - 6 - - - - - - -

 FLAT 1

650 75 25 13 - - - 1 20 -  - 4 - - - - - - -

 FLAT 3

660 75 25 13 - - - 1 20 -  - 4 - - - - - 13 15

USF2

 

 CENTRAL

101 42 42 15 - - - 2 24 -  - 4 - - - - - - -

 CENTRAL

102 42 42 15 - - - 2 24 -  - 4 - - - - - -

 DUKES

301 75 62 10 - - - 2 24 -  - 3 - - - - - 20 15

 STOCKWORKS

630 42 42 15 - - - 2 24 -  - 4 - - - - - - -

156LG VOLC1

720 42 42 15 - - - 2 24 -  - 4 - - - - - 3 10

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                       AREA

DOMAIN x axis - 1st search y axis - 1st search z axis - 1st search x axis - 2nd search y axis - 2nd search z axis - 2nd search min. samples 1st search max. samples 1st search Search Type max samples per Quadrant max. samples per hole min. samples 2nd search  max. samples 2nd search Search Type max samples per Quadrant max. samples per hole off Outlier cut - Max. 3D Search for Outliers

 LG VOLC

720 42 42 15 - - - 2 32 - - 4 -  - - - - - -

 AURORA B

 

 HAMPSHIRE

100 125 213 30 - - - 1 24 - - 4 -  - - - - 20 20

 WATERLOO

220 175 175 25 - - - 1 24 - - 5 -  - - - - - -   -

 WONGASTYLE

30 175 175 15 - - - 1 24 - - 5 -  - - - - - - -

FEDERAL ALBION

 CENTRAL

101 60 40 40 320 0 -25 1 24 Quadrant 6 - 8  48 Quadrant 12 8 - -

 CENTRAL

102 60 40 40 326 26 -31 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 CENTRAL

103 60 40 40 330 0 -45 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - - -

 DUKES

200 60 48 48 330 0 -15 1 24 Quadrant 6 - - - - - - - -

 EXTENDED

301 80 60 60 325 0 -20 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 EXTENDED

302 80 60 60 331 26 -31 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 MOONLIGHT

401 60 48 48 325 0 -15 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 MOONLIGHT

402 60 48 48 329 27 -26 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 MAGDALA

501 60 48 48 340 0 -30 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 MAGDALA

502 60 48 48 334 27 -26 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 MAGDALA

503 60 48 48 315 0 -20 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 FED ALB STK

630 64 72 40 310 0 -20 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 HW VOLC

710 50 70 30 341 26 -31 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

 VOLC

720 40 30 30 320 0 -30 1 24 Quadrant 6 - 8 48 Quadrant 12 8 - -

MARINERS

 SPIRIT

110 40 40 15 - - -    4 24 - - 4 - - - - - - -

 MARINERS

120 40 40 15 - - -    4 24 - - 4 - - - - - - -

 NEW

130 40 40 15 - - - 4 24 - - 4 - - - - - - -

OPPORTUNITY

140 40 40 15 - - - 4 20 - - 4 - - - - - 18   10

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                     AREA DOMAIN x axis - 1st search y axis - 1st search z axis - 1st search x axis - 2nd search y axis - 2nd search z axis - 2nd search min. samples 1st search max. samples 1st search Search Type max samples per Quadrant max. samples per hole min. samples 2nd search max. samples 2nd search Search Type max samples per Quadrant max. samples per hole Outlier cut - off Max. 3D Search for Outliers
CURIOSITY 150 40 40 15 - - -   4 32 -  - 4 - - - -    - - -
LG HW 300 40 40 18 - - - 4  32 - - 4 - - - - - - -
                                       
LG SPIRIT 310 40 40 18 - - - 4 28 - - 4 - - - - - - -
LG MARINERS 320 40 40 18 - - - 4 24 - - 4 - - - - - - -
                                       
LG NEW 330 40 40 18 - - - 4 24 - - 4 - - - - - - -
LG OPPORTUNITY 340 40 40 18 - - - 4 24 - - 4 - - - - - 5 10
LG FW 350 40 40 18 - - - 6 24 - - 4 - - - - - 3 10
MAGDALA S6000
CENTRAL 100
F1
63 75 20 - - - 4 32 - - 5 - - - - - - -
CENTRAL 100
F2
63 75 20 - - - 4 32 - - 5 - - - - - - -
DUKES 200
F1
75 90 18 - - - 4 26 - - 5 - - - - - - -
  200                                    
DUKES F2 75 90 18 - - - 4 25 - - 5 - - - - - - -
EXTENDED 300 75 90 18 - - - 4 32 - - 5 - - - - -  25 20
STOCKWORK 600 120 100 18 - - - 4 32 - - 5 - - - - - - -
VOLC CENT 710 30 38 10 - - - 4 32 - - 5 - - - - -  5 7
VOLC DUKES 720 40 40 18 - - - 32 - - 5 - - - - - - -
VOLC EXTEND 730 40 40 20 - - - 24 - - 5 - - - - - 5 7
BIG HILL
601 BROWN 601 80 60 20 - - - 28 - - 5 - - - - - - -
602 ORANGE 602 75 75 25 - - - 5 28 - - 5 - - - - - - -
603 BLUE 603 75 45 20 - - - 5 28 - - 5 - - - - - - -
604 GREEN 604 75 75 25 - - - 28 - -  5 - - - - - - -
605 PURPLE 605 75 75 25 - - -   5 28 - - 5 - - - -   - - -

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AREA DOMAIN x axis -
1st search
y axis -
1st search 
z axis -
1st search
 
x axis -
2nd search 
 
y axis -
2nd search 
 
z axis -
2nd search 
 
min. samples
1st search 
 
max. samples
1st search 
 
Search
Type 
 
max samples
per Quadrant 
 
max. samples
per hole 
 
min. samples
2nd search 
 
max. samples
2nd search 
 
Search
Type  
 
max samples
per Quadrant
 
max. samples
per hole 
 
Outlier
cut - off 
 
Max. 3D Search
for Outliers
  
606 GREY 606 75 75 25  -  -  -  5  28 - - 5 - - - - - - -
607 TEAL LINKS 607 85 85 25  -  -  -  5  28 - - 5 - - - - - - -
608 CYAN 608 50 50 25  -  -  -  5  28 - - 5 - - - - - - -
610 YELLOW 610 75 75 25  -  -  -  5  22 - - 5 - - - - - - -
611 PINK STOCKWORKS 611 75 75 25 - - - 5 28 - - 5 - - - - - - -
505 LG VOLC DAVIS 505 75 75 25  -  -  -  5 24 - - 5 - - - - - - -
107 IRON DUKE 107 75 75 25  -  -  - 5 30 - - 5 - - - - - 2 10
108 ALLENS 108 75 75 25  -  -  -  5  28 - - 5 - - - - - - -
109 MARINERS 1091 75 75 25  -  -  -  5  28 - - 5 - - - - - - -
109 MARINERS 1092 75 40 15  -  -  -  5  28 - - 5 - - - - - - -
109 MARINERS 1093 75 25 15  -  -  -  5  28 - - 5 - - - - - - -
109 MARINERS 1094 75 25 15  -  -  -  5  28 - - 5 - - - - - - -
110 MARINERS 110 75 75 25  -  - -  5 28 - - 5 - - - - - - -
111 MARINERS SPLAY 111 75 75 25  - -   -  5  12 - - 5 - - - - - - -
11 SM FAULT 11 160 40 10  -  -  -  5  24 - - 5 - - - - - 15 10
501 LG VOLC ALLENS 501 75 75 25  -  -  -  5  24 - - 5 - - - - - - -
502 LG VOLC IRON DUKE 502 75 75 25 - - - 5 24 - - 5 - - - - - 2 15
503 LG VOLC MARINERS 503 75 75 25 - - - 5 24 - - 5 - - - - - - -
504 LG VOLC 3 504 75 75 25  -  -  -  5  24 -  - 5 - - - - - - -

TABLE 14-7 STAWELL GOLD MINES GEOSTATISTICAL SEARCH PARAMETERS

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AREA

DOMAIN axis 1    axis 2 axis 3 Nugget    1st Structure  pri. axis (x)  sec. axis (y) ter. axis (z) variance 2nd Structure    pri. axis (x)  sec. axis (y)  ter. axis (z)  variance  Total Variance   Nugget

 

  rotation                          

 FEDERAL ALBION SOUTH

 CENTRAL

100 320 0  60 3.9 sph 30 25 5 6.1 - - - - - 10.00 39%

 CENTRAL

101 320 0  50 3.9 sph 30 25 5 6.1 - -  - - - 10.00 39%

 CENTRAL

102 327 0  70 3.9 sph 30 25 5 6.1 - -  - - - 10.00 39%

 CENTRAL

103 320 0  53 3.9 sph 30 25 5 6.1 - -  - - - 10.00 39%

 EXTENDED

300 310 0  75 3.8 sph 50 30 7 7 - -  - - - 10.80 35%

 EXTENDED

302 315 0  62 3.8 sph 50 30 7 7 - -  - - - 10.80 35%

 EXTENDED

303 317 0  55 3.8 sph 50 30 7 7 - -  - - - 10.80 35%

 STOCKWORKS

630 310 0  80 7.5 sph 25 25 5 9.5 - -  - - - 17.00 44%

 HW VOLC

710 318 0  57 0.33 sph 40 30 10 0.49 - -  - - - 0.82 40%

 FW VOLC

720 318 0  57 0.33 sph 40 30 10 0.49 - -  - - - 0.82 40%

 SM 250

 CENTRAL

100 322 -15  39 5.7 sph 10 5 3 9.3 - -  - - - 21.00 27%

 HANGINGWALL

400 320 -12  50 5.7 sph 10 5 3 9.3 - -  - - - 21.00 27%

 MAGDALA

500 338 0  50 4.6 sph 18 12 3 11.7 - -  - - - 23.40 20%

 FLAT 1

650 310 -23  11 5.7 sph 5 3 3 9.3 - - - - - 21.00 27%

 FLAT 3

660 324 -11  22 5.7 sph 5 3 3 9.3 - - - - - 21.00 27%

 LG VOLC

700 335 0  39 5.7 sph 10 5 3 9.3 - - - - -   27%

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AREA

DOMAIN axis 1 axis 2 axis 3 Nugget    1st Structure  pri. axis (x)  sec. axis (y) ter. axis (z) variance    2nd Structure  pri. axis (x)  sec. axis (y) ter. axis (z) variance Total Variance Nugget
                              21.00  

 USF2

 CENTRAL

101 315 0 90 4 sph 15 15 6 6 - - - - - 10.00 40%

 CENTRAL

102 327 0 80 4 sph 15 15 6 6 - - - - - 10.00 40%

 DUKES

300 329 -20 90 3 sph 30 25 4 12 - - - - - 15.00 20%

 STOCKWORKS

630 312 -16 90 3 sph 15 15 6 10 - - - - - 13.00 23%

 LG VOLC1

721 317 0 90 4 sph 15 15 6 6 - - - - - 10.00 40%

 LG VOLC

722 338 -36 82 4 sph 15 15 6 6 - - - - - 10.00 40%

 AURORA B

 HAMPSHIRE

110 325 -44 -88 4 sph 50 85 6 13 - - - - - 17.00 24%

 HAMPSHIRE

120 341 -35 -71 4 sph 50 85 6 13 - - - - - 17.00 24%

 HAMPSHIRE

130 341 -35 -71 4 sph 50 85 6 13 - - - - - 17.00 24%

 WATERLOO

220 317 0 90 0.04 sph 70 70 10 0.1 - - - - - 0.14 29%

 WONGASTYLE

30 317 0 -57 0.04 sph 70 70 6 0.1 - - - - - 0.14 29%

 FEDERAL ALBION

 CENTRAL

101 320 0 -25 5.7 sph 10 5 3 9.3 sph  30  20  10 6 21.00 27%

 CENTRAL

102 326 26 -31 3.3 sph 10 5 3 4.4 sph  30  20  10 4 11.70 28%

 CENTRAL

103 330 0 -45 4 sph 10 5 2.5 6.4 sph  30  20    8 4 14.40 28%

 DUKES

200 330 0 -15 0.48 sph 18 12 3 1 sph  30  24    7 0.42   25%

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AREA

DOMAIN axis 1 axis 2 axis 3    Nugget    1st Structure pri. axis (x)   sec. axis (y) ter. axis (z) variance    2nd Structure  pri. axis (x)  sec. axis (y) ter. axis (z) variance  Total Variance Nugget
                              1.90  

 EXTENDED

301 325 0 -20 3.6 sph 20 15 3 8.9 sph  40  30 5  5.4 17.90 20%

 EXTENDED

302 331 26 -31 3.1 sph 20 15 3 6.6 sph  40  30 5  5.8 15.50 20%

 MOONLIGHT

401 325 0 -15 2.4 sph 18 12 3 5 sph  30  24 7 2.25 9.65 25%

 MOONLIGHT

402 329 27 -26 3.1 sph 18 12 3 6.4 sph  30  24 7  2.9 12.40 25%

 MAGDALA

501 340 0 -30 4.6 sph 18 12 3 11.7 sph  30  24 7  7.1 23.40 20%

 MAGDALA

502 334 27 -26 1.3 sph 18 12 3 3.3 sph  30  24 7      2 6.60 20%

 MAGDALA

503 315 0 -20 4.4 sph 18 12 3 10.9 sph  30  24 7  6.5 21.80 20%

 STOCKWORKS

630 310 0 -20 3.8 sph 16 18 4 7.4 sph  32  36 10 4.85 16.05 24%

 HW VOLC

710 341 26 -31 0.2 sph 10 25 3 0.7 sph  25  35 6  0.4 1.30 15%

 VOLC

720 320 0 -30 0.45 sph 10 7 3 1.5 sph  20  15 12  0.7 2.65 17%

 MARINERS

 SPIRIT

110 325 0 50 3.9 sph 20 20 7 6.8 - - - - - 10.70 36%

 MARINERS

120 330 0 52 3.9 sph 20 20 7 6.8 - - - - - 10.70 36%

 NEW

130 333 0 52 3.9 sph 20 20 7 6.8 - - - - - 10.70 36%

 OPPORTUNITY

140 334 0 55 3.9 sph 20 20 7 6.8 - - - - - 10.70 36%

 CURIOSITY

150 332 0 55 3.9 sph 20 20 7 6.8 - - - - - 10.70 36%

 LG HW

300 333 0 60 0.045 sph 20 20 9 0.07 - - - - - 0.12 39%

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AREA

DOMAIN axis 1 axis 2 axis 3    Nugget    1st Structure pri. axis (x)   sec. axis (y) ter. axis (z) variance    2nd Structure  pri. axis (x)  sec. axis (y) ter. axis (z) variance Total Variance Nugget

 LG SPIRIT

310 333 0 52 0.045 sph 20 20 9 0.07 - - - - - 0.12 39%

 LG MARINERS

320 333 0 50 0.045 sph 20 20 9 0.07 - - - - - 0.12 39%

 LG NEW

330 335 0 54 0.045 sph 20 20 9 0.07 - - - - - 0.12 39%

 LG OPPORTUNITY

340 335 0 55 0.045 sph 20 20 9 0.07 - - - - - 0.12 39%

 LG FW

350 335 0 55 0.045 sph 20 20 9 0.07 - - - - - 0.12 39%

 MAGDALA S6000

 CENTRAL

100 F1 211 -41 -32 9.5 sph 7 10 4.2 10 sph 25 30 8.2 4 23.50 40%

 CENTRAL

100 F2 326 0 70 9.5 sph 7 10 4.2 10 sph 25 30 8.2 4 23.50 40%

 DUKES

200 F1 330 0 50 14 sph 8 12 3 22 sph 30 40 7.5 9 45.00 31%

 DUKES

200 F2 324 0 76 14 sph 8 12 3 22 sph 30 40 7.5 9 45.00 31%

 EXTENDED

300 345 0 55 14 sph 8 8   24 sph  30  40 8.5 12 50.00 28%

 STOCKWORKS

600 348 0 46 6.6 sph 5 8 4.5 5.4 sph  50  40 5.7 4.1 16.10 41%

 VOLC CENT

710 328 -34 35 0.33 sph 12 15 4 0.37 - - - - - 0.70 47%

 VOLC DUKES

720 330 0 50 0.265 sph 5 5 2 0.52 sph  12  12 5.4 0.44 1.23 22%

 VOLC EXTEND

730 300 0 55 0.53 sph 15 15 8 0.53 - - - - - 1.06 50%

 BIG HILL

 601 BROWN

601 30 50 20 1.8 sph 15 10 3 1.5 sph  40  30  20 1.5 4.80 38%

 602 ORANGE

602 330 0 41 2.5 sph 12 12 8 2.8 sph  30  30  20 1 6.30 40%

 603 BLUE

603 330 34 53 1.7 sph 10 10 2 0.9 sph  50  30  15 1.9   38%

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AREA

 DOMAIN axis 1 axis 2 axis 3  Nugget    1st Structure  pri. axis (x)  sec. axis (y) ter. axis (z)  variance    2nd Structure pri. axis (x)  sec. axis (y) ter. axis (z)  variance Total Variance Nugget
                              4.50  

 604 GREEN

604 326 0 44 1.8 sph 10 10 3 1.5 sph  30  30 20 1.5 4.80 38%

 605 PURPLE

605 317 0 36 1.8 sph 10 10 3 1.5 sph  30  30 20 1.5 4.80 38%

 606 GREY

606 291 0 46 1.8 sph 10 10 3 1.5 sph  30  30 20 1.5 4.80 38%

 607 TEAL LINKS

607 320 0 25 1.8 sph 10 10 3 1.5 sph  30  30 20 1.5 4.80 38%

 608 CYAN

608 310 0 67 1.8 sph 10 10 3 1.5 sph  30  30 20 1.5 4.80 38%

 610 YELLOW

610 326 0 46 0.38 sph 10 10 3 0.32 sph  20  20 8 0.3 1.00 38%

 611 PINK STOCKWORKS

611 325 0 45 1.8 sph 10 10 3 1.5 sph 30 30 20 1.5 4.80 38%

 505 LG VOLC DAVIS

505 315 0 52 0.015 sph 5 5 5 0.012 sph  30  30 20 0.015 0.04 36%

 107 IRON DUKE

107 314 -18 37 1.6 sph 8 8 4 0.6 sph  20  20 15 0.5 2.70 59%

 108 ALLENS

108 312 0 83 1.2 sph 10 10 6 1 sph  30  30 20 0.7 2.90 41%

 109 MARINERS

1091 325 0 48 3 sph 10 10 3 2.5 sph  45  45 15 1.9 7.40 41%

 109 MARINERS

1092 327 -26 37 3 sph 15 8 3 2.5 sph  45  30 15 1.9 7.40 41%

 109 MARINERS

1093 320 -22 -21 3 sph 12 8 3 2.5 sph  45  15 10 1.9 7.40 41%

 109 MARINERS

1094 320 -21 54 3 sph 8 8 3 2.5 sph  45  15 10 1.9 7.40 41%

 110 MARINERS

110 326 0 71 3 sph 8 8 3 2.5 sph  50  50 25 1.9 7.40 41%

 111 MARINERS SPLAY

111 365 0 70 0.2 sph 8 8 3 0.5 sph  50  50 25 0.3 1.00 20%

 11 SM FAULT

11 310 -20 -19 0.5 sph 30 20 10 0.7 - - - -    - 1.20 42%

 501 LG VOLC ALLENS

501 312 0 83 1.2 sph 10 10 6 1 sph  30  30 20 0.7   41%

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AREA

DOMAIN axis 1 axis 2 axis 3 Nugget    1st Structure pri. axis (x)   sec. axis (y) ter. axis (z) variance    2nd Structure  pri. axis (x)  sec. axis (y) ter. axis (z) variance Total Variance Nugget
                              2.90  

 502 LG VOLC IRON DUKE

502 320 0 40 1.6 sph 8 8 4 0.6 sph 20 20 15 0.5 2.70 59%

 503 LG VOLC MARINERS

503 346 0 46 3 sph 8 8 3 2.5 sph 50 50 25 1.9 7.40 41%

 504 LG VOLC 3

319 0 46 3 sph 8 8 3 2.5 sph 50 50 25 1.9 7.40 41%

TABLE 14-8 STAWELL GOLD MINES GEOSTATISTICAL VARIOGRAM PARAMETERS

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  14.3.12

BULK DENSITY

As per the Stawell Gold Mines core processing flow chart (see all whole core sample intervals that are sent for assay are bagged and weighed prior to leaving site. The weights of the core are stored in the drillhole database and an apparent dry bulk density for each interval is calculated based upon a theoretical volume for the specific core diameter for each sample interval.

Whilst there are potential limitations in this process given that the volume for each core is based upon the theoretical drill core diameter and the measured length of each interval, it can be demonstrated by reconciliation of stope volumes as surveyed by cavity monitoring systems and tonnage measured by trucked ore that the estimated density applied to specific areas is a reliable estimator of in situ density. A 30-year history of processing and reconciliation supports the density values that are currently being used in the estimates at Stawell Gold Mines.

This apparent density measurement includes some residual moisture which by measurement has been estimated at approximately 2% or less and is generally not accounted for in the estimates.

Table 14-9 shows the bulk density values that have been applied by Mineral Resource area. Specific details of the data and analysis and assumptions used to derive these values are given in each of the area specific reports.

 

     

MODEL

LODE GEOCODE ASSIGNED SG FOR MODEL

 

    FRESH TRANS OXIDE

FEDERAL

Central 100 2.7 - -

ALBION SOUTH

Extended 300 2.8 - -

 

Stockworks 630 2.75 - -

 

Hangingwall Volc 710 2.75 - -

 

Footwall Volc 720 2.75 - -

 

         

BELOW SM 250

Central 100 2.75 - -

 

Hangingwall 400 2.75 - -

 

Magdala 500 2.85 - -

 

Flat 1 650 2.75 - -

 

Flat 3 660 2.75 - -

 

Splays 670 2.75 - -

 

LowGrade Volc 700 2.75 - -

 

         

USF2

Central 100 2.75 - -

 

Dukes 300 2.95 - -

 

Volcanogenics 500 2.75 - -

 

Stockworks 720 2.75 - -

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MODEL

LODE GEOCODE ASSIGNED SG FOR MODEL

AURORA B

Hampshire 100 2.85 - -

 

Waterloo 220 2.89 - -

 

WongaStyle 30 2.77 - -

 

         

FEDERAL

         

ALBION

Extended 300 2.8 - -

 

Central 100 2.75 - -

 

Dukes 200 2.75 - -

 

Moonlight 400 2.85 - -

 

Magdala 500 2.85 - -

 

Hangingwall Volc 710 2.75 - -

 

Footwall Volc 720 2.75 - -

 

         

MARINERS

Spirit Lode 110 2.9 - -

 

Mariners Lode 120 2.8 - -

 

New Lode 130 2.8 - -

 

Opportunity Lode 140 2.9 - -

 

Curiosity Lode 150 2.75 - -

 

LowGrade        

 

Hangingwall 300 2.8 - -

 

LowGrade Spirit 310 2.8 - -

 

LowGrade Mariners 320 2.8 - -

 

LowGrade New 330 2.8 - -

 

LowGrade        

 

Opportunity 340 2.8 - -

 

LowGrade Footwall 350 2.8 - -

 

Breccia 420 2.7 - -

 

         

MAGDALA

         

S6000

Extended 300 2.9 - -

 

Central 100 2.75 - -

 

Dukes 200 2.9 - -

 

Stockworks 600 2.8 - -

 

Central Volc 710 2.8 - -

 

Dukes Volc 720 2.8 - -

 

Extended Volc 730 2.8 - -

 

         

BIG HILL

601 Brown 601 2.7 2.3 2.1

 

602 Orange 602 2.7 2.3 2.1

 

603 Blue 603 2.7 2.3 2.1

 

604 Green 604 2.7 2.3 2.1

 

605 Purple 605 2.7 2.3 2.1

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MODEL

LODE GEOCODE ASSIGNED SG FOR MODEL

 

606 Grey 606 2.7 2.3 2.1

 

607 Teal Links 607 2.7 2.3 2.1

 

608 Cyan 608 2.7 2.3 2.1

 

610 Yellow 610 2.7 2.3 2.1

 

611 Pink Stocks 611 2.7 2.3 2.1

 

505 LG Volc Davis 505 2.85 2.3 2

 

107 Iron Duke 107 2.85 2.3 2.15

 

108 Allens 108 2.85 2.3 2.15

 

109 Mariners 109 2.85 2.5 2.3

 

111 Mariners Splay 111 2.85 2.5 2.3

 

501 LG Volc Allens 501 2.85 2.3 2.1

 

502 LG Volc Iron Duke 502 2.85 2.3 2.1

 

503 LG Volc Mariners 503 2.85 2.3 2.1

 

504 LG Volc 3 504 2.85 2.3 2.1

 

         

ALL

BASALT 400 2.8 2.3 2.1

 

MINE SCHIST 1 2.8 2.3 2.1

 

PORPHYRY 3 2.7 2.65 2.45

TABLE 14-9 COMPILATION OF DENSITY G/CM3,APPLIED BY RESOURCE MODEL AREA

  14.3.13

MODEL VALIDATION

Model validation has been completed using several different methods. The first is a visual assessment of the output Kriging statistics and by visualization of the actual samples chosen in the search. Estimation parameters were then adjusted to ensure the right balance of estimation is achieved for each individual domain to match the geology.

The coding end estimation of each domain was reviewed by:

  Visualization of blocks in 3D;
     
  Visualization of grade estimates, particularly with respect to local reproduction of grades and smoothing; and
     
  Statistical reproduction- comparison of output block grade with input composite statistics.

Statistical comparisons of input data and block model outcomes for the mineralized domains are shown in Table 14-10.

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STAWELL GOLD MINE MODEL VALIDATION
                                   Mean Grade (g/t Au)

 

  Undeclust   Model  

 

  1.0m Declust. Average Variance to

 

Resource comp all 1.0m Grade g/t 1m Comp

DOMAIN

Category holes comp. Au Au%

FEDERAL ALBION SOUTH

CENTRAL

100 All 2.50 2.47 2.61 5.7%

CENTRAL

100 2.63 2.63 2.78 5.7%

CENTRAL

102 2.44 2.41 2.47 2.5%

CENTRAL

103 2.17 2.18 2.34 7.3%

EXTENDED

300 2.97 2.95 2.85 -3.4%

FED ALB STK

630 3.47 3.46 3.35 -3.2%

HW VOLC

710 0.25 0.25 0.32 28.0%

FW VOLC

720 0.43 0.43 0.25 -41.9%

SM 250

CENTRAL

100 2.88 2.59 2.57 -0.8%

HANGING WALL

400 4.64 4.63 3.61 -22.0%

FEDERAL CONTACT

501 2.44 2.75 - -

MAGDALA

500 3.02 3.05 3.08 1.0%

FLAT 1

650 3.51 3.55 3.66 3.0%

FLAT 3

660 3.24 3.27 3.06 -6.4%

LG VOLC

700 0.80 0.82 0.48 -40.8%

USF2

CENTRAL

100 2.66 2.66 2.92 8.8%

DUKES

300 5.15 5.25 4.46 -17.8%

STOCKWORKS

630 3.16 3.16 2.7894 -13.4%

LG VOLC

720 0.87 0.88 0.96 8.4%

AURORA B

HAMPSHIRE

100 2.59 2.58 2.47 -4.2%

WATERLOO

220 0.23 0.23 0.22 -4.7%

WONGA STYLE

30 1.47 1.40 1.42 1.5%

FEDERAL ALBION

CENTRAL

100 2.80 2.69 2.35 -12.7%

DUKES

200 1.01 1.06 0.91 -14.3%

EXTENDED

300 2.49 2.38 2.68 12.6%

MOONLIGHT

400 1.29 1.27 1.46 15.3%

MAGDALA

500 2.59 2.40 2.18 -9.1%

HW VOLC

710 0.65 0.60 0.65 8.0%

VOLC

720 0.82 0.76 0.76 0.5%

MARINERS

         

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  Undeclust   Model  

 

  1.0m Declust. Average Variance to

 

Resource comp all 1.0m Grade g/t 1m Comp

DOMAIN

Category holes comp. Au Au%

SPIRIT

110 1.77 1.65 2.51 41.7%

MARINERS

120 2.25 2.17 3.05 35.9%

NEW

130 2.44 2.37 2.39 -2.2%

OPPORTUNITY

140 2.74 2.64 2.33 -14.8%

CURIOSITY

150 1.67 1.67 1.89 13.3%

 

         

LG HW

300 0.11 0.11 0.12 7.1%

LG SPIRIT

310 0.19 0.17 0.19 -0.1%

 

         

LG MARINERS

320 0.24 0.23 0.24 1.1%

LG NEW

330 0.27 0.27 0.24 -9.9%

LG OPPORTUNITY

340 0.96 0.83 0.40 -58.5%

LG FW

350 0.52 0.47 0.48 -7.6%

MAGDALA S6000

CENTRAL

100 4.52 4.30 3.73 -13.1%

DUKES

200 5.49 5.35 5.64 5.4%

EXTENDED

300 5.54 5.16 4.37 -15.4%

STOCKWORK

600 2.76 2.59 2.67 3.2%

VOLC CENT

710 0.58 0.48 0.46 -2.8%

VOLC DUKES

720 0.57 0.56 0.52 -6.2%

VOLC EXTEND

730 0.60 0.60 0.60 0.2%

BIG HILL

601 BROWN

601 2.13 2.04 1.92 -5.8%

602 ORANGE

602 2.27 2.24 2.10 -6.6%

603 BLUE

603 1.60 1.61 1.56 -3.3%

604 GREEN

604 1.66 1.69 1.77 4.2%

605 PURPLE

605 1.32 1.28 1.15 -11.3%

606 GREY

606 0.89 0.90 0.90 0.4%

607 TEAL LINKS

607 1.16 1.20 1.09 -9.5%

608 CYAN

608 1.07 1.08 1.09 1.2%

610 YELLOW

610 2.82 2.36 2.48 5.0%

611 PINK STOCKWORKS

611 3.39 1.40 1.43 1.9%

505 LG VOLC DAVIS

505 0.18 0.17 0.16 -3.2%

107 IRON DUKE

107 1.23 1.20 1.14 -5.1%

108 ALLENS

108 1.16 1.18 1.19 0.8%

109 MARINERS

109 2.91 2.72 2.63 -3.6%

110 MARINERS

110 2.00 2.01 2.10 4.5%

111 MARINERS SPLAY

111 2.42 2.29 1.58 -44.7%

11 SM FAULT

11 0.33 0.33 0.28 -18.0%

501 LG VOLC ALLENS

501 0.12 0.12 0.15 16.4%

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  Undeclust   Model  

 

  1.0m Declust. Average Variance to

 

Resource comp all 1.0m Grade g/t 1m Comp

DOMAIN

Category holes comp. Au Au%

502 LG VOLC IRON DUKE

502 0.17 0.17 0.15 -16.6%

503 LG VOLC MARINERS

503 0.26 0.25 0.24 -6.3%

504 LG VOLC 3

504 0.24 0.17 0.19 11.9%

TABLE 14-10 MINERALIZED DOMAINS MODEL VALIDATION

The mineralized domain comparisons display some variation between input and outcome average grades when the total domain is reported. The variations in grade occur in areas that generally are poorly informed and in portions of domains that represent the margins of the modelling area.

  14.3.14

MINERAL RESOURCE CLASSIFICATION

The classification of Indicated and Inferred Mineral Resource material is based on geological confidence, slope of regression analysis and model validation results. The practice adopted at Stawell Gold Mines uses general guidelines for classification that utilize the following information;

 

Drilling density

 

Drillhole spacing and sample locations

 

Stage of development; ore development and final data gathering in place

 

Demonstrated geological continuity of structures and mineralized domains

 

Slope of regression of the estimate analysis (calculated value during the Kriging Process)

Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data.

Ore classification as applied to the Mineral Resources disclosed in this document have been reviewed in detail on an area by area basis and are considered by the Authors to be appropriate and within the guidelines. The Mineral Resource has been classified as Measured, Indicated or Inferred Mineral Reserves.

  14.3.15

LOGGING

The drilling data provided for the Stawell Underground mineral resource estimate is collected over a period of 100-years and thus the drilling data base contains data of various levels of descriptive explanations of the geology. Underground drill-hole data has been collected over the last 30-years, all with descriptive explanations of observed geology. The geological logging code has essentially remained the same over the past 20-years which enable consistency in geological interpretation.

The logging information was considered for each model area and is of sufficient detail and quality to be used in the estimation at the current level of confidence.

  14.3.16

DATA SPACING AND DISTRIBUTION

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The Stawell underground mineral resource model areas were subject to varying drillhole density and sample locations in relation to the lode geometry. In most domains the drilling was of regular spacing and sufficient density closer to existing infrastructure but subject to decreasing densities and irregular spacing further away. Where the data spacing is sparser, the estimation in these areas was considered to be higher risk and is classified with less confidence. For classification purposes each mineralized domain was considered individually and where sufficient data density was present a classification solid was extruded.

Data spacing requirements are different for each model and drill spacing requirements for each model is considered individually. The data correlation as modelled by the variography gives the minimum drill spacing for confidence of correlation between holes. In areas of varying domain geometry, such as in proximity to faults, tighter drill spacing is required.

  14.3.17

ORIENTATION OF DATA IN RELATION TO GEOLOGICAL STRUCTURE

The orientation of the Magdala mineralization is interpreted to dip to the west between 50º and 85º. The drilling is considered to be appropriately targeted for this geological orientation. Where drillholes are considered too oblique to the interpreted lode they have been excluded from the modelling and estimation and are noted in the respective model reports.

  14.3.18

GEOLOGICAL INTERPRETATION

The geological interpretation of the Stawell underground mineral resource areas were undertaken by Newmarket Gold geologists. The interpretations were peer reviewed by Newmarket Gold’s Senior Resource Geologist who believes that they form a reasonable interpretation within the limits of the available data. The Authors agree with Newmarket Gold’s Senior Resource Geologist’s conclusions.

  14.3.19

DEPOSIT DIMENSIONS

The mineralized portion of the Stawell underground deposit extends down to -1800m RL, however, the remnant retreat of the Stawell Gold Mines which commenced at the end of 2012 saw the closure of the lower decline and decommissioning of the lower ventilation system. The classification of mineral Resource and Reserve was adjusted to reflect this change in extractive availability and all mineral Resource previously reported below the -900mRL was depleted. All Mineral Reserves below the -900mRL was mined.

The dimensions of the mineralization are adequately defined by the available drilling with acceptable extensions beyond data.

  14.3.20

ESTIMATION AND MODELLING TECHNIQUES

Refer to Section 14.3.4 Geological Modelling.

  14.3.21

MOISTURE

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Moisture determinations performed on RC chips immediately upon delivery to surface showed in-situ moisture contents to vary between 2 and 20%. After air-drying for two days (average exposure of core to air before sampling) moisture contents varied from 0.5 -2.5% . In-situ moisture is directly related to clay percentage (and therefore parent lithology) and is not directly linked to depth.

Estimates have been made on the basis of dry tonnes.

  14.3.22

EXTERNAL FACTORS AFFECTING EXTRACTION

Mining and extraction of the Magdala Mineral Resources and Reserves has been continuous since 1981. The Authors are not aware of any known environmental, permitting, legal, title, socio-economic, political or other issues that will prevent extraction of the remaining existing Mineral Reserves and Mineral Resources other than permitting requirements for the Big Hill project, as further described in section .

Stawell Gold Mines has been able to carry out its mining activities as defined by current Work Plans and approvals without intervention since 1981.

  14.3.23

BULK DENSITY

As detailed in Section 14.3.12 the bulk density factors used in this estimate are derived from previous data, and more recent data from diamond core intervals.

  14.3.24

CLASSIFICATION

All material within the mineral resource interpretation has been classified to represent the opinion of the Authors with regard to the risk in the mineral resource estimated. Within the mineralized domains that have been defined it is assumed that some of the material will form dilution to the mining of higher grade material. The classification of the Stawell Underground Mineral Resource into Measured, Indicated and Inferred Mineral Resources as set out below reflects the Authors’ view of this deposit as it is currently defined.

  14.3.25

SELECTIVITY ASSUMPTIONS

The mineral Resource estimate contains implicit assumptions of mining selectivity represented by the block size defined for each model (Y x X x Z). Block sizes for each resource area is summarized in Table 14-4.

  14.3.26

RESOURCE AUDITS OR REVIEWS

No Mineral Resource audits or reviews have been undertaken on the current western Ffank underground Minreal Resource between January 2015 and December 2015.

A high-level independent review of the eastern flank Aurora B resource modelling parameters was undertaken by Mike Stewart from QG Group in March 2106. Consideration of material included into the Inferred Category of classification was reviewed and determined appropriate to the geological understanding of the Aurora B Eastern Flank mineralization.

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  14.3.27

DISCUSSION OF RELATIVE ACCURACY/CONFIDENCE

At this stage no quantitative testing of the accuracy of the estimate or establishment of confidence limits has been undertaken.

  14.3.28

MINERAL RESOURCE STATEMENT

The total Mineral Resource inventory for Stawell Gold Mines as at 31st December 2015 is listed in Table 14-11.

The Mineral Resource Statements account for depleted mineral resources up to December 31st 2015 as a result of mining activities.

The depletion was carried out using underground development, stoping solids, historical voids as well as an existing pit surface (Davis). The as mined solids were taken up to the 31st December 2015.

The Stawell Gold Mines classified Mineral Resource Statements for Stawell Underground is tabulated below in Table 14-12. Cut-off grades applied to the resource was variable for underground ore depending upon width, mining method and ground conditions. Cut-off grades used for Mariners and Upper Levels mineral resource outside of current pit optimization is 2.0 g/t Au. A lower cut-off grade of 2.3 g/t Au was applied to all remaining underground resources.

Stawell Gold Mines Mineral Resources up to December 2013 have been reported as Exclusive of Mineral Reserves. Stawell Gold Mines Mineral Resources between January 1 2014 and December 31, 2015 are reported as Inclusive of Mineral Reserves. An adjustment to the reporting methodology of Reserves and Resources was made at December 2014 to ensure that Stawell Gold Mines complied with Crocodile Gold/Newmarket Gold reporting practices.

There are no known situations where the Mineral Resource estimate outlined above could be materially affected by environmental, permitting, legal, title, taxation, socio-economic, marketing or political or other relevant factors. There is, however, some risk with any gold mineral resource estimates where the gold price may affect the overall economic viability of a mineral resource.

 Stawell Gold Mines Resource  
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold
(Koz)
Measured 56 2.56 5
Indicated 4,063 1.85 241
Total (Measured and Indicated only) 4,119 1.86 246
Inferred 1,164 3.16 118

TABLE 14-11 STAWELL GOLD MINES RESOURCE AS AT 31 DECEMBER 2015

NOTES:

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1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:


  a.

Gold price of A$1,500/oz

     
  b.

Cut-off Grade applied was variable for surface resources. Grades used were as follows:


  0.44 g/t for Big Hill surface
  0.35 g/t for surface Low Grade Stockpiles

  c.

Cut -off Grade applied was variable for underground resources. Grades used were as follows:


  2.0 g/t for Mariners and Big Hill outside of current pit optimisation
  2.3 g/t for all remaining underground resources

4.

Surface and Underground Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines. Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Resources under NI 43 -101.

   
5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

The quantity and grade of the reported inferred mineral resources are uncertain in nature and there has been insufficient exploration to define the inferred mineral resources as indicated or measured mineral resources and it is uncertain if further exploration will result in upgrading them to an indicated or measured mineral resource category.

   
7.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
8.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Mineral Resource estimate for the Stawell underground mine is listed in Table 14-12.

 Stawell Underground Resource 
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Measured 56 2.56 5
Indicated 669 3.49 75
Total (Measured and Indicated only) 725 3.42 80
Inferred 1,118 3.24 116

TABLE 14-12 STAWELL GOLD MINES UNDERGROUND RESOURCE AS AT 31 DECEMBER 2015

NOTES:

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:


  a.

Gold price of A$1,500/oz

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  b.

Cut -off Grade applied was variable for surface resources. Grades used were as follows:


  0.44 g/t for Big Hill surface
  0.35 g/t for surface Low Grade Stockpiles

  c.

Cut-off Grade applied was variable for underground resources. Grades used were as follows:


  2.0 g/t for Mariners and Big Hill outside of current pit optimisation
  2.3 g/t for all remaining underground resources

4.

Surface and Underground Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines. Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Resources under NI 43-101.

   
5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

The quantity and grade of the reported inferred mineral resources are uncertain in nature and there has been insufficient exploration to define the inferred mineral resources as indicated or measured mineral resources and it is uncertain if further exploration will result in upgrading them to an indicated or measured mineral resource category.

   
7.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
8.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

Additional to the reported underground Mineral Resource and Mineral Reserve are a number of unclassified potential areas of mineralization which are currently the target of further investigation. These areas are currently built into the Stawell Underground Life-of-Mine ‘remaining unclassified resource’ mining program, and as such are sensitive to issues such as unidentified depletion, poor QA/QC drill information and or sparse drill information. Insufficient exploration data and confidence in these potential areas do not allow for these areas to be defined as a Mineral Resource, and it is uncertain if further exploration will result in the target being delineated as a mineral or minable resource.

The Mineral Resource estimate for the Big Hill Surface Project at Stawell Gold Mines is listed in Table 14-13 .

 Stawell Underground Resource 
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Measured      
Indicated 2,971 1.68 160
Total (Measured and Indicated only) 2,971 1.68 160
Inferred 46 1.15 2

TABLE 14-13 BIG HILL SURFACE RESOURCE ESTIMATION AS AT DECEMBER 31, 201 5

NOTES:

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:


  a.

Gold price of A$1,500/oz

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  B.

Cut-off Grade applied for Big Hill Surface Resource is 0.44 g/t Au


4.

Surface Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines . Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Resources under NI 43-101.

   
5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

The quantity and grade of the reported inferred mineral resources are uncertain in nature and there has been insufficient exploration to define the inferred mineral resources as indicated or measured mineral resources and it is uncertain if further exploration will result in upgrading them to an indicated or measured mineral resource category.

   
7.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
8.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.


  14.3.29

RECOMMENDATIONS

In order to improve the quality of the estimated resource the following actions are also recommended:

 

Undertake targeted scoping drilling on the faulted extremities of the orebody (above the Scotchmans Fault and below the South Fault).

 

Continue underground channel sampling and digital capture of the results to assist with determination of wireframe extents and aid the consideration of Mineral Reserves.

 

Continue to review the performance of the Mineral Resource estimate through regular reconciliation between the mining and the processing facilities.

 

Rebuild the Federal Albion Model wireframes and variography to incorporate learning from Upper Levels mining, for low grade stock definition.

 

Structural review of the Magdala Flanks to target unmined flat lodes.

 

Action the outcomes of the 2016 Internal Laboratory Audit, and commence quarterly internal Laboratory inspections.


  14.4

LOW GRADE STOCKPILE

The Mineral Resource estimate for the Stawell surface Low Grade Stockpile is listed in Table 14-14.

Stawell Low Grade Stockpile Resource
Domain Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Measured - - -
Indicated 423 0.43 6
Total (Measured and Indicated only) 423 0.43 6
Inferred - - -

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TABLE 14-14 LOW GRADE STOCKPILE RESOURCE ESTIMATION AS AT 31 DECEMBER 2015

  14.4.1

MT MICKE

Mt Micke is the waste stockpile from the Wonga open cut and underground mine. The Wonga Deposit has been previously mined by both open pit (1982-1985 and 2010-2011) and by underground methods (1985 -1999). Underground mining was completed in 1999 after which the mine was allowed to flood with water. The waste stockpile for the Wonga Deposit is located on the eastern margin of the open pit, Figure 14-7.

In 2011-2012 a grid program of Reverse Circulation drilling over the stockpile identified zones of low grade economic ore within the stockpile above an economic cut off of 0.35 g/t Au. This cut-off is defined as the lowest economic cut-off grade accounting for processing and transport.

A zone of 1,078,000 tonnes at 0.5 g/t Au was initially defined in 2012. From 2012 to 2014, 1,022,091 tonnes at 0.51 g/t Au has been processed through the Stawell processing facility, matching the defined grade almost identically. Given the bulk tonnes processed through the mill of the stockpile with precise reconciliation gives confidence to classification to the remainder of the defined low grade zone of the Mt Micke deposit.

In 2015 a grid program of Reverse Circulation drilling was conducted over the undefined areas of the stockpile, which has been exposed by depletion in previous years. Drilling allowed for definition of the base of the stockpile and further delineated the zones of low grade material. An economic cut off of 0.35 g/t Au was applied. This cut-off is defined as the lowest economic cut-off grade accounting for processing and transport.

A survey was undertaken over the Mt Micke low grade stockpile in early 2016 was used in depleting the defined low grade zone of the stockpile. Remaining tonnes from this zone is 423,000 tones with a grade of 0.43 g/t Au. This material was classified as Indicated given the high confidence in reconciliation of the stocks.

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FIGURE 14-7 LOCATION OF STAWELL LOW-GRADE STOCKPILE

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  15

MINERAL RESERVES


  15.1

INTRODUCTION

The Mineral Reserve estimate for the Stawell Gold Mines operations is listed in Table 15-1.

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven 51 2.49 4
Probable 3,428 1.46 161
Total Mining Reserve 3,479 1.47 166

TABLE 15-1 STAWELL GOLD MINES MINERAL RESERVE CLASSIFICATION – EFFECTIVE DECEMBER 31, 2015

1.

All Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by NI 43-101.

   
2.

Mineral Reserves were estimated using the following economic parameters:


  a.

Gold price of A$1,450/oz

     
  b.

Cut-off grade applied was variable for underground depending upon width, mining method and ground conditions.

     
  c.

Cut-off grade applied for Big Hill surface was 0.4g/t Au

     
  d.

Cut-off grade applied for Surface LG stockpile was 0.35 g/t Au


3.

Underground and Surface LG Stockpile Mineral Reserve estimates were prepared by Stawell Gold Mines personnel under the guidance of Wayne Chapman, Technical Manager Stawell Gold Mines. Mr Chapman is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 11 years of relevant mining engineering experience and is the Qualified Person for Reserves under NI 43-101.

   
4.

Big Hill Surface Mineral Reserve estimates were prepared by Mining One personnel under the guidance of Mark Edwards General Manager Exploration Newmarket Gold. Mr Edwards is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 18 years of relevant mining experience and is the Qualified Person for Reserves under NI 43-101.

   
5.

Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.


  15.2

MINERAL RESERVE ESTIMATE UNDERGROUND

The Mineral Reserve estimate for the Stawell underground mine is listed in Table 15-2.

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven 51 2.49 4
Probable 305 2.47 24
Total Mining Reserve 356 2.47 28

TABLE 15-2 UNDERGROUND MINERAL RESERVE CLASSIFICATION – EFFECTIVE DECEMBER 31, 2015

1.

All Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by NI 43-101.

   
2.

Mineral Reserves were estimated using the following economic parameters:


  e.

Gold price of A$1,450/oz

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  f.

Cut-off grade applied was variable for underground depending upon width, mining method and ground conditions.


3.

Underground Mineral Reserve estimates were prepared by Stawell Gold Mines personnel under the guidance of Wayne Chapman, Technical Manager Stawell Gold Mines. Mr Chapman is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 11 years of relevant mining engineering experience and is the Qualified Person for Reserves under NI 43-101.

   
4.

Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.


  15.2.1

MINERAL RESERVE ESTIMATE PROCESS

The following sections outline the process undertaken to produce Mineral Reserve estimates from the available Mineral Resources. This section contains descriptions of reserve design parameters, recovery and unplanned dilution factors, cut-off grades and depletion for mined material.

There are no known situations where the underground component of the Mineral Reserve estimate outlined above could be materially affected by environmental, permitting, legal, title, taxation, socio-economic, marketing or political or other relevant factors. There is, however, some risk with any gold Mineral Reserve estimates where the gold price may affect the overall economic viability of a Mineral Reserve.

  15.2.2

MINE RESERVE DESIGN UNDERGROUND

Mineral Reserve design has occurred underground at Stawell only in the areas identified as Upper South Fault 2 (USF2), Federal Albion South (FAS) and Upper Levels 250 below Scotchmans Fault (UL250) Mineral Resource areas. Additional areas are included in the mine plan but are in areas of lower confidence and are not included in Mineral Reserve estimates. The initial stage of the Mineral Reserve estimation process was the revision of the mining method selection chart. The mining methods that were considered for the Mineral Reserve estimation process were sill driving, down hole open-stoping, up-hole open stoping and select undercut and crown stoping where level spacing requires. Stoping panels are to be filled with raw fill, cemented rock fill or remain open depending on extraction sequencing, pillar recovery or geotechnical requirements. These methods were selected based upon previous experience at Stawell Gold Mines or because they were considered suitable for the ore zone geometry and geotechnical conditions present and expected.

  15.2.3

UNDERGROUND DESIGN AND RESERVE PARAMETERS

Open stope reserve shapes were created to cover all active and planned mining areas contained within the Federal Albion South, USF2 and UL250 Mineral Resource areas. These stope shapes did not necessarily reflect the final stope strike, height and/or crown pillar dimensions. Stoping widths vary from 3m out to 12m. Mining method selection criteria and applied design parameters are described in the Mining Methods Selection process (see Section 16)

The open stope Mineral Reserve wireframe design parameters applied were:

  Strike length dictated by grade distribution in block model

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  Minimum true bench width of 3m
  Maximum benching height of 30m vertical from backs to floor
  Internal waste incorporated within the stope block design

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TABLE 15-2 MINING METHOD SELECTION

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Mining recovery from open stopes at Stawell is principally influenced by the following factors:

  Accuracy of the geological interpretation.
  Accuracy of the production hole drilling.
  Stope dimensions.
  Sill drive dimensions and position relative to bench stope.
  Presence or absence of adjacent filled voids and pillars.
  Geotechnical integrity of stope and sill drive walls.

The above factors manifest themselves as ore loss in the following ways:

  The need for planned pillars due to accessing of ore blocks in low grade areas.
  Frozen rings due to ground movement.
  Bridged stopes.
  Unplanned ore pillars left to improve ground support.

Unplanned dilution in open stopes underground at Stawell is a function of the following factors:

  Regional geotechnical conditions.
  Location of sill drives relative to the open stope.
  Width of sill drives relative to the open stope width.
  Production drilling accuracy.
  Quantity, quality and type of ground support in sill drive walls.
  Speed of ore extraction from active stopes.
  Length of time sill drives have been open before stoping commences.

In order to correctly apply recovery and dilution factors to all underground stopes in the Mineral Reserve, factors such as orebody dip, rock RQD and development and stope sequence were considered. Table 15-2 and Table 15-3 show the general recovery and dilution factors that were applied to the reserve blocks (some individual blocks have lower factors applied).

Description

Recovery Factor -
Tonnes
Dilution Factor -
Tonnes
Comments

Stoping – UL 250

95% 15% Remnant open stoping within Upper Levels mining area.

Stoping – U2

95% 15% Bottom CRF and rock fill stoping, combination of undercut and crown extraction.

Stoping – Federal Albion South

95% 15% Bottom up rock fill or CRF stoping. Undercut and crown extraction where stope geometry requires.

Strike Development

100% 10%

TABLE 15-3 RECOVERY AND DILUTION FACTORS FOR THE UNDERGROUND RESERVE BLOCKS AT STAWELL GOLD MINES

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  15.2.4

GOLD CUT-OFF GRADES UNDERGROUND

Table 15-4 shows the calculated lower cut-off grades used in the estimation of the Mineral Reserves. Cost assumptions are based on the 2015 budget.

Description g/t Au
Open Stope – full 2.15
Open Stope - marginal 1.80
Development - marginal 2.70

TABLE 15-4 GOLD CUT-OFF GRADES UNDERGROUND

For certain other situations, a lower cut-off grade is applied. For development, which is justified for other reasons (i.e. access to a higher grade block or infrastructure considerations), the cut-off grade is lowered to reflect that the material only has to cover the non-mining costs to break even. This is only applied if the development material had to be trucked to surface anyway and if it is not displacing higher-grade ore from the mill. Likewise, for incremental stoping production where the development has already been mined (i.e. for access to a higher-grade block), the cut-off grade is lowered to reflect that the development cost has already been incurred.

Underground stope and development shapes are limited in their extremity by the application of appropriate cut-off grades (see Table 15-4) and a full conceptual design is subsequently created around the resultant shapes. This design includes but is not necessarily limited to decline design, associated level infrastructure and vertical development.

  15.2.5

DEPLETION AND RESULTS

The Mineral Reserves reported above are largely the result of work undertaken to Dec 31st 2015 and reported by Newmarket Gold under Canadian reporting requirements in accordance with NI 43-101. The evaluation models have been depleted for material mined up to December 31st 2015. This process applied to only the Federal Albion South mining area.

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  15.3

MINERAL RESERVE ESTIMATE BIG HILL

The Mineral Reserve estimate for the Stawell Big Hill Surface Project is listed in Table 15-3.

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven - - -
Probable 2,700 1.51 132
Total Mining Reserve 2,700 1.51 132

TABLE 15-3 BIG HILL SURFACE MINERAL RESERVE CLASSIFICATION – EFFECTIVE DECEMBER 31, 2015

1.

All Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by NI 43-101.

   
2.

Mineral Reserves were estimated using the following economic parameters:


  g.

Gold price of A$1,450/oz

     
  h.

Cut-off grade applied was 0.4 g/t Au.


3.

Big Hill Surface Mineral Reserve estimates were prepared by Mining One personnel under the guidance of Mark Edwards General Manager Exploration Newmarket Gold. Mr Edwards is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 18 years of relevant mining experience and is the Qualified Person for Reserves under NI 43- 101.

   
4.

Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.


  15.3.1

MINERAL RESERVE ESTIMATE

The following sections outline the process undertaken to produce Mineral Reserve estimates from the available Mineral Resources for the Big Hill surface. This section contains descriptions of reserve design parameters, recovery and unplanned dilution factors, cut-off grades and depletion for mined material. This section has been compiled by Mark Edwards, who is the Qualified Person for the Mineral Reserve Estimate (see Section 2.3)

Big Hill Surface Reserves are subject to project permitting conditions. There is some risk that permitting conditions will not be realized or Mineral Reserve Estimates will need to be modified.

  15.3.2

MINE RESERVE DESIGN

Big Hill surface Mineral Reserve estimates are based upon a Lerchs-Grossman algorithm optimization technique utilizing Gemcom’s Whittle 4X software. This produced a series of nested pit shells utilizing a range of parameters from slope angles, Mining Cost Adjustment Factors (MCAF’s) all utilizing a diluted block model. Ore Reserve number are from a final pit design based on a suitable “Whittle” shell.

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  15.3.3

BIG HILL SURFACE DESIGN AND RESERVE PARAMETERS

Big Hill surface Mineral Reserves were created utilizing a diluted block model, minimum ore block size (1.5m minimum mining width) and maximum mining block size of 10m. Pit benches are designed at 5m with 2.5m mining flitch.

Final Pit Design parameters are as follows;

  Ramp width 12m single, 25m double
  Ramp grade 1:9 with 1:10 localized maximum
  Batter Height 20m
  Berm width 10m weathered zones 8.5m minimum
  Batter Slope 60°
  Overall Slope Angle 42.9°

In order to estimate the expected dilution and ore recovery for Big Hill surface Mineral Reserves Datamine Minable Shape Optimiser software (MSO) is used to create a diluted model that mimics what an engineer or geologist might do when generating dig outlines on adjacent sections. This generates optimal mining ore blocks that take into account footwall and hangingwall dilution skins to ensure mining areas remain above cutoff grade.

The following parameters are used to define a potentially mineable ore block:

  Diluted Cut-Off grade (0.4 g/t Au)
  Near and far wall dilution skin (0.5m & 0.5m)
  Minimum ore block size (1.5m minimum mining width)
  Maximum mining block size (10m)

The subsequent MSO wireframes were used as a guide for a new block model in which the diluted grades generated using the MSO were added to the new block model prior to the Whittle evaluation and final design. The subsequent dilution, grades and recovery are displayed below.

  Undiluted Diluted
Tonnes (>0.4 g/t Au) 8,106,899 9,958,474
Avg. Grade g/t Au 1.70 1.48
Avg. Dilution g/t Au - 0.15
Avg. Mining Recovery - 0.07

TABLE 15-5 MINABLE SHAPE OPTIMISER – BIG HILL DILUTION

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  15.3.4

GOLD CUT-OFF GRADES

Table 15-4 shows the calculated lower cut-off grades used in the estimation of the Mineral Reserves. Cost assumptions are based on the 2016 budget.

Description g/t Au
Big Hill Surface 0.40

TABLE 15-6 GOLD CUT-OFF GRADES, BIG HILL

Surface cut-off grade is used for determination of ore and waste based predominately on process, administration, Royalty and differential mining cost to allow for haulage to the ROM.

  15.3.5

DEPLETION AND RESULTS

The Mineral Reserves reported above are largely the result of work undertaken in January 2016 and reported by Newmarket Gold under Canadian reporting requirements in accordance with NI 43-101. The evaluation models have been depleted for material mined up to January 1st 2016.

  15.4

MINERAL RESERVE ESTIMATE SURFACE STOCKPILES

The Mineral Reserve estimate for Stawell surface Low Grade Stockpiles is listed in Table 15-4.

Classification Tonnes (Kt) Gold Grade g/t Ounces Gold (Koz)
Proven - - -
Probable 423 0.43 6
Total Mining Reserve 423 0.43 6

TABLE 15-4 MINERAL RESERVE CLASSIFICATION FOR LOW GRADE STOCKPILES– EFFECTIVE DECEMBER 31, 2015

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Surface Mineral Reserves were estimated using the following economic parameters:


  a.

Gold price of A$1,450/oz

     
  b.

Above 0.35 g/t Au cut-off


3.

Surface Low Grade Mineral Reserve estimates were prepared by Stawell Gold Mines personal under the guidance of Wayne Chapman, Technical Manager Stawell Gold Mines. Mr Chapman is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 11 years of relevant mining engineering experience and is the Qualified Person for Reserves under NI 43-101.

   
4.

Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

The Mineral Reserve cut off grade calculations for Stawell surface Low Grade Stockpiles is listed in Table 15-7.

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  Mining Cost Milling Cost Recovery Gold Price Cut-off
  ($/T) ($/T) (%) ($/Oz) ($/gm) (g/t Au)
2016 Budget $2.70 $9.90 90 $1,450.00 $46.62 0.35

TABLE 15-7 LOW GRADESTOCKPILE CUT OFF CALCULATION

Processing of the Mt Micke Low grade stockpile has returned 90%.

There are no known situations where the Mineral Reserve estimate outlined above could be materially affected by environmental, permitting, legal, title, taxation, socio-economic, marketing or political or other relevant factors. There is, however, some risk with any gold Mineral Reserve estimates where the gold price may affect the overall economic viability of a Mineral Reserve.

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  16

MINING METHODS


  16.1

INTRODUCTION

Newmarket Gold’s Australian operations are currently mining the Stawell Gold Mines as an underground mine and intend to continue mining the upper levels down to the 1065mRL. The Stawell Gold Mines workings are accessed via a haulage decline commencing from a surface portal. Current mining activity is centered on the Magdala Lode utilizing long-hole open stoping. Modern mining has extended down to -1650m RL into the Golden Gift Lodes; however, in January 2013 Stawell Gold Mines made a decision not to continue mining at depth and now all exploration and mining efforts are focused above the -1065mRL.

Surface mining operations are planned around a conventional excavator and truck operation with light blasting activity only required in the very lower portion of the excavation. Additional controls are planned for air quality and amenity considerations around hours of operation, benching configuration, dust suppression and equipment specifications.

  16.2

MINE DESIGN UNDERGROUND


  16.2.1

MINING METHOD DESCRIPTION

The underground mining activity occurs in the Magdala Mine which is accessed by a decline from a portal located adjacent to the mill. The mine access development and services are located mainly within basalt. Ground conditions are good and there is no history of major seismic activity in the current mining areas. Development follows the Magdala Lode system down plunge. At the 290RL the Federal Albion decline splits off the main and between the 470RL and 786RL the decline splits into the north and the south decline to enable further access to the Magdala System. To facilitate ore access, extraction levels are developed at approximately 20m to 25m vertical intervals. The accessible mining areas currently extend over approximately 2.5km of strike to more than 1,000m below surface, measured from the top of Big Hill.

The underground mining method used in the Magdala Mine was bench stoping with cemented rock fill pillars in primary stopes, and rock filled secondary stopes. The mining method used in the narrow Magdala ore zones is retreat open stoping with either Cement Rock Fill (CRF) if full extraction, or combinations of CRF and rock fill or all rock fill stope with pillars. In the Magdala orebody, stope sizes typically range from 2,000 to 10,000 tonnes. Stope ore is recovered using loaders under direct or remote control of an operator, with haulage by 60 tonne trucks.

The access decline is used as an intake airway. Exhaust air is drawn through the workings by a series of ventilation rises and drives by the primary ventilation fan installation located at the northern end of the mine. The mine is relatively dry. Water pumped from the upper portion of the workings is recycled for use in the mine or the treatment process.

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Since the economic completion of the Golden Gift orebodies in 2013, mining has occurred in lower grade remaining Mineral Resource areas above 1065mRL. This program commenced as remnant mining programs in known low grade and previous pillar areas. Since commencement of these programs in 2013 there has been an increasing contribution from the current planned reserves located in localized mineralized zones along the margins of the orebody or from low grade previously identified potential Mineral Resource areas. Over 50% of the mine plan is from areas of unclassified or short turn around zones of the orebody that are defined and mined within an annual reporting period.

The mining method used in the planned Mineral Reserves is retreat open stoping with either CRF if full extraction required or rock fill or open stopes with pillars. Stope panels in the Mineral Reserve and unclassified material programs typically range from 2,000 to 10,000 tonnes extracted from level intervals at approximately 20 to 25m vertical intervals.

Mine production over 2015 has produced around 500kt from underground low grade Mineral Resource, and remnant programs with a further 400kt low grade stockpile oxide feed to supplement programs for a total mined ounce profile of around 38-39k ounces. This production profile is expected to continue through 2016. Evaluation for identification of areas of remnant mining potential similar to those conducted from 2013 onwards remains ongoing through 2016. This is expected to produce additional material to be included in the 2016 mine plan with additional feed to maintain underground operations into 2017 and beyond.

Drill programs into areas of remaining lower grade Mineral Resource potential above the current Federal Albion South (FAS) and along the Upper South Fault 2 (USF2) areas could have the potential to extend underground mine life activities based on previous conversion rates of Mineral Resource target areas.

Underground mining is conducted in nominal 5m by 5m development using a conventional fleet including jumbos, production drills, ancillary equipment and loaders and trucks in the 20t and 60t ranges respectively. Current mining is undertaken as owner miner with most activity conducted in previously accessed areas.

Mining plans access around 450m per month with around 65% of this being new development and the remainder light rehabilitation of previous access areas.

  16.2.2 UNDERGROUND MINING EQUIPMENT

All equipment required for the current planned underground mining at Stawell Gold Mines is currently onsite and utilized, with this equipment being either fully owned or in the case of two load haul dump and one production drill rig being hire option to purchase arrangements.

Table 16-1 below lists the current underground mining fleet.

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Equipment Quantity
Development Jumbo 2
Production Drill 1
Load Haul Dump 4
60t Haulage Truck 4
Service/Charge unit 3

TABLE 16-1 UNDERGROUND MINING MAJOR MOBILE EQUIPMENT– EFFECTIVE DECEMBER 31, 2015

  16.3

MINE DESIGN SURFACE MINING


  16.3.1

MINING METHOD DESCRIPTION

Surface mining activities are planned as an adaptation of conventional open cut mining where benching techniques of predominately rip and dig or free dig mining has been proposed. In order to provide best practice mitigation measures for air quality and amenity, the majority of operational activity will occur as low as practicable in the pit to enable upper benches to provide an element of shielding, see Figure 16-1 below. In order to further preserve amenity operating hours are daylight hours only Monday to Friday.


FIGURE 16-1 BIG HILL SURFACE MINING METHOD SCHEMATIC

It is expected that much of the Big Hill surface Mineral Reserve will be excavated without blasting as the majority of surface drill holes rank in the “Easy Digging” to “Easy Ripping” categories of the Pettifer and Fookes Excavatability Assessment. Some areas at depth reach the “Extremely Hard Ripping” category but these are the minority and light blasting is planned to be applied to these areas.

Surface mining plans commence with the excavation of the North Pit and a two stage cutback approach to mining of the South Pit with waste material from the South Pit being used to reconstruct the North Pit. Surface mining fleets at peak operation are to consist of five 90t trucks, a 120t excavator and a 190t excavator.

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  16.3.2 SURFACE MINING SCHEDULE

Mining schedules for the extraction of the North and South Pits at Big Hill are based around levels of extraction to ensure that suitable environmental conditions can be met at any point in time. Extensive air quality and noise modelling of project parameters have provided a schedule quantity that can be operated within suitable parameters. This schedule is presented below in Figure 16-2.


FIGURE 16-2 BIG HILL SURFACE MINING SCHEDULE

  16.3.3

SURFACE MINING EQUIPMENT

Mining equipment requirements for extraction of the North and South Pits at Big Hill are based on first principle productivities using excavator specifications and benchmarks from existing operations, Table 16-2.

Excavator Productivity

Unit Value Value

Excavator Type

  EX1200 EX1900-6-BH

Truck Type

  Cat_777F Cat_777F

Digger Configuration

  Backhoe Backhoe

Material Detail

     

Dry Density

t/BCM 2.1 2.1

Moisture Content

% 2% 2%

Swell Factor

% 15% 15%

Wet Density Loose

t/m^3 1.86 1.86

Wet Bank Density

t/m^3 2.14 2.14

Shovel Details

     

Bucket Heaped Cap.

m^3 6.7 9.6

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Excavator Productivity

Unit Value Value

Fill Factor

% 85% 85%

Bucket Cap. Volume

BCM 5.0 7.1

Bucket Cap. Weight

t 14.7 19.2

Bucket Cap. Weight

BCM 6.9 9.0

Bucket Cap. Adopted

BCM 5.0 7.1

Truck Details

     

Tray Capacity

m^3 60.2 60.2

Truck Fill Factor

% 95% 95%

Volume Limit

BCM 49.7 49.7

Rated Payload

t 90.7 90.7

Assumed Overload

% 0 0

Adjusted Payload

t 90.7 90.7

Weight Limit

BCM 42 42

Adopted Capacity

BCM 42 42

Min. Bucket Fill

% 25% 25%

Calculated Passes Per load

  9 6

Calculated Passes Per load

(rounded) 9 6

Actual Truck Load

BCM 42 42

Actual Truck Load

t 90.7 90.7

Dump Time

min 1.2 1.2

Excavator Productivity

     

Cycle Time

sec 25 25

Efficiency Factor

% 100% 100%

1st Pass

sec 9 9

Truck Exchange

sec 38 38

Loading Time

min 4.12 2.87

Max. Productivity

BCM/OH 617 886

Effective Utilisation of operating hours

% 73% 76%

Productivity

BCM/OH 450 670

Productivity

t/OH 945 1,407

TABLE 16-2 PRODUCTIVITY ESTIMATES

The productivity estimates for each excavator is matched against the mining schedule to determine the required number of excavators. Figure 16-3 below illustrates the total excavator capacity for an EX1900 and EX1200 versus the mining schedule.

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FIGURE 16-3 EXCAVATOR REQUIREMENTS BIG HILL

Haulage of ore and waste were modelled in TALPAC software to determine travel time and fuel consumption. Travel time and fuel consumptions were calculated for all haulage routes from each bench to the ROM or waste dump.

In addition, travel times were used to determine the productivity of each truck and thus truck requirements were matched to the mining schedule. Figure 16-4 below illustrates the haul truck requirements.

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FIGURE 16-4 HAUL TRUCK REQUIREMENTS BIG HILL

Ancillary equipment requirements for the project are based upon best environmental management practices and include additional bulldozers to reduce drill and blast requirements as well as additional water cart capacity.

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  17

RECOVERY METHODS


  17.1

MINERAL PROCESSING

The gold processing facilities utilized at SGM comprise a standard Carbon-In-Leach (CIL) gold recovery circuit following crushing and grinding and sulphide flotation. The treatment plant consists of five unit processes Figure 17-1. These are:

  size reduction (crushing and milling)
  gravity gold recovery
  flotation/ultra-fine grinding
  leach-adsorption, and
  gold recovery

Geographically the plant can be split up into five main areas. These are:

  the primary crushing circuit
  the milling circuit
  the flotation/ultra- fine grinding circuit
  the leach-adsorption circuit, and
  the elution/electrowinning circuit

A current processing flow sheet is shown in Table 17-1. A history of tonnage and grade throughput for the processing facilities is shown in Section 6 of this report.

Coarse gold (up to 30% of the gold in mill feed) is recovered from the milling circuit in self-cleaning centrifugal gravity concentrators. Approximately 75% of the ore requires further liberation of the gold from sulphides and this is achieved in a two stage flotation circuit where gold bearing sulphides (pyrite, arsenopyrite and some pyrrhotite) are concentrated. The sulphide is ground to approximately 0.01 millimeters in an ultra-fine grinding mill to liberate enclosed gold (up to 20% of the gold in mill feed). The ground sulphides and flotation tail are recombined and sent to the CIL circuit.

Stawell ore exhibits various degrees of preg-robbing of gold. Preg-robbing occurs when naturally occurring carbon species (graphite) in the ore rob gold from the pregnant liquor in the leach circuit, thus reducing the gold recovery. To combat this, kerosene is added to foul the naturally occurring carbon before it enters the leach circuit and a simple preg-rob index developed at SGM indicates the rate of addition needed for the kerosene to be most effective.

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FIGURE 17-1 SGM TREATMENT PLANT FLOW SHEET

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An ongoing program of metallurgical test work is conducted at SGM. The program utilizes diamond drill core to determine the expected plant recovery for all ore blocks at a stope scale within the immediate and long term mine plan. Samples of the ore and estimated dilution are tested to determine the expected preg-rob index and expected gold recovery through the SGM processing circuit. The results of the test work program provide an expected plant recovery on a campaign basis. Metallurgists are able to plot the actual versus predicted plant recoveries using the test work results so as to show the relationship between actual plant recovery and expected plant recoveries for all float ore treated project to date. This validates the robustness of the metallurgical test work programs utilized by SGM and as such the robustness of the forecast metallurgical assumptions used in developing project schedules and financial forecasts.

Infrastructure and systems to meet all current and projected energy, process water and consumable requirements are in place and operating with no expected impairments over the course of the expected mining.

Total gold recovered in the reporting period is given in Table 17-1.


YEAR

PERIOD
Average.

REC %
Recovered.

OUNCES
2015 JAN - DEC 80.78 36,321

TABLE 17-1 STAWELL GOLD MINES TOTAL GOLD RECOVERED JANUARY 2015 TO DECEMBER 2015

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  18

PROJECT INFRASTRUCTURE

The Stawell Gold Mines has been operating for over thirty years and therefore has all major infrastructures in place.

Milling infrastructure is related to the Stawell Gold Mines processing facility, which is outlined above in Sections 5 and 17.

  18.1

SURFACE INFRASTRUCTURE

Stawell Gold Mines facilities are extensive and representative of a modern gold mining operation (see Figure 18-1). The main site location comprises:

 

Office and administration complex

 

Store and storage facilities

 

Heavy underground equipment workshop and light vehicle workshop

 

Surface run of mine stockpiles

 

Gold processing plant and associated facilities

 

On site assay and metallurgical test work laboratory

Four freshwater storage dams to store rainfall run-off and mine dewatering which is used in the plant or around the mine site.

Power for the plant is fed from a main transformer located adjacent to the administration complex.

 

A batch plant for preparing shotcrete for underground support

 

Core farm and core processing facility

Surface facilities include the gold processing plant, offices, core shed, laboratory and workshops. Larger infrastructure onsite includes tailings dams covering 96 ha and receiving all tailings from the processing plant. Four freshwater dams occur throughout the mine lease.

Main haul road facilities are yet to be constructed for surface mining activities.

  18.2

TAILINGS AND STORAGE FACILITIES

Since operations began in 1984, three tailings dams have been constructed and operated, two of which have since been decommissioned:

Reserve Tailings Dam. This has been decommissioned and rehabilitated to a clay target shooting complex.

 

No 1 Tailings Dam. This has been decommissioned and partially rehabilitated.

 

No 2 Tailings Dam remains in operational.

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All dams were constructed as earthen embankments with upstream sub-aerial deposition and are subject to annual integrity and operational review by an independent industry expert. No. 2 Tails Dam or TSF2 currently has an approved Work Plan Variation enabling the completion of a wall lift to 253RL (the current wall level is 250RL). Review of remaining underground and Mt Micke oxide stockpile stocks by the dam designer (as per the above independent industry expert) has shown that to treat the mineralization expected from the Big Hill pits, a lift to 252RL will be required.

  18.3

WASTE DUMPS

No active surface waste dumps are in use, with all waste material from underground operations being used for filling requirements. Previous surface waste dumps are being processed as low grade oxide stocks.

Surface mining requires the construction of a temporary waste rock storage facility. This would be located partly over the existing Davis waste dump area and extend over private land towards Landsborough Road. This dump will be in operation from commencement of mining in the North Pit until reconstruction of the South Pit is complete after approximately 5-6 years.

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FIGURE 18-1 PLAN SHOWING THE LOCATION OF MIN 5260, STAWELL GOLD MINES OPERATIONAL INFRASTRUCTURE

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  18.4

UNDERGROUND MINE INFRASTRUCTURE

Underground infrastructure within the Magdala Mine is extensive and includes a 5.5m x 5.5m access decline extending from surface to around 1600mRL as at the end of December 2015. The gradient of the decline is 1:8 down to 468mRL then changes to 1:7 to 1600mRL. A number of historic surface shafts and the main decline supply fresh air to the mine. The exhaust system comprises two parallel 500kW Korfmann AL22-2500 Primary fans exhausting through the Darlington shaft, which is the only exhaust.

Mine dewatering systems had been de-commissioned with the lower portion of the mine allowed to flood. Estimations are at around 50 years for the ground water levels to return to pre-mining levels. During 2015 mine pumping systems are being re-commissioned progressively to the 806mRL.

An underground workshop and underground crib room is available at the 800mRL level within the mine.

In addition to the fixed plant, Stawell Gold Mines owns, operates and maintains all mobile mining equipment including jumbo development drills, production drills, loaders, trucks, and ancillary equipment required to undertake mining operations.

  18.5

POWER

Stawell Gold Mines purchases power under contract from Origin Energy Australia.

Supply from the National Electricity Grid to Stawell Gold Mines is via high voltage installations in two locations;

 

Moonlight Substation (10 Mega Watts feed) which supplies the Magdala Underground operation, and

Reefs Road (7 Mega Watts feed) that supplies the gold processing plant, administration, workshop facilities and parts of the upper levels of the Magdala mine.

Power to underground from the Moonlight substation is supplied through a 990m steel cased borehole, 400m borehole and that from the Reefs Road substation via the Magdala decline.

The total facilities for electrical power input distribution to site is 5,833,600 kwh/month

The main underground demands are outlined below,

  1.

Primary ventilation fan sizes are

 

482L 1000kw

  2.

Compressor sizes are

 

1300 cfm x 4(250kw x 2 units)

 

1000 cfm x 2(132kw x 1 unit underground)

 

650 cfm x 3(110kw x 3 units)

The total volume of ventilation supplied to the mine is 180m3/sec.

The current power availability is sufficient to meet the needs of the current life of mine operating plan.

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  18.6

WATER

Water supply is from harvested rainfall runoff, mine dewatering, recycling of process water from the tailings facility, and by way of a 1ML/day raw water right entitlement and urban customer access to potable supply from GWMWater. This water is sourced from the nearby Grampians Mountains. The capacity of the site water storages is approximately 690ML.

The GWMWater water is supplied as raw or potable and is preferentially used, when required in the processing operations as it improves gold recovery.

  18.7

EXISTING PUBLIC INFRASTRUCTURE

For the mining operations to commence there is a number of items of public and private infrastructure to be replicated elsewhere and then removed or decommissioned from the project area.

These items include:

  Fire watch tower
  Town water supply system
  Optic fiber route across the North Pit
  Power reticulation for communications and fire watch facility
  Pioneer Memorial Rotunda and access roads
  Arboretum gates and various other monuments from 1930’s through to 1975

Feasibility level studies have been conducted on the replication, removal and re-establishment of the above items with costing included in the Capital assessment of the project.

Fire-watch tower - Replacement facilities will be constructed adjacent to the existing facility with necessary access off Scenic Road. Detailed design is to be progressed.

   

Town water supply system - As a risk mitigation strategy the current Stawell gravity feed treated water supply system will be decommissioned, with a pressure pumping system to be installed on the regional water authority managed crown land. This system will involve the installation of 600m of mains delivery system; construction of 10 Meg L treated water storage capacity and pumping capability for the service of both high and low elevation town water feed. Detailed design and site investigation is to be progressed.

   

Optic fiber – As short length of an optic fiber data and communications cable is routed across the North Pit mining area. This is to be re-routed around the pit operation area to the existing communications tower.

   

Power reticulation - Power reticulation will need to be re- routed to the communications and firwatch facilities.

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Memorial Rotunda and access road - A memorial rotunda currently serviced by three roads is situated close to the crest of Big Hill, this will be deconstructed and stored for the duration of mining of the North Pit. This facility with one access road and car parking area is to be re-established following reconstruction of the Big Hill land form.

   

Other Monuments - Provision has been made for the deconstruction and reconstruction of a number of monuments spanning a range of structures, ages and construction methodology. These structures range from stone fascia entrance gates, memorial, stone construction memorial seating and a medium solid stone monument.


FIGURE 18-2 PLAN SHOWING THE EXISTING AND REPLACEMENT WATER SUPPLY SYSTEM

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  19

MARKET STUDIES AND CONTRACTS


  19.1

MARKETS

Gold doré produced by Newmarket Gold is currently shipped to the Perth Mint for smelting. On notification of produced amounts of gold dore, Newmarket Gold notifies Auramet Trading LLC (Auramet) of the upcoming gold shipment and deals for future delivery of gold to the mint. During this process a given number of ounces will be included in the sale amount for delivery directly after smelting has occurred. Generally, a period of one week is required for this process to occur.

Auramet is a New Jersey based company, which specializes in the sale of both base metals such as copper, nickel and zinc and precious metals such as gold, silver, platinum and palladium.

During the smelting process the mint can extract other minerals. The main other economic mineral that is recovered is silver, which is sold to Perth Mint.

  19.2

GOLD PRICE

To determine the Australian denominated gold price to use in the Mineral Resource and Mineral Reserve calculations, reference was made to publicly available price forecasts by industry analysts for both the gold price in US dollar terms and the AU$/US$ foreign exchange rate.

This exercise was completed in December 2015, and yielded the following average gold forecast prices and corresponding average forecast US$:A$ FX rates.

For Mineral Reserve purposes, a US$1,100/oz gold price was used and an FX rate of $0.76 for an approximate Australian dollar gold price of A$1,450 per ounce.

For Mineral Resource purposes, a US$1,125/oz gold price was used and an FX rate of $0.75 for an approximate Australian dollar gold price of A$1,500 per ounce.

The average US$ gold price per ounce for the last three years was as follows:

2013 - US$1,411
2014 - US$1,266
2015 - US$1,160

The Qualified Persons are not aware of any agreements that are not within market parameters.

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  20

ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT

Stawell Gold Mines is committed to maintaining effective management systems with respect to environmental matters on site. At Stawell Gold Mines, site management monitors and regularly reviews the environmental and social impacts of the operations, such as water quality, air quality, blast vibration, noise and community attitudes. Site environmental performance results are shared with regulatory authorities and local communities quarterly through the Environmental Review Committee (ERC) meeting. Stawell Gold Mines promotes responsible environmental behavior among all employees and contractors. Operations are managed in accordance with the Stawell Gold Mines Environmental Management Plan (EMP) which assesses Stawell Gold Mines’ environmental risks and provides the standards, procedures, and guidelines required to mitigate and manage any risks.

  20.1

ENVIRONMENTAL RESEARCH

As a practice, Stawell Gold Mines personnel conducts research in all environmental fields of Stawell Gold Mines to better understand any potential impacts and ensure best practices for closure and remediation. Environmental research is done by Stawell Gold Mines staff, consultants or though university partnerships. The main areas of research are in tailings storage facility management both during operations and post closure, noise of operations, rehabilitation of vegetation and air quality of operations.

Stawell Gold Mines has three tailings storage facilities (TSF) with one operational (TSF2) and two rehabilitated (TSF1 and the reserve tailings dam) to a planned end use. In 2009, Stawell Gold Mines engaged O’Kane Consultants, a world leader capping design consultant, to design a store and release cap for the long term stability of the tailing storage facilities. This design has been installed on the two rehabilitated TSF’s. Ongoing research by O’Kane Consultants in monitoring capping design methodology occurs and is reported on annually.

Stawell Gold Mines has had a long partnership with the University of Melbourne and Curtin University to undertake research and trials on remediation and rehabilitation. The University of Melbourne has led studies in both restoration ecology of TSF’s and waste rock stockpiles and bioremediation of groundwater. The results of the restoration ecology trials has found that the local native vegetation can survive and flourish in the proposed capping design and there are no impacts on the vegetation or subsequent food chain from the vegetation roots entering tailings. The results of the bioremediation work has isolated a naturally forming local bacteria that degrades thiocyanate (SCN) and have been undertaking laboratory trials with the intention of field trials in 2016. The work undertaken by Curtin University has investigated the acid generating potential of the waste rock and tailings and determined that Stawell Gold Mines tailings were non-acid forming due to a high degree of buffering minerals available.

Stawell Gold Mines has detected SCN in some monitoring bores immediate to TSF2. The groundwater in several areas is also elevated due to hydraulic pressure from TSF2. Stawell Gold Mines has undertaken multiple studies of the area and the impacts from the seepage (Coffey (2008), LanePiper (2008), Rockwater (2010), NQ Groundwater and Environment (2011), NQ Groundwater and Environment (2012), Cardno LanePiper (2014)). An environmental audit of groundwater (Cardno LanePiper, 2014) was completed for the tailings storage facilities at the direction of the Environment Protection Authority Victoria (EPAV) in 2014. The conclusions to the audit environmental risk assessment based on the source-path-receptor (Conceptual Site Model) and rehabilitation proposed for TSF2 indicated that the risk to the beneficial use of stock watering to be “moderate”, given the unlikely use based on groundwater salinity although it is possible due to the presence of an extraction bore (775m north-east) in the vicinity. The risk to primary contact recreation and buildings and structures to be “low” and “no risk” to the maintenance of aquatic ecosystems and industrial water use. The closest water receptor to the TSF is Concongella Creek and the audit (Cardno LanePiper, 2014) stated that under present operating conditions seepage would take between 150 to 240 years to reach the distance required (1.3km) .

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A remediation feasibility study was undertaken in 2015 by URS Australia in compliance with a notice from EPAV. The study concluded that hydraulic containment (pump and treat) would be the best option for remediation of SCN detected in groundwater immediate to TSF2. Stawell Gold Mines is currently working with EPAV under a notice to develop and implement a clean-up plan in 2016.

In a review conducted by NQ Groundwater and Environment in 2012 during the installation of five new seepage interception bores it has been indicated that due to very high clay content and high weathering rates it was demonstrated that seepage mitigation from TSF2 is very slow ranging from 0.0016m/day to 0.067m/day. Increasing distance, dilution and absorption of cyanide to geological materials down gradient will slow migration. During mining operations, to manage elevated water levels and seepage from TSF2, rock filled cut-off seepage trenches have been installed in several locations around the dam. Extraction bores have been installed in areas of seepage and groundwater rise to mitigate any impacts in accordance with the regulatory approved site groundwater management plan.

It is expected that with capping and closure of TSF 2, the tailings will progressively dewater resulting in groundwater levels slowly re-establishing to near pre-TSF levels. This is supported by the modelling works undertaken by O’Kane Consultants (O’Kane 2010). Therefore the need for active groundwater level management is anticipated to decline once the operation of TSF2 and closure is complete.

Research on the noise of operations has been undertaken with modelling of the site operations including milling and haulage against the mining license requirements due to the proximity of the mine site against the town. Results of the investigations have found opportunities to further reduce noise from site with vehicle alterations and barriers around selective fixed plant. A temporary barrier has been installed in the mill area in front of the ball mill and been effective in a 12dB reduction at source.

Ambient air quality monitoring stations (AAQMS) were installed within the community in 2013 to take continuous results for PM10 and PM2.5 as well as one week per month for metals and respirable crystalline silica (RCS). Previous monitoring was undertaken for dust deposition only. This more detailed investigation was undertaken initially for the permitting of the Big Hill Enhanced Development program and has now been incorporated into the site monitoring program. All results (apart from some offsite bushfire influenced events) have been below statutory criteria.

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There are currently no new permit requirements for the underground operations that would require any change in operation or effect the current environmental management. The TSF is current permitted for another lift (1.7m) that will hold the waste of the known Mineral resources/reserves. The TSF is being actively managed for detected seepage and is being regulated to manage any associated risks.

  20.2

MONITORING

Monitoring is undertaken on site for a range of environmental parameters as required by state and federal government regulations, the mining license and work plans. A report on monitoring data is presented to the ERC on a quarterly basis and an annual report is provided to the State regulators as a condition of the mining license. Table 20-1 lists environmental monitoring requirements for Stawell Gold Mines.

Component

Action Frequency

Air Quality

Dust deposition at 9 locations surrounding site. Monthly

Real time PM10 and PM2.5 within immediate community at 2 locations. Continuous

RCS and metals within immediate community at 2 locations. 1 week per month

Emissions from exhaust vents and kiln for odor and gases. Biannual.

Blasting

Surface vibration at 4 locations within community above working areas. Continuous

Airblast at 4 locations within community above working areas. Continuous

Noise

Noise monitoring at 4 closest sensitive receptors. Annual

Flora

Flora surveys of rehabilitation sites. Annual

 

Weed surveys of site. Quarterly

Fauna

Record of fauna spotted. Daily

 

Pest animal survey of site. Annual

Heritage

Survey of new areas As required

Air Emissions (including greenhouse gases).

Reporting through statutory reporting for the National Greenhouse and Energy reporting Scheme and National Pollutant Inventory. Annual

Water

Groundwater and surface water monitoring. Monthly, Quarterly and

 

Monitoring for any changes from background. Annually.

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Component

Action Frequency

Tailings Storage Facility

Monitoring of SWL, drainage flows, piezometers and deformation surveys. Monthly and Annually.

TABLE 20-1 ENVIRONMENTAL MONITORING REQUIREMENTS FOR STAWELL GOLD MINES

Dust deposition monitoring stations are located at nine locations around the site. Two AAQMS are located within the immediate community and monitor ambient dust emissions in accordance with the EPA protocols for environmental management. Monitoring is undertaken in accordance with the Australian standard. Monitoring of the afterburner inlet and outlet of the mill kiln and the four ventilation stacks is undertaken on a biannual basis by an independent emissions consultant and analysis performed by a NATA certified laboratory for gaseous emissions and odor.

Noise control is an integral part of site design and a number of mitigation measures are in place to reduce the impact of noise on local residents. As described in the site mining license, the noise limits applicable for the Stawell Gold Mines operations were derived from procedures outlined in the EPA State Environment Protection Policy (Control of Noise from Commerce, Industry & Trade) No. N-1. Monitoring is undertaken annually by an independent consultant. All blast events are monitored for ground vibration using monitors located at sensitive sites. A portable vibration monitor can be utilized if the need arises. All results of the air blast and ground vibration monitoring are recorded and maintained by Stawell Gold Mines. Flora and fauna surveys are conducted yearly by consultants in rehabilitation areas and surrounding the TSF. The rehabilitation areas are surveyed to ensure progress is meeting the planned end land use and the flora surrounding the TSF is monitored to ensure there are no impacts on vegetation.

Monitoring of surface waters and groundwater on site and off site is undertaken to ensure compliance with the State Environmental Planning Policies (SEPP, Waters of Victoria), the SEPP (Groundwater), Work Plan conditions and the mining license conditions. Stawell Gold Mines monitors 83 groundwater monitoring bores installed around the site with them being most densely installed around the TSF’s. Also monitored are all surface water catchments including surrounding farm dams and creeks.

  20.3

PERMITTING

Stawell Gold Mines underground operations are fully permitted with no outstanding permitting requirements for current Mineral Resources or Reserves.

In October 2014 the Big Hill Project EES was provided a recommendation from the Minister for Planning to the Minister for Energy and Resources that further considerations were required, predominately in the areas of air quality and public health. The Big Hill Project was modified to meet the concerns raised by the Minister for Planning and SGM has requested the Department of Economic Development, Job, Transport and Resources Victoria to review the changes and to provide the permitting pathway for the project to progress to a work plan.

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  20.4

COMMUNITY ENGAGEMENT

Stawell Gold Mines is closely associated with the Township of Stawell (to the south and east of the mine) and is bordered by farming land to the north and east, and state forest to the west. Stawell Gold Mines has been operating in Stawell since 1981 and is an integral part of the town. The mine is a large contributor to the regional economy and therefore has many stakeholders connected directly or indirectly to the operation. The expectation by the local community is that the Newmarket Gold Inc. will continue to operate in accordance with its license conditions and modern community standards and the local community expects and appreciates being informed of changes to the operations.

Stawell Gold Mines has taken a proactive approach to community relations and has endeavored to keep the community up to date with all changes occurring at the mine site and within their community. Community attitudes and expectations have been identified through a range of techniques including:

 

One on one discussions

 

Community surveys

 

Information sessions

 

Community meetings

 

Environmental Review Committee

 

Annual Open Day


  20.5

REHABILITATION AND CLOSURE

Stawell Gold Mines is regulated through the Mineral Resources (Sustainable Development) Act 1990 (MRSDA). It is a requirement of the MRSDA to have a rehabilitation plan and a rehabilitation bond held by the State government. Stawell Gold Mines is required to regularly review their closure plan and rehabilitation bond with the state regulators. Closure costs are derived using a combination of consultants, contractors and the regulator rehabilitation bond calculator. The current bond held by the state for rehabilitation is $4.803 million ($0.9 million for demolition and removal of surface infrastructure has been deferred from the bond by the state after an asset valuation for second hand and scrap of the existing infrastructure was shown to be greater than the demolition cost). The bond amount covers the decommissioning and rehabilitation of the surface and underground facilities including;

 

Demolition and removal of all surface infrastructure (plant, buildings, roads, services)

Capping and rehabilitation of tailings storage facilities to native bush (carting and spreading growth material and topsoil, structural works, surface contouring, drainage, seed and seedlings)

 

Rehabilitation of open pits (stabilization of pit walls, bunds/fencing, re-vegetation and signage)

 

Rehabilitation of underground workings (sealing of any underground accesses and ventilation shafts)

 

Rehabilitation of stockpiles/waste dumps (shaping, spreading, contouring and re-vegetation)

Decommissioning and Post closure costs (To cover the on-going assessment, monitoring and management of the site)

 

Contingency (Imposed by the regulator for unforeseen costs)

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The Stawell Gold Mines is a reasonably small compact site and there are very few opportunities to conduct progressive rehabilitation whilst it is in operation. Two closed TSF’s have been filled in and rehabilitated over the life of the mine along with other small areas around site.

The Reserve Tailings Dam has been rehabilitated into the home of the Stawell Clay Target Complex. In 2000 Stawell Gold Mines established a Tailings Experimental Research Facility (TERF) on TSF1 as a field trial to investigate different closure mechanisms for the tailings dams. The University of Melbourne, Curtin University and O’Kane Consultants have undertaken extensive research on this TERF to determine the performance of the cover design system on a number of aspects including:

  Acid Mine Drainage
  Ecotoxicology
  Soil stabilization
  Arsenic uptake,
  Net percolation and water balance, and
  Phyto-stabilization using trees.

The results of these trials have been used to finalize rehabilitation and create an end land use available to community activities.

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  21

CAPITAL AND OPERATING COSTS


  21.1

UNDERGROUND COSTS

Underground costs detailed in this section have been derived from the 2016 Stawell Gold Mines budget and include current costs for underground mining and processing.

  21.1.1

CAPITAL COSTS

The capital costs for Stawell Gold Mines in 2016 are presented below in Table 21-1 SGM 2016 CAPITAL COST SUMMARY. These costs are based upon budgeted works planned to be undertaken through the course of 2016.

Underground capital development planned is access development to the Federal Albion South Mineral Reserve area.

2016 Capital Costs

$,000’s Cost

Mobile Plant and Equipment

200

Mobile Plant and Equipment Under Finance

0

Processing Plant

1,016

Surface Infrastructure

345

UG Infrastructure

0

Land and Buildings

0

Underground Development

1,380

Total

2,941

TABLE 21-1 SGM 2016 CAPITAL COST SUMMARY

The capital costs for Stawell Gold Mines beyond 2016 budget period are expected to be similar to current values.

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  21.1.2

OPERATING COSTS

Operating costs for Stawell Gold Mines in 2016 are presented below inTable 21-2. These costs are based upon the budgeted quantity of works planned to be undertaken through the course of 2016.

Underground mining costs are inclusive of all development, production, haulage and associated power and running costs excluding Capital Development allowance in Table 21-1, above.

Geology costs include all underground geological programs including diamond drilling programs to further evaluate or convert target areas on the western flank of the Magdala basalt.

Surface trucking costs include all activities for haulage of LG Surface stockpile Mineral Reserves for processing.

2016 Operating Costs

$,000’s Cost

UG Mining

22,582

Geology

1,503

Surface Trucking

1,281

Processing

17,207

Administration

4,948

Total

48,072

TABLE 21-2 SGM 2016 OPERATING COST SUMMARY

Underground mining operating costs are indicated to remain similar for future years.

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  21.2

SURFACE COSTS

Surface mining costs are based on a Project Feasibility Study for the Big Hill Enhanced Development Project prepared by Dean Basile, MAusIMM (CP) and Stuart Hutchin, MAIG, MAusIMM dated June 2014 (the Big Hill Technical Report).

Operating and capital cost estimates were developed by Mining One in conjunction with Newmarket Gold for the Big Hill project. The operating costs used in the feasibility study were based on a contractor tendered mining rates. A shadow bid cost estimate using first principles was carried out to verify the tender rates used are reasonable for a project of this nature and size. Capital cost estimates include initial and sustaining expenditures and have been estimated and obtained from a variety of sources. They are detailed in the following sections.

  21.2.1

CAPITAL COSTS

Overall project capital requirements are assumed to be similar for both Contractor and Owner mining cost models. Owner mining cost models assume that equipment is either fully leased or on a dry/maintained hire agreement.

Capital costs estimates are based on both onsite set up requirements for the surface mining activities, environmental monitoring programs and removal and re-establishment of infrastructure items as per section 18.7.

Capital Cost Estimate Big Hill

$,000’s Cost

Permitting

 

Project Site Set Up

2,704

Environmental

7,229

Public Infrastructure

898

GWM Infrastructure

2,798

Communications

773

Mining Set Up

1,639

Mining Equipment

1,380

Operation Fixed Capital

187

Total

17,608

TABLE 21-3 CAPITAL COST ESTIMATE BIG HILL

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FIGURE 21-1 BIG HILL CAPITAL EXPENDITURE

Capital cost estimates were reviewed during 2015 are in line with modified project considerations.

  21.2.2

OPERATING COSTS

Operating costs for the Big Hill project were reviewed during 2015 with modified project schedules and material movements.

Operating cost estimates for the Big Hill project are based upon owner mining costs models with equipment leased for the duration of the project.

Operating Costs Estimate Big Hill $,000’s Cost
Surface Mining 1.24
Geology 0.11
Surface Trucking 0.39
Fixed & Labour 2.02
Lease & Hire 1.24
Total 4.99

TABLE 21-4 OPERATING COST ESTIMATE BIG HILL

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Operating cost estimates for the Big Hill project showing proportionment of cost drivers below in Figure 21-2.


FIGURE 21-2 OPERATING COST BREAK DOWN

Mining cost estimates are driven by productivity analysis on the mining schedule to determine cycle times and associated equipment and related labor requirements.

    21.2.2.1

LOAD AND HAUL OPERATING COSTS

Load and haulage cost estimates are based upon surface mining schedules as per Section 16.3.2. First principle productivities are calculated using excavator specifications and benchmarks from existing operations.

Table 21-5 outlines the productivity estimates for the load and haulage fleet proposed for the operation.

Excavator Productivity

Unit Value Value

Excavator Type

  EX1200 EX1900-6-BH

Truck Type

  Cat_777F Cat_777F

Digger Configuration

  Backhoe Backhoe

Material Detail

     

Dry Density

t/BCM 2.1 2.1

Moisture Content

% 2% 2%

Swell Factor

% 15% 15%

Wet Density Loose

t/m^3 1.86 1.86

Wet Bank Density

t/m^3 2.14 2.14

Shovel Details

     

Bucket Heaped Cap.

m^3 6.7 9.6

Fill Factor

% 85% 85%

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Excavator Productivity

Unit Value Value

Bucket Cap. Volume

BCM 5.0 7.1

Bucket Cap. Weight

t 14.7 19.2

Bucket Cap. Weight

BCM 6.9 9.0

Bucket Cap. Adopted

BCM 5.0 7.1

Truck Details

     

Tray Capacity

m^3 60.2 60.2

Truck Fill Factor

% 95% 95%

Volume Limit

BCM 49.7 49.7

Rated Payload

t 90.7 90.7

Assumed Overload

% 0 0

Adjusted Payload

t 90.7 90.7

Weight Limit

BCM 42 42

Adopted Capacity

BCM 42 42

Min. Bucket Fill

% 25% 25%

Calculated Passes Per load

  9 6

Calculated Passes Per load

(rounded) 9 6

Actual Truck Load

BCM 42 42

Actual Truck Load

t 90.7 90.7

Dump Time

min 1.2 1.2

Excavator Productivity

     

Cycle Time

sec 25 25

Efficiency Factor

% 100% 100%

1st Pass

sec 9 9

Truck Exchange

sec 38 38

Loading Time

min 4.12 2.87

Max. Productivity

BCM/OH 617 886

Effective Utilisation of operating hours

% 73% 76%

Productivity

BCM/OH 450 670

Productivity

t/OH 945 1,407

TABLE 21-5 OPERATING PRODUCTIVITY ESTIMATE BIG HILL

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The productivity estimates for each excavator is matched against the mining schedule to determine the required number of excavators. Figure 21-1 below illustrates the total excavator capacity for an EX1900 and EX1200 versus the mining schedule.


FIGURE 21-1 EXCAVATOR REQUIREMENTS BIG HILL

Overall excavator operating costs excluding operating labor for the project period equate to $5,313,000for a unit rate averaged across the project of $0.24/t mined and an additional $0.58/t of rehandle material for reconstruction purposes.

Haulage of ore and waste were modelled in TALPAC software to determine travel time and fuel consumption. Travel time and fuel consumptions were calculated for all haulage routes from each bench to the ROM or waste dump.

In addition, travel times were used to determine the productivity of each truck and thus truck requirements were matched to the mining schedule. Figure 21-2 below illustrates the haul truck requirements.

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FIGURE 21-2 HAUL TRUCK REQUIREMENTS BIG HILL

Overall haulage truck operating costs excluding operating labor for the project period equate to $11,881,000 for a unit rate averaged across the project of $0.39/t mined and an additional $1.64/t of rehandle material reconstruction purposes.

    21.2.2.2

ANCILLARY EQUIPMENT OPERATING COSTS

Ancillary equipment was selected in order to adequately support mining operations. The type and quantity selected over and above what would be required at based on experience with operations of a similar size and nature. Additional ancillary equipment costs are incurred for the project mainly in the increased use of additional bulldozers for excavation, stockpile and re-establishment purposes. Additional water cart and dust suppression equipment and labor allowances have also been added to the ancillary equipment schedule. Ancillary equipment requirements equate to $9,955,000 over the life of the project or a unit rate of $0.79/t mined and an additional $0.47/t of rehandle material for reconstruction purposes.

    21.2.2.3

DRILL AND BLAST OPERATING COSTS

Mining will predominantly be via rip and dig with minimal drill and blast activity scheduled. Due to the relatively small quantity of material expected to be drilled and fired, drill and blast operations will be carried out via a contractor as the quantity does not justify the capital cost.

Drill and blast contractor costs equate to $2,038,000 over the life of the project or $2.74/BCM for drill and blast zones.

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    21.2.2.4

EQUIPMENT LEASING AND HIRE

For the project major items of mobile equipment would be leased with additional equipment hired on an as needs basis. Lease rate calculations have been based on the duration of the project at 10% interest terms.

Leasing, insurance and hire costs for the duration of the project equate to $19,147,000 or a unit rate of $1.24/t mined and an additional $1.46/t of rehandle material for reconstruction purposes.

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  22

ECONOMIC ANALYSIS


  22.1

UNDERGROUND ECONOMIC ANALYSIS

Section 22 for underground mining has been excluded on the basis that the property is currently in production and there are no plans for material expansion of current production, other than in connection with the Big Hill Project.

  22.2

BIG HILL ECONOMIC ANALYSIS

Economic Analysis for the Big Hill Project are based upon Feasibility Cost Estimates compiled during 2014 with adjustment as necessary for Project modifications conducted to-date. Cashflow streams consist of Revenue, Capital Costs, Operating Costs and Taxes.

  22.2.1

PRODUCTION AND REVENUE

Production and revenue estimates for the Big Hill Project are presented below in Table 22-1.

Description

Unit Big Hill Project

Ore Production

Mt 2.73

Processing Recovery

% 90

Gold

koz 119

Revenue @ A$1,450/oz gold price

$M 172.9

TABLE 22-1 BIG HILL PRODUCTION AND REVENUE SUMMARY

  22.2.2

OPERATING COSTS

Operating costs are comprised of Mining, Waste Rehandle, Processing, Royalties and Administration with further detail provided in Section 16.3.2 and Section 21.2.2. Overall operating costs for the Project are presented below in Table 22-2.

Description

Unit Big Hill Project

Mining

$M 48.34

Waste Rehandle

$M 17.23

Processing and General Admin

$M 40.25

Royalties

$M 0.24

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Description

Unit Big Hill Project

Pre-tax

   

Undiscounted CashFlow

 $ 000 46,915

Discounted CashFlow @ 8%

 $ 000 37,262

IRR

% 121.4%

Payback Period

quarter 6.38

Post-tax

   

Undiscounted CashFlow

 $ 000 20,584

Discounted CashFlow @ 8%

 $ 000 20,007

IRR

% 74.5%

Payback Period

quarter 7.35

TABLE 22-3 BIG HILL CASHFLOW MODEL OUTPUTS SUMMARY

The quarterly cashflow, cash streams and operating costs of the Big Hill Project are depicted in Figure 22-2 and Figure 22-3 respectively.


FIGURE 22-2 BIG HILL CASHFLOW PRE-TAX NPV

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FIGURE 22-3 BIG HILL FOUR CASH STREAMS

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  23

ADJACENT PROPERTIES

There are no adjacent properties of significance to the Stawell Gold Mines.

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  24

OTHER RELEVANT DATA AND INFORMATION

The Stawell Gold Mines underground mining plan is based upon mining of both classified Mineral Resource and the associated Mineral Reserve and also a proportion of material that is from “remaining unclassified Mineral Resource” areas. This material can contribute up to and over 50% of the mined ounces in a given period. The remaining “unclassified Mineral Resource” areas consist of lower grade portions of the mineralization that has been uneconomic to mine in previous years of operation and is often in areas of older data and lower geological confidence.

In June 2014, Crocodile Gold Corp. released the results of a Feasibility Study for the Big Hill Enhanced Development Project located adjacent to the Stawell Gold Mine. The Feasibility Study had been completed in accordance with the CIM Definition Standards on Mineral Resources and Mineral Reserves referred to in the National Instrument 43-101.

Table 24-1 to Table 24-6 summarizes the findings of the Feasibility Study as of June 2014 on the Big Hill Deposit.

Key Project Elements.

Financial Analysis

Pre-Tax Post Tax

Gold Price*

A$1,415 A$1,415

Undiscounted Cash Flow (A$)(M)

49.2 30.3

NPV @ 8% Discount (A$)(M)

38.5 22.6

IRR

125.3% 79.1%

Payback Period (Years)

1.5 1.9

TABLE 24-1 BIG HILL FINANCIAL ANALYSIS RESULTS BASED ON A USD$1,225/OZ PRICE AND A US$:A$ EXCHANGE RATE OF 0.87

Capital Costs

Pre-production Capital (A$)(M)

A$11.99

Total Project Capital (A$(M)

A$19.60

TABLE 24-2 CAPITAL COST SUMMARY

Operating Costs

Mining Cost (A$/t ore)

A$4.83

Waste Rehandle (A$/m3 waste)

A$4.98

Processing Costs (A$/t ore)

A$14.75

Royalty (AUD$/oz)

A$2.00

TABLE 24-3 OPERATING COST SUMMARY

Unit Costs

Operating Cash Cost (A$/oz)

A$886

All-in Sustaining Cash Cost (A$/oz)

A$1035

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TABLE 24-4 UNIT COSTS SUMMARY

Operating Plan

Mining Duration (Years)

3.5 – 4

Mining (Days/Yr)

260

Landform Re-establishment (Years)

1-1.5

TABLE 24-5 OPERATING PLAN SUMMARY

Processing

Metallurgical Recovery

90 %

Recovered Grade (g/t Au)

1.5

Recovered Gold (oz)

131,000

TABLE 24-6 BIG HILL PROCESSING SUMMARY

The Feasibility Study confirmed the technical and economic viability of the Project and highlights the measures taken to achieve levels of compliance beyond that of conventional open pit mining in consideration of proximity to the Stawell community. Self-imposed modified practices have been adopted to mitigate environmental and social impacts, even if not economically optimal.

The Project is currently the subject of an Environmental Effects Statement and Ministerial Assessment. These modified work practices have been adopted in consideration for favorable permitting and works approval.

Substantial work has been done to confirm operational costs, technical support and social impacts. Based on a forecast gold price of A$1,415 per ounce, the Feasibility Study base case, pre-tax NPV (8%) is A$39M with an IRR of 125%. The most positive characteristic of the Project is its high operating margin; economic sensitivity analyses demonstrate the Project NPV to be resilient to downward movement in gold price and potential upward movement in costs. Table 24-7 and Table 24-8 highlight the Project’s ability to withstand potential market pressures and capacity to capitalize on opportunities.

Gold price sensitivity (Pre-tax)

A$/oz

1,300 1,350 1,400 1,450 1,500

NPV (A$M)

26.3 31.6 36.9 42.2 47.6

IRR(%)

97.7% 110.2% 121.9% 133.1% 143.7%

Pre-Tax Payback Period Undiscounted (Years)

1.7 1.6 1.6 1.5 1.5

TABLE 24-7 GOLD PRICE SENSITIVITY ANALAYSIS

Change in NPV Pre-tax (%)

-20 -10 0 10 20

Capital costs (A)($M)

42.7 40.6 38.5 36.4 34.3

Operating costs (A)($M)

56.7 47.6 38.5 29.4 20.3

TABLE 24-8 COST SENSITIVITY ANALYSIS

The Big Hill Project is part of an Australian tax-consolidation group, which has non-capital operating losses, which can be applied to reduce taxable income in future years. It is expected that a portion of these losses will be available during the Big Hill time frame which would effectively increase the after-tax value of the project.

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The total pre-production capital cost is estimated at A$11.99 million and include costs associated with site establishment, relocation of existing infrastructure and environmental impact mitigation. The total capital cost for the Big Hill Enhanced Development Project is estimated at A$19.6 million, which includes rehabilitation and end land use amenity consideration. Mine fleet requirements will be pursued through either leasing or dry hire arrangement.

Capital Cost Summary

Unit Cost (A$M)

Permitting

1.85

Project Site Set up

2.56

Public Infrastructure

0.90

Town Water Infrastructures

2.67

Communication and Fire Watch

2.05

Mining Set Up

1.62

Mining Equipment and Operation Fixed Capital

0.34

Pre-production Capital Costs

11.99

Environmental

7.61

Total Project Capital Costs

19.60

TABLE 24-9 CAPITAL COST SUMMARY AND PRE-PRODUCTION CAPITAL COSTS

Environmental Bonds will reflect incremental payment at key stages of the project and will also reflect bond reduction on completion of progressive rehabilitative works.

A comprehensive first principle mining cost model was developed to provide a shadow bid estimate and compared to the result of a formal tender process. This undertaking forms the basis of mine operating costs used in the Feasibility Study. Processing costs are based on actual costs as realized at Stawell Gold Mines for the treatment of Big Hill Enhanced Development Project ore types.

At the end of mining operations, re-handling and rehabilitation of the pits and waste dumps will be undertaken. Approximately 3.6 million cubic meters of material will be re-handled back into the pit voids as part of the rehabilitation. This will complete the progressive rehabilitation program of events.

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Operating Cost Summary

Unit Cost

Mining cost (A$/t rock)

4.83

Waste re-handle cost (A$/m3)

4.98

Processing & Administration cost (A$/t ore)

14.75

TABLE 24-10 OPERATING COSTS SUMMARY

Additional operating cost allowances have been incorporated to reflect the adoption of leading practice mine operations and were estimated in the order of more than A$10.6 million over the life of the project (Table 24-9 and Table 24-10). These initiatives are in addition to those environmental management initiatives included within capital costs, and are included within operating unit rates presented.

  24.1

BIG HILL MINERAL RESOURCE ESTIMATE

The Mineral Resource model used in the Feasibility Study was current as at March 2014, Table 24-11. Drill spacing over the considered Mineral Resource area is largely 20m x 25m, with drilling inclusive of additional RC and diamond drilling programs from 2008, 2012 and 2013. The estimate incorporates an updated geological interpretation, update of the Mineral Resource estimate, review and update of void models, utilization of a pit shell (constrained) at A$1,425 gold price and an in situ cut-off for reporting of 0.35 g/t Au. The Indicated Mineral Resources are inclusive of those Mineral Resources modified to produce the Ore Reserves.


Domain
Tonnes
(Mt)
Gold Grade
(g/t Au)
Ounces Gold
(,000’s)
Big Hill Measured - - -
Big Hill Indicated 3.0 1.7 160
Total (Measured and Indicated) 3.0 1.7 160
Inferred 0.2 1.2 7

TABLE 24-11 BIG HILL MINERAL RESOURCES ESTIMATE AS OF MARCH 2014

NOTES:

1.

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by National Instrument 43-101.

   
2.

Mineral Resources are inclusive of Mineral Reserves.

   
3.

Mineral Resources were estimated using the following parameters:

   
  a. Gold price of A$1,500/oz
   

b. Cut-off Grade applied for Big Hill Surface Mineral Resource is 0.44 g/t Au

   
4.

Surface Mineral Resource estimates were prepared by Justine Tracey, Senior Resource Geologist, Stawell Gold Mines. Ms Tracey is a member of the Australian Institute of Geoscientists and a Charted Professional member of the Australasian Institute of Mining and Metallurgy, and has over 12 years of relevant geological experience and is the Qualified Person for Mineral Resources under NI 43-101.

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5.

Ms. Tracey believes that the stated Mineral Resources is a realistic inventory of mineralization which, under the assumed technical, political, legal, environmental and economic development conditions, is economically extractable. If these conditions change then the Mineral Resources, either in whole or part, may not be economically extractable.

   
6.

The quantity and grade of the reported inferred Mineral Resources are uncertain in nature and there has been insufficient exploration to define the inferred Mineral Resources as indicated or measured Mineral Resources and it is uncertain if further exploration will result in upgrading them to an indicated or measured Mineral Resource category.

   
7.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

   
8.

Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

A series of dilution modelling routines were applied to the resource model to develop a model that would reflect the impact, on grade and tonnes, of practical mining constraints.

  24.2

BIG HILL MINERAL RESERVE ESTIMATE

The pit optimization and design were developed considering, detailed slope parameter analysis, potential infrastructure interaction, safe operating widths for the proposed equipment, business objectives, rehabilitation commitments, previous mining activities (both open pit and underground), minimization of dust and noise impact on the surrounding community and rehabilitative commitments. The Big Hill ore reserve as at March 2014 is detailed in Table 24-12.

Big Hill Enhanced Development Project
Reserve
Classification
tonnes
(Mt)
grade (g/t)
Au
ounces
(000's)
North Pit Probable 1.4 1.7 77
South Pit Probable 1.5 1.4 68
Total   2.9 1.5 145

TABLE 24-12 BIG HILL ORE RESERVES AS OF MARCH 2014

All Mineral Resources and Mineral Reserves have been estimated in accordance with the JORC Code and have been reconciled to CIM Standards as prescribed by NI 43-101.

2.

Mineral Reserves were estimated using the following economic parameters:


  a.

Gold price of A$1,450/oz

     
  b.

Cut-off Grade applied was 0.4 g/t Au


3.

Big Hill Surface Mineral Reserve estimates were prepared by Mining One personal under the guidance of Mark Edwards, General Manager Exploration Newmarket Gold. Mr Edwards is a member and Chartered Professional of the Australasian Institute of Mining and Metallurgy, has over 18 years of relevant mining experience and is the Qualified Person for Mineral Reserves under NI 43-101.

   
4.

Mineral Resources and Mineral Reserves are rounded to 1,000 tonnes, 0.01 g/t Au and 1,000 ounces. Minor discrepancies in summations may occur due to rounding.

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  24.3

GEOLOGY, MINING, METALLURGY AND PROCESSING

Underground mining operations at Stawell Gold Mines has been continuous since 1981 with the extraction of gold bearing ore from the Magdala and Golden Gift systems to a vertical depth of -1200mRL and -1600mRL respectively.

Big Hill is the up dip extension of the Magdala system, which has been historically mined from underground. It contains Basalt Contact mineralization, Central Lode mineralization and Stockwork mineralization, all typically seen in the Magdala System. Big Hill geology and mineralization can be broken into 4 main domains: Mariner’s, Allen’s, Iron Duke and Magdala Flank. All except Mariner’s and Allen’s are separated by faults.

The project proposes the mining of two open cut pits, the North Pit and the South Pit (Figure 24-1). The North Pit will be mined first and provide near immediate production contribution in light of outcropping ore zones. On completion of the North Pit, the pit will be filled and progressively rehabilitated as the South Pit is mined. The South Pit is effectively an extension to economic depth of the former “Davis Pit” which was mined in the late 80’s early 90’s. On completion of the South Pit, all waste rock will be returned to the pit voids and a full re-establishment and rehabilitation program undertaken with additional waste rock and top soil contributions from surface stockpiles.

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FIGURE 24-1 BIG HILL SURFACE PLAN

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The mine life is anticipated to be in the order of 3.5 to 4 years (excluding backfilling of the South Pit) using a mining fleet consisting of one 120 tonne and one 190 tonne excavator coupled with an average of five 90t trucks. Test work indicates material can be extracted using a rip and dig or free dig mining methods, with some drill and blast planned in the lower 30m of both pits.

An extensive metallurgical test work program has been completed for both the Big Hill North and South (Davis Extension) Pit areas. This test work was based on the leaching of a large number of RC drill intercepts during the period of 1998 to early 1999 and several new diamond drill holes completed in 2013.

Overall, the gold recovery from testing averaged 91.7%, with recovery consistent with depth, level of weathering of the rock and mine section.

The processing plant at Stawell has been in operation since 1984 and has a well-established flow path to accommodate the treatment of both oxidized and fresh ore types.

Ore from the Davis Open Cut has been previously processed at SGM with close to 500kt milled, with an overall gold recovery of 90% achieved. Testing of the mineralization in Big Hill has demonstrated that the mineralization elsewhere in the deposit will behave in a similar manner to the Davis Open Cut mineralization previously processed. A gold recovery of 90% is forecast.

The tailings facility at SGM has a current approved work plan to cover all additional requirements for the Project.

  24.4

ENVIRONMENTAL, PERMITTING, SOCIAL AND COMMUNITY CONSIDERATIONS

To achieve leading practice operations, which mitigate environmental and social impacts, the project taken into account a number of additional initiatives to conventional practice. Strategies associated with mining method modification (no blasting until the bottom portion of the Pits), alignment of ramps relative residential dwellings, dust mitigation measures, pit development, waste dump construction, noise modelling and monitoring, re-establishment of Big Hill, back filling of the South Pit, etc. has required the mine plan to delivered to a high level of detail..

Pit and stockpile benching configurations have been proposed that shield nearby residences from noise dust emissions, reducing some overall efficiency of operations. A staged mining approach is to be used in South Pit to reduce the active footprint and noise and dust emissions in the upper levels. Mining operations will be restricted to daylight hours only on Monday to Friday, with additional modified practices to mitigate social impact:

  Reduced use of blasting
  Sealing of haul roads, additional water carts and use of dust suppressants
  Staged clearing and grubbing
  Noise attenuation suppressant equipment
  Real time weather monitoring and predictive management practices

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Initiative Impact AUD $M Cost Implication
Mining, stockpile and reconstruction sequence and schedule. Decreased efficiency of mining, increased equipment moves, reduced equipment and schedule for the North Pit. Additional equipment for noise bunding and increased haulage distance for staged pit. A$4.9M
Reduced Blasting Rip and Dig requirement increased to defer drill and blast activity. Costing allowance in ancillary equipment for an extra D11 bulldozer over normal project allowances. A$1.9M
Additional watering capacity Additional water carts required for project duration, ancillary fleet contains two additional water carts to contractor requirements. A$2.2M
Landform Re- construction Additional bulldozer and compactor maintained during re-construction activities. Ancillary equipment for land form push-up and compaction. AU$1.6M
Total   A$10.6M

TABLE 24-13 ADDITIONAL OPERATING COSTS FOR COMMUNITY AND ENVIRONMENTAL CONSIDERATIONS

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  25

INTERPRETATION AND CONCLUSIONS

Operations continue at the Stawell Gold Mine with the processing facility at Stawell Gold Mine continuing to perform as expected. The cost structures are well understood through local experience through mining since operations commenced in 1981, and the subsequent operation down-size in 2012.

The Stawell Project has a significant production history and the site personnel have significant knowledge of the orebody and the mining process, which adds additional support to the methodologies used to estimate the Mineral Resources and Mineral Reserves.

Stawell Gold Mines has in place rigorous processes for the estimation of Mineral Resources and Mineral Reserves. These processes are managed on site and appropriate resources (personnel and drilling) are allocated to this function.

The conclusion of the Authors are that the Mineral Resources and Mineral Reserves as stated in this document are valid and supported by appropriate data collection, sampling, processing, interpretation and estimation methodologies and conform to NI 43-101 guidelines.

Exploration activities are incorporated into ongoing life of mine plans and it is the opinion of the Authors that additional exploration is warranted as per the descriptions included in Section 9 of this technical report.

There are no known risks or uncertainties, other than previously noted, (outside a significant reduction in gold price assumptions) which could reasonably be expected to affect the reliability or confidence in the exploration information, Mineral Resource or Mineral Reserve estimates or projected economic outcomes discussed in this technical report. The risk of the gold price assumptions has been mitigated by the use of an industry standard approach to estimating the price to be used in all estimations. Further work is required to maintain the understanding of the Mineral Resources and Mineral Reserves of the deposit, which will be completed as required to ensure the successful development of the project.

There is some uncertainty regarding the permitting process for the Pine Creek deposits, however the Newmarket team understand the obligations to obtaining these approvals and will work within the requirements of the mining department to gain approvals. There is some risk that this could delay the commencement of the operations but it is not of the opinion of the author that this approval will not be granted.

The Authors have made the following interpretations and conclusions:

The Stawell Mineral Resource increased in inventory over the past 12 months. A small increase in Indicated ounces in the Magdala Western Flank mineralisation does not reflect the extent of conversion during the year. This is a result of a short time frame between delivery of the mineral Resource and mine scheduling, thus material that was converted during the reporting period is already in production during the same reporting period.

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An increase to the Inferred Mineral Resource in the reporting year is result of targeted drilling in the unmined margins of the Magdala orebody (Federal Albion South, Below Scotchmans 250 and Upper South Fault 2) and exploration drilling on the Eastern Flank (Aurora B).

   

Prior to the reporting period the a significant portion of the Inferred Mineral Resource inventory underwent conversion to Indicated Mineral Resource in order to sustain Mine Operations. Drill programs through 2015 have enabled this material to be replaced through the course of the reporting period. Continued drilling and investigation into the Inferred resource areas in 2016 gives capacity for conversion to Indicated, and will provide increased understanding into the along strike and down plunge potential to the mineralisation.

   

 

Underground Mineral Reserve has matched depletion during the reporting period.

   

 

Big Hill Surface Mineral Reserve was reduced in line with the Adjusted Big Hill mining plan.



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  26

RECOMMENDATIONS

The key recommendations of the Principal Author, Justine Tracey are in respect of ongoing geological testing and mining evaluation of the Mineral Resources and exploration areas at Stawell Gold Mines.

Maintain the mine geological program as currently implemented to supplement currently available geological information in the preparation of detailed mine design. It is the opinion of the Authors that there remains sufficient prospectivity within the Stawell Gold Mines tenement to support the proposed programs and budgeted expenditure and it is recommended that ongoing evaluation of this potential be continued and re-evaluated as exploration results are received.

   

A detailed exploration program is presented in Section 9 of this report, including a proposed program for 2016, which it is recommended should be implemented.

   

Further investigation of the Eastern Flank mineralization should be undertaken, including a detailed structural study of the diamond drill core to understand the down plunge position of Aurora B and the potential of the mineralized shoots. This understanding will aid in the targeting and design of further diamond drilling campaigns with the intent to extend the Inferred Resource and infill and convert to Indicated Resources.

   

Undertake infill diamond drilling on the upper southern extents of priority lodes to confirm the assumptions of geological continuity inherent in the current Mineral Resource estimate. Undertake targeted resource definition drilling on the faulted extremities of the mineralization (above the Scotchmans Fault and below the South Fault).

   

Continue to build and improve geological models over the Magdala orebody where there is no current digital model to aid in targeting and aid geological understanding. Continued investigation of the defined depletion model and underground ground truthing coupled with a diligent review of paper records is expected to add to life of mine stocks.

   

Continue to review the performance of the Mineral Resource estimate through regular reconciliation between geological modelling, mining and the processing facility. The implementation of consistent underground channel sampling and digital capture of the results to assist with determination of wireframe extents will aid with the consideration and reconciliation of recoverable reserves.

   

Continuation of diligent QAQC programs. The outcomes of the 2016 Internal Laboratory Audit will need to be actioned in early 2016 and quarterly internal Laboratory inspections commenced.

     
 

Continuation of collection of density measurements and geometallurgical testwork.

   

 

Continue to further permitting applications for the Big Hill surface mining project.

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  27

REFERENCES

Arne, D.C., Bierlein, F.P., McNaughton, N., Wilson, C.J.L., Morland, V.J. (1998). Timing of gold mineralization in western and central Victoria, Australia: New constraints from SHRIMP II analysis of zircon grains from felsic intrusive rocks. Ore Geology Review, 13, 251-273.

Basile, D., Hutchins, S. (2014). NI 43-101Technical Report – Big Hill Enhanced Development Project at Stawell Gold Mine, Mineral Resources and Reserves for Crocodile Gold Corp. Prepared by Mining One Consultants.

Blythman, R. (2015). Upper South Fault 2 Geology and Resource Model Report. SGM Internal Report.

Cas, R.A.F. (1983). A review of the palaeogeographic and tectonic development of the Palaeozoic Lachlan Fold Belt of southeastern Australia. Geological Society of Australia, Special Publication, 10.

Crawford, A.J. (1988). The Cambrian system in Victoria. Geology of Victoria. J. G. Douglas, and Ferguson, J.A. Melbourne, Geology Society of Australia, Victoria Division Special Publications: 37-62.

Coventry, D., Llorca, J., Keeling, S., Haydon, M., Weymess, P., Brenchley, P., Ross, J., Ankerson, A (2012).Technical Report on Stawell Gold Mines, Victoria, Australia.

Coffey Environments (2008), Stawell Gold Mines – Stage 4(a) – Hydrogeological Assessment, Tailings Storage Facility No 2 Factual Report (Draft). 621.0/10/1

Doronila, A. (2006). Phytostabilisation of arsenic – rich sulphidic gold mine tailings in Victorian goldfields, Australia. Ph.D. Thesis submitted to the School of Botany, University of Melbourne. Melbourne, Australia.

Dowsley, K. (2001). Characterization of a partially oxidized sulphidic tailings dam. Honors Thesis (unpublished). School of Earth Sciences. University of Melbourne.

Foster D.A., Gray, D.R., Kwak, T.A.P., Bucher, M. (1998). Chronology and tectonic framework of turbidite hosted gold deposits in western Lachlan Fold Belt, Victoria: 40Ar-39Ar results: Ore Geology Review, 13, 229-250

Fredericksen, D., Miller, G., Dincer, T. (2008). Technical Report on Stawell Gold Mines, Victoria, Australia.

Gane, M.J. (1998) Gold mineralization within the basalt contact ore zones, Magdala Mine, Stawell, Victoria. Masters Thesis (unpublished). The University of Melbourne, 208p

Gedge, L. (1997). The relationships between structure, gold mineralization and intrusive events at Stawell, Victoria. Unpublished thesis, Melbourne, Australia, The University of Melbourne, 95p

Heard, S. (2013). Mariners Underground Geology and Resource Model Report. SGM Internal Report.

Jupp, B. (2003). Hydrothermal alteration and lithogeochemistry of the Kewell and Wallup prospects and their comparison with the Magdala gold, Stawell, Victoria. Unpublished thesis, Melbourne, Australia, The University of Melbourne, 193p

Kaufman, A. (2003). The volcano-sedimentary and structural evolution of the Wildwood prospect, Western Lachlan Orogen. Unpublished thesis, Melbourne, Australia, The University of Melbourne, 85p

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LanePiper. (2008). 207119 Report 01.2 Environmental Audit Report Stawell Gold Mine Leviathan Rd, Stawell Vic, January 2008

Lllorca, J.P., Schunke, N., (2012). Crocodile Gold Corp. NI 43-101 Report Stawell Gold Mines, Victoria, Australia.

Mapani, S.E., Wilson, C.J.L. (1994). Structural evolution and gold mineralization in the Scotchmans Fault Zone, Magdala Gold Mine, Stawell, Western Victoria, Australia. Economic Geology, 89, 566-583

Miller, J.McL., Wilson, C.J.L, Dugdale, L.J. (2006). Stawell gold deposit: a key to unravelling Cambrian to Early Devonian structural evolution of the western Victorian goldfields, Australian Journal of Earth Sciences, 53, 677-695

Miller, J.McL., Wilson, C.J.L. (2004a). Structural Analysis of faults related to a heterogeneous stress history: reconstruction of a dismembered gold deposit, Stawell, western Lachlan Fold Belt, Australia. Journal of Structural Geology, 26, 1231-1256

Miller, J.McL., Wilson, C.J.L. (2004b). Stress controls on intrusion-related gold lodes: Wonga Gold Mine, Stawell, Western Lachlan Fold Belt, southeastern Australia. Economic Geology, 99, 941-963

Miller, J. McL., Wilson, C.J.L. (2002). The Magdala Lode System, Stawell, Southeastern Australia: Structural style and relationship to gold mineralization across the western Lachlan Fold Belt. Economic Geology, 97, 325-349

NQ, Groundwater & Environment (2011). Modelling of physical groundwater control options and cost estimates for ozone remediation at Stawell Gold Mines, July 2011.

NQ, Groundwater & Environment (2012). Review of Conceptual Groundwater Model with reference to the detection of SCN in Bore PS585.

Oldmeadow, D. (2008). Geochemical evolution of experimental tailings rehabilitations systems at Stawell Gold Mine, Victoria, Australia: implications for the use of thin composite cover in storage of sulphidic gold mine tailings. Thesis, Curtin University of Technology.

Pritchard, E.G. (2001). The Magdala Basalt in the east exploration decline, Magdala Gold Mine, Stawell Victoria: Petrography, alteration paragenesis and structural style. School of Earth Sciences, The University of Melbourne, 68p

Rockwater. (2010). Summary report on hydrogeological assessment and modelling of possible seepage from tailings storage facility No2. 621.0/101, March 2010.

Robinson, J.A., Wilson, C.J.L, Rawling T.J. (2006). Numerical modelling of an evolving gold-lode system: structural and lithological controls on ore-shoot formation in the Magdala Gold Mine, western Victoria, Australian Journal of Earth Sciences, 54, 799 – 823.

Robinson, J.A. (2005). Nature of the Mineralized (ore-shoot) environment within the Magdala Gold Deposit, western Lachlan Fold Belt, Australia, Unpublished thesis, Melbourne, Australia, The University of Melbourne, 373p

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Schaubs, P.M., Wilson, C.J.L. (2002). The relative roles of folding and faulting in controlling gold mineralization along the Deborah anticline, Bendigo, Victoria, Australia. Economic Geology, 97, 351-370

Squire, R.J., (2004). Stawell Au Deposit – ARC Linkage Project: June 2004 progress report, Unpublished

Squire, R.J., Wilson, C.J.L. (2005). Tectonic responses to super-continent formation: correlation of Cambrian geological events along proto-Pacific margin of East Gondwana. Journal of the Geological Society, London, 162, 749-761 Mineral Resources and Reserves.

Tracey, J (2012). Mid Magdala Geology and Resource Model Report. SGM Internal Report.

Tracey, J (2013). Big Hill Davis Geology and Resource Model Report. SGM Internal Report.

Tracey, J (2015). Below Scotchmans 250 Geology and Resource Model Report. SGM Internal Report.

Tracey, J (2015). Federal Albion South Geology and Resource Model Report. SGM Internal Report.Tracey, J., Blythman, R. (2014). Magdala S6000 Geology and Resource Model Report. SGM Internal Report.

Tracey, J., Chapman, W. (2014). Technical Report on Stawell Gold Mines, Victoria, Australia. Vandenberg, A.H.M., Willman, C.E., Maher, S., Simons, B.A., Cayley, R.A., Taylor, D.H., Morland, V.J., More, D.H., and Radojkovic A. (2000). The Tasman Fold Belt System in Victoria, Geological Survey of Victoria Special Publication

Vann, J, Jackson, S, and Bertoli, O. (2003). Quantitative Kriging Neighborhood Analysis for the Mining Geologist — A Description of the Method with worked case examples, in Proceedings of the, 5th International Mining Geology Conference (Australasian Institute on Mining and Metallurgy, Melbourne)

Watchorn, R.B., Wilson, C.J.L (1989). Structural setting of the gold mineralization at Stawell, Victoria, Australia. Economic Geology Monographs, 6, 292-309

Wilson, C.J.L., Will, T.M., Cayley, R.A., Chen, S. (1992). Geologic framework and tectonic evolution in Western Victoria, Australia. Tectonophysics, 214, 93-127

Xu, G., Powell, R., Wilson, C.J.L., Will, T.M. (1994). Contact metamorphism around the Stawell granite, Victoria, Australia. Journal of Metamorphic Geology, 12, 609-624

Websites used http://www.jorc.org

http://www.kitco.com/

http://www.vistagold.com/

 http://www.abs.gov.au

http://web.cim.org/UserFiles/File/CIM_DEFINITON_STANDARDS_Nov_2010.pdf

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  28

SIGNATURE PAGE

CERTIFICATE OF QUALIFIED PERSON

I, Justine Tracey, MAusIMM (CP), do hereby certify that:

1.

I am Senior Resource Geologist of Newmarket Gold Inc., with offices at:

Stawell Gold Mines
Leviathan Road
Stawell, 3380

Telephone: 61-3-5358-1022

Email: Justine.Tracey@ newmarketgoldinc.com

2.

I graduated with a Bachelor of Science (Geology) degree with H1 Honors from Melbourne University in 1998.

   
3.

I am a Member of the Australasian Institute of Mining & Metallurgy (MAusIMM - CP) – Membership No. 318313.

   
4.

I have worked as both mine and exploration geologist since graduation and have over 14 years’ experience in Australian gold deposits. I have worked as a Resource Geologist at Stawell Gold Mines for the past 6 years.

   
5.

I have read the definition of “Qualified Person” set out in National Instrument 43-101 (NI 43- 101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

   
6.

I am responsible for Sections 1-14, 17, 19, 20 and 23-28 of the Technical Report titled “REPORT ON THE MINERAL RESOURCES & RESERVES OF THE STAWELL GOLD MINE IN VICTORIA, AUSTRALIA FOR NEWMARKET GOLD INC.” effective December 31, 2015 and dated March 16, 2016 (the Technical Report).

   
7.

I work directly for Newmarket Gold Inc. at the Stawell Gold Mine and authored the previous NI43 101 Technical Report. I regularly visit the Stawell underground mine as required in my role as Senior Resource Geologist.

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8.

I am not independent of Newmarket Gold Inc. pursuant to NI 43-101, applying the test set out in Section 1.5 of NI 43-101.

   
9.

As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

   
10.

I have read NI 43- 101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 16th Day of March 2016

 

Justine Tracey, B.Sc.H (Geology), MAIG, MAusIMM (CP),

SENIOR RESOURCE GEOLOGIST

NEWMARKET GOLD AUSTRALIAN OPERATIONS



Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

CERTIFICATE OF QUALIFIED PERSON

I, Wayne Chapman, MAusIMM (CP), do hereby certify that:

1.

I am Mine Technical Manager of Newmarket Gold Inc., at the Stawell Gold Mines with offices at:

Stawell Gold Mines
Leviathan Road
Stawell, 3380

Telephone: 61-3-5358-9354

Email: wayne.chapman@newmarketgoldinc.com

2.

I graduated with a Bachelor of Engineering (Mining) degree from the University of Ballarat in 2000.

   
3.

I am a Member and Chartered Engineer of the Australasian Institute of Mining & Metallurgy (MAusIMM (CP)) – Membership No. 314911.

   
4.

I have worked for more than 16 years in underground mining including more than 11 years in gold mining operations. I have worked at a variety of Newmarket Gold Operations for the previous 9 years.

   
5.

I have read the definition of “Qualified Person” set out in National Instrument 43-101 (NI 43- 101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

   
6.

I am responsible for Sections 1-3, 15.1-15.2, 15.4, 16.1- 16.2, 18, 21.1, 22.1, 25-27 and 28 of the Technical Report titled “REPORT ON THE MINERAL RESOURCES & RESERVES OF THE STAWELL GOLD MINE IN VICTORIA, AUSTRALIA FOR NEWMARKET GOLD INC.” effective December 31, 2015 and dated March 16, 2015 (the Technical Report).

   
7.

I work directly for Newmarket Gold Inc. at the Stawell Gold Mine and authored the previous NI43 101 Technical Report. I regularly visit the Stawell underground mine as required in my role as Mine Technical Manager.

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8.

I am not independent of Newmarket Gold Inc. pursuant to NI 43-101, applying the test set out in Section 1.5 of NI 43-101.

   
9.

As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

   
10.

I have read National Instrument 43-101 and Form 43-101F1, and this Technical Report has been prepared in compliance with that instrument and form.

   
11.

The effective date of this Technical Report is 31st December 2015.

Dated this 16th Day of March 2016

 

Wayne Chapman, B.Eng.(Mining), MAusIMM (CP)

Mine Technical Manager

NEWMARKET GOLD AUSTRALIAN OPERATIONS



Technical Report
December 2015
Newmarket Gold
Stawell Gold Mines

CERTIFICATE OF QUALIFIED PERSON

I, Mark Edwards, BSc. MAusIMM (CP), MAIG, MAICD do hereby certify that:

1.

I am the General Manager of Exploration for Newmarket Gold Inc. (Newmarket Gold), with offices at:

1/48 Smith Street
Darwin,
Northern Territory, 0800
Australia

Telephone: 61-8-8982-4444

Email: medwards@crocgold.com

2.

I graduated with a Bachelor of Science (Geology) degree from Flinders University in 1997.

   
3.

I graduated with a Bachelor of Science (Geology) Honors Degree from the University of Tasmania in 1998.

   
4.

I am a Member of the Australasian Institute of Mining & Metallurgy (MAusIMM - CP) – Membership No. 220787.

   
5.

I am a Member of the Australian Institute of Geoscientists (MAIG) – Membership No. 3655

   
6.

I have worked as a geologist since graduation and have over 18 years’ experience in the exploration, development and mining of mineral properties in, Australia and Botswana and am familiar with and have visited a variety of styles of mineral deposits worldwide, with particular emphasis on precious metals.

   
7.

I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

   
8.

I am responsible for Sections 1-3, 15.3, 16.3, 18, 21.2, 22.2, 24-27 and 28 of the Technical Report titled “REPORT ON THE MINERAL RESOURCES & RESERVES OF THE STAWELL GOLD MINE IN VICTORIA, AUSTRALIA FOR NEWMARKET GOLD INC.” effective December 31, 2015 and dated March 16, 2015 (the Technical Report).

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9.

I work directly for Newmarket Gold Inc. in the region and have regular visits to site as required in my role as General Manager of Exploration. I regularly work on site and I last visited the site in February 2016.

   
10.

I am not independent of Newmarket Gold Inc. applying the test set out in Section 1.5 of NI 43- 101.

   
11.

As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

   
12.

I have read NI 43-101 and Form 43- 101F1, and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 16th Day of March, 2016.


Mark Edwards, B.Sc., MAusIMM (CP), MAIG, MAICD GENERAL MANAGER .

NEWMARKET GOLD INC

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Kirkland Lake Gold Ltd. - Exhibit 99.25 - Filed by newsfilecorp.com

Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

REPORT ON THE
MINERAL RESOURCES & MINERAL RESERVES
OF THE
NORTHERN TERRITORY OPERATIONS
In the Northern Territory, Australia
Prepared for
NEWMARKET GOLD INC.
 
Effective Date December 31, 2015
Dated March 21, 2016

                                


- i -

This technical report has been prepared as a National Instrument 43-101 Technical Report, as prescribed in Canadian Securities Administrators’ National Instrument 43-101, Standards of Disclosure for Mineral Projects (NI 43-101) for Newmarket Gold Inc.). The data, information, estimates, conclusions and recommendations contained herein, as prepared and presented by the Authors, are consistent with

  i)

information available at the time of preparation;

  ii)

data supplied by outside sources, which has been verified by the authors as applicable; and

  iii)

the assumptions, conditions and qualifications set forth in this technical report.

CAUTIONARY NOTE WITH RESPECT TO FORWARD LOOKING INFORMATION

This document contains “forward-looking information” as defined in applicable securities laws. Forward looking information includes, but is not limited to, statements with respect to the future production, costs and expenses of the project; the other economic parameters of the project, as set out in this technical report, including; the success and continuation of exploration activities, including drilling; estimates of mineral reserves and mineral resources; the future price of gold; government regulations and permitting timelines; requirements for additional capital; environmental risks; and general business and economic conditions. Often, but not always, forward-looking information can be identified by the use of words such as “plans”, “expects”, “is expected”, “budget”, “scheduled”, “estimates”, “continues”, “forecasts”, “projects”, “predicts”, “intends”, “anticipates” or “believes”, or variations of, or the negatives of, such words and phrases, or statements that certain actions, events or results “may”, “could”, “would”, “should”, “might” or “will” be taken, occur or be achieved. Forward-looking information involves known and unknown risks, uncertainties and other factors which may cause the actual results, performance or achievements to be materially different from any of the future results, performance or achievements expressed or implied by the forward-looking information. These risks, uncertainties and other factors include, but are not limited to, the assumptions underlying the production estimates not being realized, decrease of future gold prices, cost of labour, supplies, fuel and equipment rising, the availability of financing on attractive terms, actual results of current exploration, changes in project parameters, exchange rate fluctuations, delays and costs inherent to consulting and accommodating rights of local communities, title risks, regulatory risks and uncertainties with respect to obtaining necessary permits or delays in obtaining same, and other risks involved in the gold production, development and exploration industry, as well as those risk factors discussed in Newmarket Gold Inc.’s latest Annual Information Form and its other SEDAR filings from time to time. Forward-looking information is based on a number of assumptions which may prove to be incorrect, including, but not limited to, the availability of financing for Newmarket Gold Inc.’s production, development and exploration activities; the timelines for Newmarket Gold Inc.’s exploration and development activities on the property; the availability of certain consumables and services; assumptions made in mineral resource and mineral reserve estimates, including geological interpretation grade, recovery rates, price assumption, and operational costs; and general business and economic conditions. All forward-looking information herein is qualified by this cautionary statement. Accordingly, readers should not place undue reliance on forward-looking information. Newmarket Gold Inc. and the authors of this technical report undertake no obligation to update publicly or otherwise revise any forward-looking information whether as a result of new information or future events or otherwise, except as may be required by applicable law.

NON-IFRS MEASURES
This technical report contains certain non-International Financial Reporting Standards measures. Such measures have non standardized meaning under International Financial Reporting Standards and may not be comparable to similar measures used by other issuers. 

i 


- ii -

TABLE OF CONTENTS

1 Executive Summary 1
  1.1 INTRODUCTION 1
  1.2 PROPERTY DESCRIPTION AND LOCATION 1
  1.3 GEOLOGY & MINERALIZATION 3
  1.4 EXPLORATION, DEVELOPMENT AND OPERATIONS 4
  1.5 MINERAL RESOURCES AND MINERAL RESERVES 6
  1.6 CONCLUSIONS AND RECOMMENDATIONS 10
     
2 Introduction and Terms of Reference 15
  2.1 INTRODUCTION 15
  2.2 SCOPE OF WORK 16
  2.3 AUTHORS, QUALIFICATIONS AND RESPONSIBILITIES 16
  2.4 DEFINITIONS 17
  2.5 MINERAL RESOURCE AND MINERAL RESERVE DEFINITIONS 19
     
3 Reliance on other Experts and Disclaimer 25
  3.1 LEGAL ISSUES – AGREEMENTS, LAND TENURE, SURFACE RIGHTS, ACCESS & PERMITS 25
  3.2 HISTORICAL INFORMATION 26
  3.3 ENVIRONMENTAL ISSUES 26
     
4 Property Description and Location 27
  4.1 LOCATION 27
  4.2 MINERAL RIGHTS, MINING LAWS AND REGULATIONS 28
  4.3 ADMINISTRATION 32
  4.4 MINERAL TENURE 32
  4.5 AGREEMENTS 38
  4.6 SURFACE RIGHTS – LAND ACCESS 40
  4.7 OPERATING AUTHORIZATIONS 41
  4.8 MISCELLANEOUS LICENSES & ACCESS 42
  4.9 NATIVE TITLE 42
  4.10 ROYALTIES 46
  4.11 ENVIRONMENTAL MANAGEMENT PLAN 53
  4.12 WASTE DISCHARGE LICENSES 59
     
5 Accessibility, Climate, Local Resources, Infrasture and Physiography 60
  5.1 TOPOGRAPHY 60
  5.2 ACCESS 61
  5.3 CLIMATE AND VEGETATION 61
  5.4 LOCAL RESOURCES AND INFRASTRUCTURE 63
  5.5 POWER 63
  5.6 WATER 64
  5.7 COMMUNICATIONS 64
  5.8 MINING PERSONNEL 64
  5.9 ACCOMMODATION 65
  5.10 PROCESSING FACILITIES 65
     
6 History 66
  6.1 COSMO MINE AND SURROUNDING AREAS 67
  6.2 UNION REEFS AREA 70
  6.3 PINE CREEK GOLD PROJECT 75
     
7 Geological Setting and Mineralization 83
  7.1 REGIONAL GEOLOGY 83
  7.2 LOCAL COSMO MINE GEOLOGY 87
  7.3 UNION REEFS GEOLOGY 96
  7.4 PINE CREEK GEOLOGY 107
  7.5 BURNSIDE GEOLOGY 118
  7.6 MINERALIZATION 128
  7.7 DEPOSIT DIMENSIONS 143
     
8 Deposit Types 144
  8.1 MINERALIZATION DEPOSIT MODELS 144
  8.2 STRUCTURAL MODELS 149

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  8.3 COSMO MINE MODELS 153
  8.4 UNION REEFS MODELS 159
  8.5 PINE CREEK MODELS 160
  8.6 GOLD-URANIUM MINERALIZATION 160
  8.7 POLYMETALLIC DEPOSITS 161
     
9 Exploration 163
  9.1 COSMO EXPLORATION 163
  9.2 EXPLORATION PLANS FOR 2016 178
  9.3 UNION REEFS AREA 180
  9.4 PINE CREEK EXPLORATION 193
  9.5 BURNSIDE AREA EXPLORATION 195
     
10 Drilling 229
  10.1 COSMO MINE DRILLING 229
  10.2 UNION REEFS DRILLING 236
  10.3 PINE CREEK AREA 244
  10.4 BURNSIDE AREA 248
     
11 Sample Preparation, analysis and security 254
  11.1 REVERSE CIRCULATION DRILLING SAMPLING 254
  11.2 DIAMOND DRILLING SAMPLING 255
  11.3 COSMO MINE FACE SAMPLING PROCEDURE 257
  11.4 SAMPLING PREPARATION 258
  11.5 SAMPLE SECURITY 259
  11.6 QUALITY ASSURANCE/ QUALITY CONTROL 262
     
12 Data Verification 300
     
13 Mineral Processing and Metallurgical Testing 301
  13.1 UNION REEFS PROCESSING FACILITY 301
  13.2 METALLURGY 303
     
14 Mineral Resource Estimations 327
  14.1 INTRODUCTION 327
  14.2 COSMO MINE MINERAL RESOURCE 329
  14.3 UNION REEFS DEPOSIT 355
  14.4 PINE CREEK DEPOSITS 432
  14.5 BURNSIDE AREA 455
     
15 Mineral Reserves 508
  15.1 COSMO MINE 508
  15.2 UNION REEFS UNDERGROUND - PROSPECT 509
  15.3 UNION REEFS OPEN PIT – ESMERALDA 510
  15.4 PINE CREEK OPEN PITS 511
  15.5 CONCLUSION ON MINERAL RESERVES 512
     
16 Mining Methods 513
  16.1 COSMO MINE 513
  16.2 UNION REEFS UNEDERGROUND (PROSPECT) 528
  16.3 UNION REEFS OPEN PIT (ESMERALDA) 537
  16.4 PINE CREEK OPEN PITS 544
     
17 Recovery 553
  17.1 UNION REEFS GOLD PLANT 553
  17.2 UNION REEFS PLANT OPERATIONS 555
     
18 Project Infrastructure 560
  18.1 INTRODUCTION 560
  18.2 COSMO MINE 560
  18.3 UNION REEFS UNDERGROUND (PROSPECT) 560
  18.4 UNION REEFS OPEN PIT (ESMERALDA) 561
  18.5 PINE CREEK OPEN PITS 561
     
19 Market Studies and Contracts 563
  19.1 MARKETS 563

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  19.2 GOLD PRICE 563
  19.3 MATERIAL CONTRACTS 563
     
20 Environmental Studies, Permitting and Social or Community Impact            566
  20.1 NOTICE OF INTENT (NOI) 566
  20.2 ENVIRONMENTAL IMPACT ASSESSMENT (EIA) 567
  20.3 PUBLIC ENVIRONMENTAL REPORT (PER) 567
  20.4 NAF/PAF ENVIRONMENTAL TEST WORK PROCEDURE: COSMO MINE 569
  20.5 ENVIRONMENTAL ISSUES & LIABILITIES 571
  20.6 COMMUNITY CONSULTATION 578
  20.7 MINE CLOSURE REQUIREMENTS 578
  20.8 COMMENTS ON ENVIRONMENTAL ISSUES AND LIABILITIES 581
     
21 Capital and Operating Costs            582
  21.1 COSMO MINE 582
  21.2 UNION REEFS UNDERGROUND (PROSPECT) 583
  21.3 UNION REEFS OPEN PIT (ESMERLADA) 584
  21.4 PINE CREEK OPEN PITS 585
     
22 Economic Analysis            587
  22.1 GENERAL COMMENTS 587
  22.2 SCHEDULING STRATEGY AND ASSUMED OPERATING PARAMETERS 587
  22.3 NET PRESENT VALUE AND INTERNAL RATE OF RETURN 590
  22.4 CASH FLOWS 590
  22.5 SENSITIVITIES 594
     
23 Adjacent Properties            596
  23.1 NON-NEWMARKET GOLD DEPOSITS 596
     
24 Other Relevant Data and Information            604
     
25 Interpretation and Conclusions            605
     
26 Recommendations            607
  26.1 COSMO MINE 607
  26.2 UNION REEFS 608
  26.3 PINE CREEK 609
  26.4 BURNSIDE AREA 610
     
27 References            611
     
28 Signature Page            618

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TABLES PAGE
TABLE 1- 1 SUMMARY OF MINERAL TITLES FOR NEWMARKET GOLD NT OPERATIONS  2
TABLE 1- 2 MINERAL RESOURCES FOR NT OPERATIONS, AS AT DECEMBER 31, 2015  6
TABLE 1- 3 COSMO MINE MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015  7
TABLE 1- 4 MINERAL RESERVE CLASSIFICATION PROSPECT DEPOSIT UNDERGROUND AS AT DECEMBER 31, 2015  8
TABLE 1- 5 MINERAL RESERVE CLASSIFICATION ESMERALDA OPEN PIT AS AT DECEMBER 31, 2015  9
TABLE 1- 6 MINERAL RESERVE CLASSIFICATION FOR PINE CREEK AS AT DECEMBER 31, 2015  9
TABLE 1- 7 PROPOSED EXPLORATION PROGRAMS FOR COSMO MINE FOR 2016. 11
TABLE 2- 1 DEFINITIONS 19
TABLE 3- 1 SITE EXPERTS WHO CONTRIBUTED TO THE TECHNICAL REPORTS 25
TABLE 4- 1 SUMMARY OF MINERAL TITLES NEWMARKET GOLD NT OPERATIONS 27
TABLE 4- 2 SUMMARY OF MINERAL TITLES BURNSIDE (* MINERAL LEASES ARE INCLUDED IN EXPLORATION LICENSES) 33
TABLE 4- 3 SUMMARY OF MINERAL TITLES - UNION REEFS (* MINERAL LEASES ARE INCLUDED IN EXPLORATION LICENSES) 34
TABLE 4- 4 SUMMARY OF MINERAL TITLES PINE CREEK (* MINERAL LEASES ARE INCLUDED IN EXPLORATION LICENSES) 35
TABLE 4- 5 SUMMARY OF MINERAL TITLES FOR NT OPERATIONS OUTSIDE BURNSIDE, UNION REEFS AND PINE CREEK (* MINERAL LEASES ARE INCLUDED IN EXPLORATION LICENSES) 36
TABLE 4- 6 SUMMARY, MINERAL TITLES, NORTHERN TERRITORY, AUSTRALIA 37
TABLE 4- 7 LIST OF UNION REEFS ROYALTY’S CURRENTLY REQUIRED BY NEWMARKET GOLD 48
TABLE 4- 8 LIST OF PINE CREEK ROYALTIES CURRENTLY REQUIRED BY NEWMARKET GOLD 49
TABLE 4- 9 LIST OF ALL ROYALTYIES CURRENTLY REQUIRED BY NEWMARKET GOLD 52
TABLE 4- 10 NEWMARKET GOLD PERFORMANCE BONDS – 2016 54
TABLE 6- 1 HISTORICAL GOLD PRODUCTION – PINE CREEK OROGEN 66
TABLE 6- 2 ESTIMATED HISTORICAL GOLD MINED. COSMO HOWLEY GOLD PROJECT 67
TABLE 6- 3 SUMMARY OF HISTORIC OWNERSHIP OF COSMO HOWLEY MINING AREA 69
TABLE 6- 4 RECONCILIATION FIGURES FOR CROCODILE GOLD/NEWMARKET GOLD MILLING - 2009-2015 69
TABLE 6- 5 ESTIMATED HISTORIC GOLD PRODUCTION PINE CREEK REGION 1985-2007 72
TABLE 6-6 HISTORIC GRADE COMPARISON OF PROSPECT DEPOSIT MAIN LODE AT VARIOUS AU CUT- OFF GRADES (NB1) 74
TABLE 7- 1 BROCKS CREEK LODE TYPES AND GRADES 127
TABLE 7- 2 DEPOSIT DIMENSIONS 143
TABLE 8- 1 PINE CREEK OROGEN MINERALIZATION MODELS 148
TABLE 9-1 LANTERN TARGET DRILL INTERSECTION ASSAY RESULTS 175
TABLE 9- 2 ROCK CHIP SAMPLING INFORMATION FOR ELIZABETH 182
TABLE 9- 3 ROCK CHIP SAMPLING ANALYTICAL RESULTS FOR ELIZABETH 182
TABLE 9- 4: 2014 ESMERALDA GRAB SAMPLE RESULTS AU G/T 191
TABLE 9- 5 MT BONNIE EAST GRAB SAMPLE ICP RESULTS 199
TABLE 9- 6 MT ELLISON EAST ROCK CHIP AU RESULTS IN PPM 205
TABLE 9- 7 MODEL PARAMETERS FOR CONDUCTIVE PLATES BLT_021 AND BLT_022. 210
TABLE 9- 8 MODEL PARAMETERS FOR CONDUCTIVE PLATES BLT_020 AND BLT_026. 213
TABLE 9- 9 SNAKEBITE - SOIL SAMPLING CORRELATION MATRIX 219
TABLE 9- 10 NORTH CULLEN - RESULTS OF ROCK CHIP SAMPLING, AU G/T 222
TABLE 10-1 DIAMOND DRILL STATISTICS FOR THE COSMO MINE 230
TABLE 10-2 RC DRILL STATISTICS FOR THE COSMO MINE 233
TABLE 10-3 HISTORIC DRILLING BY COMPANY – DIAMOND DRILLING 235
TABLE 10-4 HISTORIC DRILLING BY COMPANY – RC DRILLING 235
TABLE 10-5 COSMO MINE HISTORIC DRILLING BY YEAR – DIAMOND DRILLING 236
TABLE 10-6 DIAMOND DRILL STATISTICS FOR UNION REEFS AREA 237
TABLE 10-7 SUMMARY OF 2011-12 DRILLING AT UNION REEFS AREA 239
TABLE 10-8 SUMMARY OF 2011-12 DRILLING AT UNION REEFS AREA 239
TABLE 10-9 SUMMARY OF DRILL RESULTS FROM PROSPECT AND CROSSCOURSE DEPOSITS -2012 241
TABLE 10-10 RC DRILL STATISTICS FOR UNION REEFS AREA 243
TABLE 10-11 HISTORIC DRILLING BY PROJECT – DIAMOND DRILLING 244
TABLE 10-12 UNION REEFS HISTORIC DRILLING BY PROJECT – RC DRILLING 244
TABLE 10-13 2012 DIAMOND DRILL STATISTICS FOR INTERNATIONAL DEPOSIT 245
TABLE 10-14: SIGNIFICANT INTERCEPTS FROM INTERNATIONAL DRILLING 245
TABLE 10-15 RC DRILL STATISTICS FOR PINE CREEK AREA 246
TABLE 10-16 HISTORIC DRILLING TYPES AT PINE CREEK 248
TABLE 10-17 2011- 12 DIAMOND DRILL STATISTICS FOR BURNSIDE AREA 248
TABLE 10-18 DRILL HOLE CO-ORDINATES MT BONNIE EAST 249
TABLE 10-19 2011- 12 RC DRILL STATISTICS FOR BURNSIDE AREA 249
TABLE 10-20: HISTORIC RC DRILLING AT RISING TIDE DEPOSIT 253
TABLE 10-21 HISTORIC DIAMOND DRILLING AT RISING TIDE DEPOSIT 253
TABLE 10-22 HISTORIC RC DRILLING AT YAM CREEK DEPOSIT 253
TABLE 11-1 RATE OF QA/QC SAMPLING FOR COSMO OPERATION 1 JANUARY 2010 TO 31 DECEMBER 2015 263
TABLE 11-2 LIST OF STANDARD SAMPLES USED AT COSMO MINE 263
TABLE 11-3 COSMO MINE STANDARD SR535 COMPLIANCE TABLE 265
TABLE 11-4 STATISTICAL RESULTS FOR COSMO MINE INTER-LAB REPEATS - 100, 200 & 300 LODES 268
TABLE 11-5 STASTISTICAL RESULTS FOR COSMO MINE INTER-LAB REPEATS - 101, 400, 500, 600 AND WESTERN LOADS 269
TABLE 11-6 ESMERALDA DEPOSIT QA/QC SAMPLING RATES 272
TABLE 11-7 ESMERALDA DEPOSIT INTER-LABRATORY STANDARD PERFORMANCE CHECK DATA 273
TABLE 11-8 ESMERALDA DEPOSIT ORIGINAL ASSAY VS REPEAT ASSAYS CORRELATION BETWEEN GRADE RANGES FOR LAB REPEATS, AU G/T 275
TABLE 11-9 ESMERALDA DEPOSIT TABLE OF STASTICS FOR 2015 LAB REPEATS, AU G/T 276
TABLE 11-10 ESMERALDA DEPOSIT ORIGINAL ASSAY VS REPEAT ASSAYS CORRELATION BETWEEN GRADE RANGES FOR INTER-LABRATORY CHECKS 277
TABLE 11-11 ESMERALDA DEPOSIT SUMMARY OF STASTICS FOR 2015 INTER-LABRATORY CHECK SAMPLES 277
TABLE 11-12 RATE OF QA/QC SAMPLING FOR UNION REEFS RC AND DIAMOND DRILLING 278

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TABLE 11-13 UNION REEFS STANDARD ST48/9278 COMPLIANCE TABLE 279
TABLE 11-14 UNION REEFS DIAMOND DUPLICATE ANALYSIS TABLE 281
TABLE 11-15 UNION REEFS DIAMOND PROGRAM DUPLICATE CORRELATION TABLE 281
TABLE 11-16 DUPLICATE R TABLE UNION REEFS DIAMOND PROGRAM 282
TABLE 11-17 UNION REEFS DIAMOND NAL:ALS INTER-LABORATORY REPEAT SUMMARY TABLE 283
TABLE 11-18 UNION REEFS DIAMOND INTER-LABRATORY REPEATS CORRELATION TABLE NTEL: ALS 283
TABLE 11-19 NAL:ALS INTER-LABRATORY REPEAT R TABLE UNION REEFS DIAMOND PROGRAM 284
TABLE 11-20 RATE OF QA/QC SAMPLING FOR INTERNATIONAL DEPOSIT DIAMOND DRILLING 287
TABLE 11-21 INTERNATION STANDARD ST08/8225 COMPLIANCE TABLE 288
TABLE 11-22 INTERNATIONAL DIAMOND DUPLICATE ANALYSIS TABLE 289
TABLE 11-23 INTERNATION DUPLICATE TABLE 289
TABLE 11-24 INTERNATION DIAMOND INTER-LAB REPEAT SUMMARY TABLE NAL:ALS 290
TABLE 11-25 INTERNATION DIAMOND NTEL:ALS INTER-LAB REPEATS CORELATION TABLE 291
TABLE 11-26 NAL:ALS INTER-LABS REPEAT R TABLE INTERNATIONAL DIAMOND PROGRAM 291
TABLE 11-27 RATE OF QA/QC SAMPLING FOR RISING TIDE & YAM CREEK RC AND DIAMOND DRILLING 293
TABLE 11-28 RATE OF QA/QC SAMPLING FOR RISING TIDE & YAM CREEK RC AND DIAMOND DRILLING 294
TABLE 11-29 RISING TIDE RC DUPLICATE ANALYSIS TABLE 295
TABLE 11-30 RISING TIDE RC DUPLICATE CORRELATION TABLE 295
TABLE 11-31 RISING TIDE DUPLICATE TABLE 296
TABLE 11-32 RISING TIDE RC INTER-LAB REPEAT SUMMARY TABLE NTEL:ALS 297
TABLE 11-33 RISING TIDE INTER-LAB REPEATS CORRELATION TABLE NTEL:ALS 297
TABLE 11-34 NTEL: ALS INTER-LAB REPEAT R TABLE RISING TIDE RC PROGRAM 298
TABLE 13-1 SUMMARY OF REPORTS AVAILABLE FOR COSMO MINE METALLURGICAL TEST WORK 304
TABLE 13-2 HEAD ASSAY RESULTS FOR COSMO MINE SAMPLES 304
TABLE 13-3 COSMO MINE RESULTS OF BOND WORK INDEX TEST WORK 304
TABLE 13-4 COSMO MINE BOND ABRASION INDEX RESULTS 304
TABLE 13-5 GRAVITY/DIRECT CYANIDATIOIN LEACH TEST WORK ON COSMO ORE 305
TABLE 13-6 DIRECT CYANIDATION TIME LEACH RESULTS ON COSMO ORE 305
TABLE 13-7 MINERALOGY OF COSMO ORE 305
TABLE 13-8 RESULTS OF OXYGEN UPTAKE TESTING COSMO MINE ORE 306
TABLE 13-9 HEAD ASSAYS FOR COSMO MINE SAMPLES 306
TABLE 13-10 DIRECT AND CIL CYANIDATION TESTWORK RESULTS FOR COSMO MINE ORE 307
TABLE 13-11 HEAD ASSAY RESULTS FOR COSMO MINE ORE 307
TABLE 13-12 GRAVITY/CYANIDATION LEACH TESTWORK RESULTS COSMO ORE 308
TABLE 13-13 COSMO MINE GRAVITY - CIL CYANIDATION LEACH TEST RESULTS COSMO MINE ORE 308
TABLE 13-14 COSMO MINE HEAD ANALYSIS 309
TABLE 13-15 COSMO MINE GOLD EXTRACTION RESULTS 310
TABLE 13-16 COSMO MINE HEAD ASSAYS COMPOSITES SUMMARY 311
TABLE 13-17 GOLD EXTRACTION AND PREG ROBBING RESULTS 311
TABLE 13-18 COSMO MINE OXYGEN UPTAKE RESULTS SUMMARY 312
TABLE 13-19 COSMO MINE BOND ABRASION (AI) DETERMINATIONS SUMMARY 312
TABLE 13-20 COSMO MINE 1 HEAD ASSAYS: SUMMARY 313
TABLE 13-21 COSMO MINE GRIND ESTABLISHMENT TEST WORK SUMMARY 313
TABLE 13-22 COSMO MINE SUMMARY OF GOLD EXTRACTION TEST WORK 314
TABLE 13-23 COSMO MINE OXYGEN UPDATE RATE DETERMINATIONS, TEST CONDITIONS 314
TABLE 13-24 COSMO MINE SUMMARY OF OXYGEN UPTAKE RATE TEST WORK 315
TABLE 13-25 SUMMARY OF REPORTS AVAILABLE FOR UNION REEFS DEPOSITS METALLURGICAL TEST WORK 315
TABLE 13-26 PROSPECT DEPOSIT HEAD ASSAYS COMPOSIT ES SUMMARY 316
TABLE 13-27 PROSPECT DEPOSIT SUMMARY OF GOLD EXTRACTION TEST WORK 316
TABLE 13-28 ESMERALDA PROSPECT HEAD ASSAYS COMPOSITES SUMMARY 317
TABLE 13-29 ESMERALDA PROSPECT SUMMARY OF GRIND RETENTION TIMES IN MINUTES 317
TABLE 13-30 ESMERALDA PROSPECT SUMMARY OF GOLD EXTRACTION TEST WORK 317
TABLE 13-31 ESMERALDA PROSPECT - SUMMARY OF ABRASION, ROD AND BALL MILL WORK INDEX RESULTS 318
TABLE 13-32 PINE CREEK MILL PRODUCTION FIGURES FROM OPEN FILE REPORTS 319
TABLE 13-33 METALURGICAL TEST-WORK FOR INTERNATIONAL DEPOSIT 319
TABLE 13-34 INTERNATIONAL DEPOSIT HEAD ASSAY RESULTS 319
TABLE 13-35 INTERNATIONAL DEPOSIT RESULTS OF BOND WORK INDEX TEST-WORK 320
TABLE 13-36 INTERNATIONAL DEPOSIT RESULTS OF GRAVITY SEPERATION TEST-WORK 320
TABLE 13-37 INTERNATIONAL DEPOSIT HEAD ASSAY RESULTS OF NAL TESWORK 321
TABLE 13-38 METALLURGICAL SAMPLES FROM INTERNATIONAL DEPOSIT 321
TABLE 13-39 RESULTS OF THE BOTTLE ROLL TEST-WORK ON INTERNATIONAL DEPOSIT MINERALIZATION 322
TABLE 13-40 CONSUMABLE REQUIREMENTS FOR INTERNATIONAL DEPOSIT MINERALIZATION WITH RESIDUE TIMINGS 322
TABLE 13-41 TEST WORK SUMMARY COMPLETED ON RISING TIDE DEPOSIT 323
TABLE 13-42 SAMPLE DETAILS FOR RISING TIDE DEPOSIT 323
TABLE 13-43 RISING TIDE DEPOSIT - SUMMARY OF DETECTED GOLD-SILVER MINERAL 324
TABLE 13-44 HEAD ASSAY FOR RISING TIDE SAMPLES 325
TABLE 13-45 GRAVITY AND CYANIDE EXTRACTION RESULTS FOR RISING TIDE DEPOSIT 326
TABLE 13-46 PREG-ROBBING CHARATERISTICS FOR RISING TIDE DEPOSIT SAMPLES 326
TABLE 14-1 NEWMARKET GOLD MINERAL RESOURCE STATEMENT – DECEMBER 31, 2015 328
TABLE 14-2 MINERAL RESOURCE ESTIMATIONS COSMO MINE PROJECT NORTHERN TERRITORY DEPLETED TO 31ST DECEMBER 2015 329
TABLE 14-3 RECONCILIATION RESULTS FOR COSMO MINE JANUARY - DECEMBER 2015, AU G/T 330
TABLE 14-4 BULK DENSITY FOR LODES AT COSMO MINE 334
TABLE 14-5 SUMMARY OF SAMPLE LENGTHS BY MINERALIZED DOMAIN 335
TABLE 14-6 MINERALIZED DOMAIN NOMENCLATURE 336
TABLE 14-7 SATISTICAL SUMMARY, SAMPLE LENGTH ALL MINERALISZED DOMAINS (FOOTWALL AND HANGING WALL) 337
TABLE 14-8 STATISTICAL SUMMARY, GOLD PPM - FOOTWALL DOMAINS 338

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TABLE 14-9 STATISTICAL SUMMARY, GOLD PPM – HANGINGWALL DOMAINS 338
TABLE 14-10 STATISTICAL SUMMARY FOR HIGH GRADE CUT COMPOSITES, GOLD G/T - FOOTWALL DOMAINS 339
TABLE 14-11 STATISTICAL SUMMARY FOR HIGH GRADE CUT COMPOSITES, GOLD G/T – HANGINGWALL DOMAINS 340
TABLE 14-12 ISOTROPIC VARIOGRAM MODELS FOR GOLD – FOOTWALL 341
TABLE 14-13 ISOTROPIC VARIOGRAM MODELS FOR GOLD – HANGINGWALL 341
TABLE 14-14 DYNAMIC KRIGING SEARCH PARAMETERS FOR GOLD – FOOTWALL DOMAINS – MINERALIZED AND WASTE 344
TABLE 14-15 DYNAMIC KRIGING SEARCH PARAMETERS FOR GOLD – HANGINGWALL DOMAINS – MINERALIZED AND WASTE 344
TABLE 14-16 INVERSE DISTANCE WEIGHTED SEARCH PARAMETERS FOR GOLD – FOOTWALL DOMAINS – MINERALIZED AND WASTE 345
TABLE 14-17 INVERSE DISTANCE WEIGHTED SEARCH PARAMETERS FOR GOLD – HANGINGWALL DOMAINS – MINERALIZED AND WASTE 345
TABLE 14-18 COSMO_UNDERGROUND_NI43101_EOY2015_DEPLETED.MDL BLOCK MODEL DEFINITION 346
TABLE 14-19 3D BLOCK MODEL ATTRIBUTES 346
TABLE 14-20 MINERALIZED DOMAIN AVERAGE GOLD GRADE (G/T) COMPARISONS 347
TABLE 14-21 MINERAL RESOURCE STATEMENT FOR COSMO MINE COMBINED HANGINGWALL AND FOOTWALL LODES AT 2.0 G/T GOLD CUT OFF, EFFECTIVE DEC 31 2015 353
TABLE 14-22 MINERAL RESOURCE ESTIMATIONS NEWMARKET GOLD DEPOSITS UNION REEFS AREA 355
TABLE 14-23 MODEL SUMMARY FOR UNION REEFS DEPOSITS 357
TABLE 14-24 UNION REEFS DEPOSITS MODEL SUMMARY OF MODEL INPUTS 358
TABLE 14-25 PROSPECT DEPOSIT LOAD SUMMARY 360
TABLE 14-26 PROSPECT DEPOSIT SUMMARY OF SAMPLE LE NGTHS BY MINERALIZED DOMAIN 361
TABLE 14-27 PROSPECT DEPOSIT HIGH -GRADE COMPOSITE CUTS BY DOMAIN 366
TABLE 14-28 PROSPECT DEPOSIT STATISTICAL SUMMARY, STOCKWORK DOMAINS 366
TABLE 14-29 PROSPECT DEPOSIT FINAL VARIGRAM MODELS BY DOMAIN 367
TABLE 14-30 PROSPECT DEPOSIT VEIN DOMAIN ESTIMATION VARIOGRAM MODELS 368
TABLE 14-31 PROSPECT DEPOSIT VEIN DOMAIN ESTIMATION PARAMETERS 368
TABLE 14-32 PROSPECT DEPOSIT STOCKWORK DOMAIN EST IMATION VARIOGRAM MODELS 369
TABLE 14-33 PROSPECT DEPOSIT STOCKWORK DOMAIN EST IMATION PARAMETERS 369
TABLE 14-34 PROSPECT DEPOSIT VEIN DOMAIN PROJECTION BLOCK MODEL DEFINITI ON 370
TABLE 14-35 PROSPECT DEPOSIT VEIN DOMAIN PROJECTION BLOCK MODEL ATTRIBUTES 371
TABLE 14-36 PROSPECT DEPOSIT STOCKWORK AND FINAL 3D BLOCK MODEL DEFINITION 371
TABLE 14-37 PROSPECT DEPOSIT STOCKWORK AND FINAL 3D BLOCK MODEL ATTRIBUTES 371
TABLE 14-38 PROSPECT DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK 372
TABLE 14-39 PROSPECT DEPPOSIT SPECIFIC GRAVITY DATA STATISTICS BY OXIDATION STATE 372
TABLE 14-40 PROSPECT DEPOSIT STOCKWORK DOMAIN AVERAGE GOLD GRADE COMPARISONS 373
TABLE 14-41 PROSPECT DEPOSIT VEIN DOMAIN BACK CALCULATED AVERAGE GOLD GRADE COMPARISONS 375
TABLE 14-42 SUMMARY OF DRILLING STATISTICS FOR THE DATA SET COVERING THE CROSSCOURSE DEPOSIT LODES 380
TABLE 14-43 SUMMARY OF SAMPLE LENGTHS BY HOLE TYPE FOR CROSSCOURSE DEPOSIT MINERALIZED DOMAINS 381
TABLE 14-44 CROSSCOURSE AND UNION REEFS WEST DEPOSITS STATISTICAL SUMMARY MINERALIZED DOMAINS 383
TABLE 14-45 CROSSCOURSE AND UNION REEFS WEST HIGH GRADE COMPOSITE STATISTICS BY MINERALIZED DOMAIN 384
TABLE 14-46 CROSSCOURSE AND UNION REEFS WEST DEPOSITS FINAL VARIOGRAM MODELS FOR LODE DOMAIN 386
TABLE 14-47 STATISTICS FOR CROSSCOURSE DEPOSIT E-LENS DOMAINS GOLD INDICATORS 389
TABLE 14-48 CROSSCOUSE DEPOSIT E-LENS DOMAINS MIK ESTIMATION SEARCH PARAMETERS 389
TABLE 14-49 CROSSCOURSE DOMAINS OK ESTIMATION PARAMETERS 390
TABLE 14-50 URW DOMAIN 2D PROJECTION BLOCK MODEL DEFINITION 390
TABLE 14-51 URW DOMAIN PROJECTION BLOCK MODEL ATTRIBUTES 391
TABLE 14-52 CROSSCOURSE DEPOSIT E-LENS DOMAINS AND FINAL 3D BLOCK MODEL DEFINITION 392
TABLE 14-53 CROSSCOURSE DEPOSIT E-LENS DOMAINS AND FINAL 3D BLOCK MODEL ATTRIBUTES 392
TABLE 14-54 CROSSCOURSE DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK 393
TABLE 14-55 CROSSCOURSE DEPOSIT BULK DENSITY DATA STATISTICS BY OXIDATION STATE 393
TABLE 14-56 DEPLETION CODES ASSIGNED ABOVE AND BELOW TOPOGRAPHIC SURFACES 394
TABLE 14-57 URW DOMAIN BACK CALCULATED AVERAGE GOLD GRADE COMPARISONS 394
TABLE 14-58 E-LENS DOMAINS AVERAGE GOLD GRADE COMPARISONS 395
TABLE 14-59 ORINOCO DEPOSIT - HIGH GRADE SAMPLE CUTS BY DOMAIN 406
TABLE 14-60 ORINOCO DEPOSIT FINAL VARIOGRAM MODELS BY DOMAIN 407
TABLE 14-61 ORINOCO DEPOSIT MINERALIZATION DOMAIN ESTIMATION PARAMETERS 407
TABLE 14-62 ORINOCO DEPOSIT FINAL 3D BLOCK MODEL DEFINITION 408
TABLE 14-63 ORINOCO DEPOSIT FINAL 3D BLOCK MODEL ATTRIBUTES 408
TABLE 14-64 ORINOCO DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK 409
TABLE 14-65 ORINOCO DEPOSIT BULK DENSITY (BD) DATA STATISTICS BY OXIDATION STATE 409
TABLE 14-66 ORINOCO DEPOSIT MINERALIZED DOMAIN AVERAGE GOLD GRADE COMPARISONS 410
TABLE 14-67 ESMERALDA OPEN PIT MINERAL RESOURCE ESTIMATION 414
TABLE 14-68 ESMERALDA UNDERGROUND MINERAL RESOURCE ESTIMATION 414
TABLE 14-69 COMBINED MINERAL RESOURCE FOR ESMERALDA PROSPECT (OPEN PIT AND UNDERGROUND) 417
TABLE 14-70 ESMERALDA PROSPECT DRILLHOLE SUMMARY BY TYPE 418
TABLE 14-71 ESMERALDA PROSPECT SUMMARY OF SAMPLE LENGTHS BY HOLE TYPE FOR ESMERALDA MINERALIZED DOMAINS 418
TABLE 14-72 ESMERALDA PROSPECT MINERALIZED DOMAIN SAMPLE LENGTH STATISTICS 419
TABLE 14-73 ESMERALDA PROSPECT COMPOSITE STATISTICS BY DOMAIN 419
TABLE 14-74 ESMERALDA PROSPECT HIGH GRADE RESTIRCTIONS BY DOMAIN 420
TABLE 14-75 ESMERALDA PROSPECT VARIOGRAM MODELS BY DOMAIN 421
TABLE 14-76 ESMERALDA DEPOST ORDINARY KRIGING ESTIMATION PARAMETERS 421
TABLE 14-77 BLOCK SIZE AND MODEL DIMENSIONS 422
TABLE 14-78 ESMERALDA MODEL CODING 422
TABLE 14-79 ESMERALDA PROSPECT 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK 423
TABLE 14-80 ESMERALDA PROSPECT BULK DENSITY DATA STATISTICS BY OXIDISATION STATE 423
TABLE 14-81 ESMERALDA PROSPECT DOMAIN AVERAGE GOLD GRADE COMPARISON 424
TABLE 14-82 ESMERALDA PROSPECT DISTANCES TO COMPOSITES USED IN THE MINERAL RESOURCE MODEL 426
TABLE 14-83 UNION REEFS DEPOSITS - BLOCK MODEL SET UP PARAMETERS 427

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TABLE 14-84 UNION REEFS MODEL PARAMETERS 428
TABLE 14-85 MINERAL RESOURCES FOR PINE CREEK DEPOSIT AS OF DEC 31 2015 432
TABLE 14-86 MODEL SUMMARY FOR PINE CREEK DEPOSITS 434
TABLE 14-87 PINE CREEK MODEL SUMMARY OF MODEL INPUTS 434
TABLE 14-88 INTERNATIONAL DEPOSIT SUMMARY OF SAMPLE LENGTHS BY MINERALIZED DOMAIN 438
TABLE 14-89 INTERNATIONAL DEPOSIT HIGH-GRADE COMPOSITE CUTS BY DOMAIN 441
TABLE 14-90 INTERNATIONAL DEPOSIT STATISTICAL SUMMARY BY DOMAIN 441
TABLE 14-91 INTERNATIONAL DEPOSIT FINAL VARIOGRAM MODELS BY DOMAIN 442
TABLE 14-92 INTERNATIONAL DEPOSIT MINERALIZED DOMAIN ESTIMATION VARIOGRAM MODEL 443
TABLE 14-93 INTERNATIONAL DEPOSIT MINERALIZED DOMAIN ESTIMATION PARAMETERS 443
TABLE 14-94 INTERNATIONAL DEPOSIT FINAL 3D BLOCK MODEL DEFINITION 444
TABLE 14-95 INTERNATIONAL DEPOSIT FINAL 3D BLOCK MODEL ATTRIBUTES 444
TABLE 14-96 INTERNATIONAL DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK 445
TABLE 14-97 INTERNATIONAL DEPOSIT SPECIFIC GRAVITY VALUES BY OXIDATION STATE 445
TABLE 14-98 INTERNATIONAL DEPOSIT MINERALIZED DOMAIN AVERAGE GOLD GRADE COMPARISONS 446
TABLE 14-99 INTERNATIONAL DEPOSIT COMPARISON OF TOTAL METAL BETWEEN OK AND MIK METHODS FOR ALL MINERALIZED DOMAINS 448
TABLE 14-100 INTERNATIONAL DEPOSIT COMPARISON OF HISTORIC PRODUCTION DATA AND CURRENT MODEL (OVERALL RECOVERY OF 79%) 451
TABLE 14-101 RESOURCE ESTIMATIONS NEWMARKET GOLD DEPOSITS NORTHERN TERRITORY 455
TABLE 14-102 COMMENTS ON MINERAL RESOURCES ESTIMATIONS OF NEWMARKET GOLD DEPOSITS, NORTHERN TERRITORY 457
TABLE 14-103 MINERAL RESOURCE SUMMARY FOR BURNSIDE AREA 457
TABLE 14-104 RISING TIDE DEPOSIT SUMMARY OF SAMPLE LENGTHS BY MINERALIZED DOMAIN 460
TABLE 14-105 RISING TIDE DEPOSIT STATISTICAL SUMMARY FOR GOLD IN PPM BY MINERAL RESOURCE ESTIMATION DOMAIN 462
TABLE 14-106 RISING TIDE DEPOSIT HIGH-GRADE COMPOSITE CUTS FOR GOLD IN G/T BY MINERAL RESOURCE ESTIMATION DOMAIN 463
TABLE 14-107 RISING TIDE DEPOSIT VARIOGRAM MODELS FOR GOLD BY MINERALIZED DOMAIN 463
TABLE 14-108 RISING TIDE DEPOSIT ESTIMATION PARAMETERS FOR GOLD BY ESTIMATION DOMAIN 464
TABLE 14-109 RISING TIDE DEPOSIT 3D BLOCK MODEL DEFINITION (M) 464
TABLE 14-110 RISING TIDE DEPOSIT 3D BLOCK MODEL ATTRIBUTES 464
TABLE 14-111 RISING TIDE DEPOSIT 3D BLOCK MODEL TO WIREFRAME VOLUMES CHECK 465
TABLE 14-112 RISING TIDE DEPOSIT SPECIFIC GRAVITY VALUES BY OXIDATION STATE 465
TABLE 14-113 RISING TIDE DEPOSIT MINERALIZED DOMAIN AVERAGE GOLD GRADE COMPARISONS 466
TABLE 14-116 BLOCK MODEL SUMMARY FOR KAZI DEPOSIT 476
TABLE 14-117 DRILLHOLE SUMMARY FOR KAZI 1989-1996 481
TABLE 14-198 HOWLEY DEPOSIT - BLOCK MODEL SUMMARY 482
TABLE 14-119 MOTTRAM DEPOSIT - SUMMARY STATISTICS 486
TABLE 14-120 BLOCK MODEL SET UP AND DIMENSIONS 487
TABLE 14-121 MOTTRAMS DEPOSIT - BLOCK MODEL SUMMARY 487
TABLE 14-122 MOTTRAMS DEPOSIT GLOBAL VALIDATION, G/T AU 488
TABLE 14-123 STATISTICAL SUMMARY FOR GOLD IN G/T BY MINERAL RESOURCE ESTIMATION DOMAIN 489
TABLE 14-124 3D BLOCK MODEL DEFINITION (M) 490
TABLE 14-125 3D BLOCK MODEL ATTRIBUTES 490
TABLE 14-126 VARIOGRAM MODELS FOR GOLD BY MINERALIZED DOMAIN 490
TABLE 14-127 ESTIMATION PARAMETERS FOR GOLD BY ESTIMATION DOMAIN 491
TABLE 14-128 PRINCESS LOUISE DEPOSIT – CUT STATISTICAL SUMMARY BY ZONE 492
TABLE 14-129 PRINCESS LOUISE DEPOSIT – BLOCK MODEL EXTENTS 492
TABLE 14-130 PRINCESS LOUISE DEPOSIT – BLOCK MODEL ATTRIBUTES 492
TABLE 14-131 PRINCESS LOUISE DEPOSIT – INTERPOLATION PARAMETERS 493
TABLE 14-132 FOUNTAIN HEAD UNCUT COMPOSIT STATISTICS BY LODE (G/T AU) 497
TABLE 14-133 FOUNTAIN HEAD PROJECT BLOCK MODEL PARAMETERS 498
TABLE 14-134 TALLY HO DEPOSIT UNCUT COMPOSIT STATISTICS BY LODE DOMAIN (G/T AU) 502
TABLE 14-135 TALLY HO DEPOSIT BLOCK MODEL PARAMETERS 503
TABLE 14-136 TALLY HO DEPOSIT SEARCH ELLIPSE DIMENSIONS FOR EACH LODE 504
TABLE 14-137 TALLY HO DEPOSIT GLOBAL STATISTICAL VALIDATION OF AU INTERPOLATED GRADES G/T 505
TABLE 15-1 NT OPERATIONS MINERAL RESERVE SUMMARY – EFFECTIVE DEC 31, 2015 508
TABLE 15-2 MINERAL RESERVE CLASSIFICATION FOR COSMO AS AT DECEMBER 31, 2015 508
TABLE 15-3 MINERAL RESERVE CLASSIFICATION PROSPECT DEPOSIT UNDERGROUND AS AT DECEMBER 31, 2015 509
TABLE 15-4 MINERAL RESERVE CLASSIFICATION ESMERALDA OPEN PIT AS AT DECEMBER 31, 2015 510
TABLE 15-5 MINERAL RESERVE CLASSIFICATION FOR PINE CREEK AS AT DECEMBER 31, 2015 511
TABLE 16-1 SUMMARY OF ROCK MASS QUANITIES OF COSMO GEOTECHNICAL DOMAINS (A.M.C 2014) 516
TABLE 16-2 RECOMMENDED PRIMARY GROUND SUPPORT SYSTEM AT COSMO MINE (A.M.C 2014) 517
TABLE 16-3 COSMO MINE CUT-OFF GRADE CALCULATIONS 521
TABLE 16-4 MINE OPERATING COSTS 522
TABLE 16-5 DESIGN PARAMETERS 522
TABLE 16-6 LATERAL DEVELOPMENT 523
TABLE 16-7 AIRFLOW REQUIREMENTS 525
TABLE 16-8 COSMO MINE DEVELOPMENT SCHEDULE 526
TABLE 16-9 COSMO MNE PRODUCTION SCHEDULE 526
TABLE 16-10 MINING FLEET 527
TABLE 16-11 NEWMARKET GOLD PERSONNEL REQUIREMENTS 527
TABLE 16-12 CONTRACTOR PERSONNEL REQUIREMENTS – COSMO MINE 528
TABLE 16-13 PROSPECT STOPING CUT- OFF GRADE CALCULATION 532
TABLE 16-14 PROSPECT DEVELOPMENT CUT- OFF GRADE CALCULATION 532
TABLE 16-15 DEVELOPMENT MINERALIZATION INVENTORY 534
TABLE 16-16 WASTE DEVELOPMENT QUANTITIES 535
TABLE 16-17 STOPE INVENTORY 535
TABLE 16-18 DEVELOPMENT SCHEDULE 537

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TABLE 16-19 MINERALIZATION PRODUCTION SCHEDULE 537
TABLE 16-20: WHITTLE PROCESSING PARAMETERS 539
TABLE 16-21: WHITTLE REVENUE PARAMETERS 539
TABLE 16-22: PIT DESIGN PARAMETERS 540
TABLE 16-23 ESMERALDA A PIT DESIGN RESULTS 541
TABLE 16-24 ESMERALDA B PIT DESIGN RESULTS 542
TABLE 16-25: MINING EQUIPMENT LIST 542
TABLE 16-26: PINE CREEK PIT DESIGN PARAMETERS 547
TABLE 16-27: KOHINOOR PIT DESIGN RESULTS 548
TABLE 16-28: COX PIT DESIGN RESULTS 549
TABLE 16-29: INTERNATIONAL PIT DESIGN RESULTS 550
TABLE 16-30: SOUTH ENTERPRISE PIT DESIGN RESULTS 551
TABLE 16-31: MINING EQUIPMENT LIST 552
TABLE 17-1 PRODUCTION FIGURES FOR UNION REEFS PLANT SINCE RESUMPTION OF OPERATIONS IN 2009. 559
TABLE20-1 LIST OF CURRENT MMP’S FOR NEWMARKET GOLD NT OPERATIONS 566
TABLE 20-2 EXAMPLE OF NAF/PAF SAMPLE COMPOSITE INFORMATION TO BE COLLECTED – HOLE CW92008 569
TABLE 20-3 TYPE AND ANALYTES TESTED 569
TABLE 20-4 LIST OF BONDING HELD BY NT OPERATIONS 578
TABLE 20-5 MINE CLOSURE REQUIREMENTS FOR COSMO MINE 580
TABLE 20-6 MINE CLOSURE REQUIREMENTS FOR UNION REEFS OPERATION 580
TABLE 20-7 MINE CLOSURE REQUIREMENTS FOR PINE CREEK SITE 580
TABLE 20-8 MINE CLOSURE REQUIREMENTS FOR NORTH POINT 581
TABLE 21-1 COSMO MINE CAPITAL COST SUMMARY 582
TABLE 21-2 COSMO MINE OPERATING COST SUMMARY 582
TABLE 21-3 CAPITAL COST SUMMARY 583
TABLE 21-4 OPERATING COST SUMMARY 583
TABLE 21-5 CAPITAL COSTS FOR ESMERALDA MINERAL RESERVES 584
TABLE 21-6 OPERATING COSTS FOR ESMERALDA OPERATIONS 585
TABLE 21-7 CAPITAL COSTS FOR PINE CREEK OPERATIONS 586
TABLE 21-8 OPERATING COSTS FOR PINE CREEK OPERATIONS 586
TABLE 22-1: MINING OPERATING COSTS 588
TABLE 22-2 ROYALTIES AND PROCESSING PARAMETERS 588
TABLE 22-3 ECONOMIC ANALYSIS RESULTS 590
TABLE 22-4 COMBINED OPERATIONAL CASH FLOWS 593
TABLE 22-5 PRE-TAX SENSITIVITIES 594
TABLE 23-1 FRANCIS CREEK MINERAL RESERVES AND MINERAL RESOURCES, 2011 601
TABLE 23-2 IRON BLOW DEPOSIT MINERAL RESOURCE ESTIMATE 602
TABLE 26-1 EXPLORATION PLANS FOR COSMO 607
TABLE 26-2 EXPLORATION PLANS FOR PINE CREEK 609
TABLE 26-3 EXPLORATION PLANS FOR BURNISDE AREA 610

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FIGURES PAGE
FIGURE 1-1 LOCATION PLAN OF NT OPERATIONS  2
FIGURE 1-2 EXPLORATION TARGETS OF THE COSMO DEPOSIT 12
FIGURE 2-1 CONVERSION OF MINERAL RESOURCES TO MINERAL RESERVES CIM STANDARDS 2014  24
FIGURE 4-1 NEWMARKET GOLD'S NORTHERN TERRITORY PROPERTIES 27
FIGURE 4-2 NEWMARKET GOLD’S NORTHERN TERRITORY PROPERTIES 28
FIGURE 4-3 BURNSIDE AREA TENEMENTS 33
FIGURE 4-4 UNION REEFS AREA TENEMENTS 34
FIGURE 4-5 PINE CREEK AREA TENEMENTS 35
FIGURE 4-6 NEWMARKET GOLD’S ROCKLAND RESOURCES – JV AGREEMENTS AREAS 38
FIGURE 4-7 NEWMARKET GOLD-PHOENIX COPPER (NOW PNX METALS) JOINT VENTURE AREAS 40
FIGURE 4-8 BIDDLECOMBE ROYALTY AGREEMENT FOR ELIZABETH TENEMENTS 48
FIGURE 4-9 TITLES COVERED BY THE PINE CREEK ROYALTY AGREEMENT 49
FIGURE 4-10 TENEMENTS COVERED BY FRANCO-NEVADA AGREEMENT 50
FIGURE 4-11 TENEMENTS COVERED BY FREEPORT-MCMORAN AGREEMENT 51
FIGURE 4-12 TITLES COVERED BY ON AND GROVES ROYALTY AGREEMENT NEAR NORTH POINT 51
FIGURE 4-13 TENEMENTS COVERED BY MMP #0546-03 54
FIGURE 4-14 TENEMENTS COVERED BY MMP #0539-03 55
FIGURE 4-15 TENEMENTS COVERED BY MMP #0538-01 56
FIGURE 4-16 TENEMENTS COVERED BY AUTHORIZATION #0525-02 57
FIGURE 4-17 TENEMENTS COVERED BY AUTHORIZATION #0526-01 57
FIGURE 4-18 TENEMENTS COVERED BY AUTHORIZATION #0528-01 58
FIGURE 4-19 TENEMENTS COVERED BY AUTHORIZATION #0530-01 58
FIGURE 5-1 GEOGRAPHICAL REGIONS OF THE NORTHERN TERRITORY 60
FIGURE 7-1 CRUSTAL SUB-DIVISIONS OF AUSTRALIA 83
FIGURE 7-2 GEOLOGY - NORTHERN TERRITORY (AHMAD, WYGRALAK AND FERENCZI 2009) 85
FIGURE 7-3 REGIONAL GEOLOGY, PINE CREEK OROGEN (AHMAD, WYGRALAK AND FERENCZI 2009) 86
FIGURE 7-4 STRATIGRAPHIC COLUMN, PINE CREEK OROGEN (GILLMAN, ET AL. 2009) 87
FIGURE 7-5 SIMPLE GEOLOGICAL MODEL FOR THE COSMO DEEPS DEPOSIT (J. MILLER 2014) 88
FIGURE 7-6 (A) ISOCLINAL FOLDS OF EARLY SILICA-PYRITE VEINS 89
FIGURE 7-7 (A) ISOCLINAL TO ELASTICA NON-CYLINDRICAL FOLDS IN MUDSTONE IN A DECOUPLED PRIMARY LAYER; 89
FIGURE 7-8 3D VIEW LOOKING DOWN TOWARDS THE NW UPON THE COSMO (NORTHERN) AND PHANTOM (SOUTHERN) OPEN PITS SHOWING THE RELATIVE LOCATIONS OF MAIN DOLERITE BODIES (GREEN) 90
FIGURE 7-9 COSMO DOLERITES WITH MINERALIZATION LOCATION SHOWN IN BLUE. LOOKING EAST 91
FIGURE 7-10 SCHEMATIC DIAGRAM SHOWING A CROSSECTION OF THE SEDIMENTARY CYCLES AND THE LOCATION OF THE 100 AND 200 LODES WITHIN THE CYCLE 92
FIGURE 7-11 SOUTH WALL OF THE WESTERN COSMO ANTICLINE LIMB IN THE COSMO OPEN PIT.  92
FIGURE 7-12 MAIN ROCK TYPES ASSOCIATED WITH GOLD MINERALIZATION AT COSMO DEEPS MINE 93
FIGURE 7-13 HANGINGWALL MINERALIZATION. SHOWING MACRO-SCALE FOLDING COMPLEXITY AND LOCATION OF THE EASTERN MINERALIZATION LODES AND CORRESPONDING WESTERN LODES (YELLOW) 95
FIGURE 7-14 LONGSECTION OF THE COSMO MINE 96
FIGURE 7-15 DEFORMATION SUMMAR Y UNION REEFS 98
FIGURE 7-16 LOCAL GEOLOGY FOR UNION REEFS GOLD PROJECT  99
FIGURE 7-17 MINERALIZATION AND STRUCTURE, NORTH W ALL OF CROSSCOURSE MINE 100
FIGURE 7-18 BLOCK MODEL DIAGRAM OF POTENTIAL STRUCTURES AT CROSSCOURSE 101
FIGURE 7-19 1.0G/T AU GRADE SHELLS IN THE CROSSCOURSE PIT. (A) PLAN VIEW. (B) LONGITUDINAL SECTION 102
FIGURE 7-20 1.65G/T AU GRADE SHELLS IN THE CROSSCOURSE PIT (A) PLAN VIEW. (B) LONG SECTION. 103
FIGURE 7-21 UNION REEFS DEPOSITS AREA 105
FIGURE 7-22 UNION REEFS MINERALIZED ZONES AND DEPOSITS – PLAN VIEW 106
FIGURE 7-23 SCHEMATIC LONG SECTION OF PINE CREEK MINERALIZED ZONES ( (GILLMAN, ET AL. 2009)) 108
FIGURE 7-24 LOCAL GEOLOGY FOR PINE CREEK AREA (AFTER KRUSE ET AL 1994) 109
FIGURE 7-25 STRATIGRAPHIC LOG OF PINE CREEK SEQUENCE SHOWING POSITION OF GOLD DEPOSITS. (MCGUIRE 2007) 110
FIGURE 7-26 MINERALIZATION IN THE ENTERPRISE DEPOSIT ( (AHMAD, WYGRALAK AND FERENCZI 2009)) 111
FIGURE 7-27 PINE CREEK DEPOSITS LOCATION OVER AEROMAGNETIC IMAGE 115
FIGURE 7-28 BURNSIDE – PROSPECTS AND DEPOSIT LOCATIONS 119
FIGURE 7-29 HOWLEY LINE DEPOSITS - SURFACE GEOLOGY 120
FIGURE 7-30 HOWLEY DEPOSIT SIMPLIFIED GEOLOGICAL SECTION (LOOKING NORTH) – HOWLEY ANTICLINE (GILLMAN, ET AL. 2009) 122
FIGURE 7-31 FOUNTAIN HEAD AND TALLY HO DEPOSITS –PLAN VIEW 125
FIGURE 7-32 SECTION 9825E THROUGH THE FOUNTAIN HEAD AND TALLY HO DEPOSITS – LOOKING WEST 125
FIGURE 7-33 MINERALIZATION EXAMPLES FROM COSMO MINE. LHS: EXAMPLE OF THE LINEAR EASTERN LODES. RHS: FOLDED EXAMPLE FROM THE HINGE ZONE 129
FIGURE 7-34 FOOTWALL MINERALIZATON. SHOWING PLAN OF THE FOUR EASTERN LODES. 130
FIGURE 7-35 COSMO MINERALIZATION - LONGSECTION OF BLOCK MODEL >2G/T AU ALL 100 LODES 132
FIGURE 7-36 COSMO MINERALIZATION - LONGSECTION OF BLOCK MODEL >2G/T AU SLIVER LODE 133
FIGURE 7-37 COSMO MINERALIZATION - LONGSECTION OF BLOCK MODEL >2G/T AU ALL 200 LODES 133
FIGURE 7-38 COSMO MINERALIZATION - LONGSECTION OF BLOCK MODEL >2G/T AU ALL 300 LODES 134
FIGURE 7-39 COSMO MINERALIZATION - LONGSECTION OF BLOCK MODEL >2 G/T AU ALL 400 LODES 134
FIGURE 7-40 COSMO MINERALIZATION - LONGSECTION OF BLOCK MODEL >2G/T AU ALL 500 LODES 135
FIGURE 7-41 COSMO MINERALIZATION - LONGSECTION OF BLOCK MODEL >2G/T AU ALL 600 LODES 135
FIGURE 7-42 NORTH POINT DEPOSIT SIMPLIFIED CROSS SECTION – LOOKING NORTH 141
FIGURE 7-43 PRINCESS LOUISE DEPOSIT SIMPLIFIED CROSS SECTION - LOOKING NORTH 141
FIGURE 8-1STRUCTURAL – STRATIGRAPHIC MODEL FOR NEWMARKET GOLD DEPOSITS – PINE CREEK OROGEN (A. K. SENER 2004, A. K. SENER 2004) 147
FIGURE 8-2 PINE CREEK OROGEN STRUCTURAL INTERPRETATION 152
FIGURE 8-3 LOCATION AND INTERPRETED DISPLACEMENT OF THE F1 FAULT LOOKING WEST. 154

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FIGURE 8-4 LONGSECTION SHOWING LOCATION OF THE F8 AND F9 FAULTS - LOOKING WEST. 155
FIGURE 8-5 PLAN VIEW TAKEN AT THE 940RL SHOWING DISPLACEMENT OF THE FOOTWALL MINERALIZATION LODES DUE TO THE DEXTRAL STRIKE SLIP MOVEMENT ON THE F9 FAULT 156
FIGURE 8-6 COSMO PIT SOUTHEAST PORTAL WALL LOOKING SOUTHEAST SHOWING DEEPLY INCISED SLOTS AT THE TRACES OF FAULTS: F10 AND OTHER FAULTS. DOLERITES ARE IN GREEN STIPPLE. 157
FIGURE 8-7 DIAMOND CORE DEMONSTRATING F-10 FAULT FABRIC AT DEPTH IN THE FOOTWALL OF THE EASTERN LIMB 158
FIGURE 8-8 MINERALIZATION DEVELOPMENT HEADING ON THE 955-100 SOUTH DRIVE SHOWING THE 100 LODE ON THE LEFT, THE F10 FAULT, AND THE 200 LODE ON THE RIGHT (LOOKING SOUTH). 158
FIGURE 8-9 FLAT NORTH DIPPING FAULTS SEEN IN THE EAST WALL OF THE COSMO PIT. 159
FIGURE 8-10 REGIONAL GEOLOGY AND BASE METAL DEPOSITS OF THE PINE CREEK OROGEN 162
FIGURE 9-1 CONCEPTUAL ILLUSTRATIONS TAKEN FROM (J. MILLER 2014) 164
FIGURE 9-2 PLAN OF UNDERGROUND GEOLOGY 165
FIGURE 9-3 LOCATION OF COSMO MINE NEAR MINE TARGETS, DRILL TESTED IN 2015 165
FIGURE 9-4 PERSPECTIVE VIEW DISPLAYING TARGET AREAS BENEATH, AND NORTH DOWN PLUNGE FROM THE COSMO DEPOSIT LOOKING TOWARDS THE WNW 166
FIGURE 9-5 PLAN VIEW – WESTERN LODES GOLD INTERCEPTS 169
FIGURE 9-6 PLAN VIEW SHOWING LOCATION OF THE WESTERN LODE DRILLING WITH RESPECT TO THE EASTERN LODE MINERALIZATION AND MAJOR STRUCTURES INTERPRETED PRIOR TO THE CURRENT STUDY 170
FIGURE 9-7 OBLIQUE 3D VIEW LOOKING APPROXIMAT DOWN PLUNGE OF THE MINERALIZATION (I.E. VIEWED AT APPROXIMATELY 50O>335O IN MINE GRID) 171
FIGURE 9-8 EXAMPLE OF HEMATITE-SILICA ALTERATION AT LANTERN AND THE PROXIMAL ASSOCIATION WITH GOLD MINERALIZATION (RHS VG = VISIBLE GOLD) FROM MILLER 2015A – HOLE CW101002 172
FIGURE 9-9 SECTION STRUCTURAL INTERPRETATION AND MINERALIZATION STYLE SUMMARY FOR LANTERN TARGET HOLE CW101002 (– FROM MILLER 2015A) 173
FIGURE 9-10 EXAMPLE OF LANTERN MINERALIZATION SHOWING FOLDED LAMINATIONS OF CHLORITE-SILICA-HEMATITE ALTERATION WITH PYRITE AND CARBONATE BEDS. 174
FIGURE 9-11 EXAMPLE OF LANTERN MINERALIZATION SHOWING HOW CHLORITE-SILICA-HEMATITE LAMINATIONS ARE COMMONLY FOUND TO THE MARGINS OF MARLY CARBONATE BEDS. 174
FIGURE 9-12 SCHEMATIC CROSS SECTION LOOKING NNW ACROSS THE LANTERN TARGET AREA 175
FIGURE 9-13 GRAPH OF HANDHELD XRF IRON RESULTS COMPARED TO FIRE ASSAY GOLD GRADES IN LANTE RN HOLE CW101006 176
FIGURE 9-14 LOCATION OF HANGINGWALL AND FOOTWALL LODES IN RELATION TO THE F1 FAULT MOVEMENT. LEFT HAND DIAGRAM SHOWS CURRENT MINERALIZATION. RIGHT HAND DIAGRAM SHOWS ESTIMATED LODE LOCATIONS PRE F1 FAULT MOVEMENT. 178
FIGURE 9-15 LONG SECTION ILLUSTRATING TARGETS FOR MINERAL RESOURCE & MINERAL RESERVE GROWTH IN 2016 179
FIGURE 9-16 INCLINED PLAN VIEW SHOWING THE MAJOR EXPLORATION GROWTH TARGETS FOR 2016 AT THE COSMO DEEPS MINE 179
FIGURE 9-17 PINE CREEK ELIZABETH MINE AREA REGIONAL STRUCTURAL INTERPRETATION 180
FIGURE 9-18 ELIZABETH MINE AREA BIDDLECOMBE (1985) OUTINE OF SPECULATIVE MINERAL RESOURCE 181
FIGURE 9-19 ELIZABETH MINE - UNION REEFS AREA AEROMAGNETIC SURVEY RTP 1ST VD BASE. 183
FIGURE 9-20 LOCATION PLAN FOR ESMERALDA DEPOSIT 184
FIGURE 9-21 ESMERALDA GEOLOGICAL MAPPING CARRIED OUT BY W P KARPETA 188
FIGURE 9-22 SCHEMATIC EAST-WEST SECTION THROUGH ESMERALDA A AND B WITH MINERALIZATION MARKED IN ORANGE 188
FIGURE 9-23 SCHEMATIC PLAN SHOWING THE FORMATION OF MINERALIZED BEDDING PLANE PARALLEL QUARTZ VEINS DURING THE F3 SINISTRAL STRIKE-SLIP DEFORMATION 189
FIGURE 9-24 ELR10 GRAB SAMPLES LOCATIONS AND RANGE OF ASSAYS - AU G/T 190
FIGURE 9-25 SURFACE DRILLING WITH GAS PIPELINE LOCATION 192
FIGURE 9-26 ENVIRONMENTAL DRILLING AT INTERNATIONAL DEPOSIT 194
FIGURE 9-27 STREAM SEDIMENT SAMPLE LOCATIONS, COSMO SOUTH 196
FIGURE 9-28 VTEM SURVEY CHANNEL 42 RESULTS OVER THE COSMO SOUTH AREA 198
FIGURE 9-29 MT BONNIE EAST VTEM CONDUCTOR – CHANNEL 35 WITH PROFILES 199
FIGURE 9-30 AIRBORNE CONDUCTORS ON TOTAL MAGNETIC INTENSITY AEROMAGNETIC BASE, MT ELLISON EAST AREA 201
FIGURE 9-31 MT ELLISON EAST GEOLOGY AND VTEM CONDUCTOR AXIS 206
FIGURE 9-32 VTEM SURVEY RESULTS, BFIELD CHANNEL 35 OVERLAIN WITH SOIL SAMPLE RESULTS 207
FIGURE 9-33 BONS RUSH EAST VTEM PROFILES 208
FIGURE 9-34 CONDUCTOR AXES FOR BLT_021 AND BLT_022. VTEM SURVEY LINE PATH IS SHOWN WITH BLACK LABELS, AND LATE TIME VTEM RESPONSE PROFILE IS SUPERIMPOSED. BACKGROUND IMAGE IS FIRST VERTICAL DERIVATIVE OF TMI MAGNETICS. VT EM LINE SPACING 150M 209
FIGURE 9-35 LINE 10600, LATE TIME VTEM CHANNELS (39-48). MODELED (RED) AND OBSERVED (BLACK) RESPONSES OF BLT_021 AND BLT_022 210
FIGURE 9-36 PLAN VIEW (A) OF CONDUCTIVE PLATES WITH LATE TIME (CH40) PROFILE SUPERIMPOSED ON FLIGHT PATHS FOR LINES 10600 AND 10610. SECTION VIEW (B) LOOKING FROM THE SOUTH WITH LATE TIME (CH40) PROFILE SUPERIMPOSED ON FLIGHT PATHS, AND SURFACE ELEVATION PROFILE JUST BELOW FLIGHT PATH 211
FIGURE 9-37 CONDUCTOR AXIS FOR BLT_020 AND BLT_026. VTEM SURVEY LINE PATH IS SHOWN WITH B LACK LABELS, AND LATE TIME VTEM RESPONSE PROFILE IS SUPERIMPOSED. BACKGROUND IMAGE IS FIRST VERTICAL DERIVATIVE OF TMI MAGNETICS. 212
FIGURE 9-38 LINE 10600, LATE TIME VTEM CHANNELS (40-48). MODELED (RED) AND OBSERVED (BLACK) RESPONSE OF BLT_026 213
FIGURE 9-39 LINE 10480, LATE TIME VTEM CHANNELS (40-48). MODELED (RED) AND OBSERVED (BLACK) RESPONSE OF BLT_020 214
FIGURE 9-40 PLAN VIEW (A) OF CONDUCTIVE PLATES AND FLIGHT PATHS FOR LINES 10470 AND 10480. SECTION VIEW (B) LOOKING FROM THE SOUTH WITH FLIGHT PATHS, AND SURFACE ELEVATION PROFILE 214
FIGURE 9-41 AG (PPB) IONIC LEACH SOIL SAMPLE RESULTS – VTEM ANOMALY AREA 217
FIGURE 9-42 SNAKEBITE AND EL25748 LOCATION MAP 218
FIGURE 9-43 NORTH CULLEN AREA VTEM ANOMALIES OVERLAIN WITH SOIL SAMPLE RESULTS AU-PPB 221
FIGURE 9-44 1:10K MAPPING OF JENKINS AREA. 223
FIGURE 9-45 BAN BAN AREA GEOLOGY WITH VTEM CONDUCTORS OVERLAIN 225
FIGURE 9-46 NORTHERN GOLD MAP DEPICTING THE BON’S RUSH DEPOSIT AREA FOLD NOSE 227
FIGURE 10-1 3D COSMO MINE DRILLING MODEL 230
FIGURE 10-2 CROSS SECTION OF COSMO MINE LOOKING SOUTH 233
FIGURE 10-3 PLAN OF COSMO MINE SURFACE RC DRILLING 2015 234
FIGURE 10-4 COSMO MINE LOCAL GRID CONVERSION PLAN 234
FIGURE 10-5 UNION REEFS LOCAL GRID CONVERSION PLAN 237

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FIGURE 10-6 ESMERALDA LOCAL GRID CONVERSION PLAN 238
FIGURE 10-7 2011-12 DRILLING LOCATIONS AT UNION REEFS AREA 242
FIGURE 10-8 2015 DRILLING LOCATIONS PLAN AT ESMERALDA A AREA 243
FIGURE 10-9 2012 DRILLING LOCATION AT INTERNATIONAL AREA 246
FIGURE 10-10 PINE CREEK MINE LOCAL GRID CONVERSION PLAN 247
FIGURE 10-11 2011 DRILLING LOCATION PLAN FOR THE RISING TIDE DEPOSIT 250
FIGURE 10-12 YAM CREEK AREA AND EXPLORATION DRILLING LOCATION FOR 2011 252
FIGURE 11-1 RC DRILL SAMPLING FLOW SHEET 255
FIGURE 11-2 DIAMOND DRIL RIG SAMPLING FLOW SHEET 256
FIGURE 11-3 COMOS MINE UNDERGROUND FACE WITH IDEAL MARK UP SAMPLING 258
FIGURE 11-4 NAL ANALYTICAL CONTROL SCHEME 260
FIGURE 11-5 ALS ANALYTICAL CONTROL SCHEME 261
FIGURE 11-6 NTEL ANALYTICAL CONTROL SCHEME 262
FIGURE 11-7 BOX AND WHISKER PLOT OF STANDARDS USED AT COSMO MINE IN 2015 264
FIGURE 11-8 COSMO MINE STANDARD ST535 COMPLIANCE CHART 265
FIGURE 11-9 BLANK ASSAY RESULTS - COSMO MINE 266
FIGURE 11-10 ORIGINAL (NAL) VS REPEAT (ALS) GRAPH, COSMO MINE 2015 DATA - 100,200 & 300 LODES 267
FIGURE 11-11 ORIGINAL (NAL) VS REPEATS (ALS) GRAPH, COSMO MINE 2015 DATA - 101, 400, 500, 600 & WESTERN LOADS 268
FIGURE 11-12 10Q-Q ORIGINAL VS REPEAT GRAPH, COSMO MINE 2015 DATA - ALL SAMPLES 269
FIGURE 11-13 ESMERALDA DEPOSIT NORMALISED STANDARD PERFORMANCE OF ALL STANDARDS 273
FIGURE 11-14 ESMERALDA DEPOSIT BLANK SAMPLE PERFORMANCE 274
FIGURE 11-15 ESMERALDA DEPOSIT 2015 LAB REPEATS 275
FIGURE 11-16 ESMERALDA DEPOSIT INTER-LABORATORY CHECK SAMPLES 276
FIGURE 11-17 UNION REEFS STANDARD ST48/9278 COMPLIANCE CHART 279
FIGURE 11-18 UNION REEFS DIAMOND DRILL BLANKS 280
FIGURE 11-19 UNION REEFS DUPICATE CORRELATION PLOT; RANGE ALL 282
FIGURE 11-20 UNION REEFS DIAMOND INTER-LABORATORY REPEATS FOR ALL RANGES NAL FA50:ALS AA25 284
FIGURE 11-21 UNION REEFS DIAMOND ASSAY VS SCREEN FIRE ASSAY 285
FIGURE 11-22 INERNATIONAL STANDARD ST 08/8225 COMPLIANCE CHART 287
FIGURE 11-23 INTERNATIONAL DEPOSIT DIAMOND BLANKS 288
FIGURE 11-24 INTERNATIONAL DUPLICATE CORRELATION PLOT; RANGE ALL 290
FIGURE 11-25 PINE CREEK DIAMOND INTER-LAB REPEATS FOR ALL RANGES NAL FA50: ALS AA25 291
FIGURE 11-26 RISING TIDE STANDARD ST48/9278 COMPLIANCE CHART 293
FIGURE 11-27 RISING TIDE RC BLANKS RESULTS 294
FIGURE 11-28 RISING TIDE DUPLICATE CORRELATION PLOT; RANGE <0.20PPM AU 296
FIGURE 11-29 RISING TIDE RC INTER-LAB REPEATS FOR ALL RANGES NTEL FA25: ALS AA25 298
FIGURE 13-1 UNION REEFS PROCESSING PLANT LAYOUT 303
FIGURE 14-1 DRILLING AND MINERALIZED WIREFRAMES OBLIQUE VIEW LOOKING NORTHWEST 336
FIGURE 14-2 APPLICATION OF DYNAMIC GRID ESTIMATE FOR LODES 100, 110, 120 AND 130 (10° DIP INCREMENTS) 342
FIGURE 14-3 APPLICATION OF DYNAMIC GRID ESTIMATE FOR LODES 100, 110, 120 AND 130 (10° AZIMUTH INCREMENTS) 343
FIGURE 14-4 G-T CURVE FOR THE FOOTWALL MINERALIZED LODES 348
FIGURE 14-5 G-T CURVE FOR THE HANGINGWALL MINERALIZED LODES 349
FIGURE 14-6 G-T CURVE FOR 100-LODE 349
FIGURE 14-7 G-T CURVE FOR 150-LODE 350
FIGURE 14-8 G-T CURVE FOR 101-LODE 350
FIGURE 14-9 G-T CURVE FOR 550-LODE 351
FIGURE 14-10 LOCATION OF MINERAL RESOURCES AT UNION REEFS DEPOSITS 356
FIGURE 14-11 PLAN OF PROSPECT DEPOSIT DRILLING 359
FIGURE 14-12 PROSPECT DEPOSIT CROSS SECTION VIEW (LOOKING NORTH) DRILLING 7325MN 360
FIGURE 14-13 PROSPECT DEPOSIT LOG TRANSFORMED GOLD PROBABILITY PLOT –GOLD BY OXIDATION 1 TO 5 362
FIGURE 14-14 PROSPECT DEPOSIT DRILLING AND STOCKWORK WIREFRAMES OBLIQUE VIEW LOOKING NORTHWEST 363
FIGURE 14-15 PROSPECT DEPOSIT VEIN WIREFRAMES OBLIQUE VIEW LOOKING NORTHWEST 364
FIGURE 14-16 PROSPECT DEPOSIT DOMAIN 40 WIREFRAME AND COMPOSITE DATA – LONG SECTION VIEW LOOKING WEST 370
FIGURE 14-17 PROSPECT DEPOSIT DOMAIN 300 BLOCK MODEL AND COMPOSITE DATA – LONG SECTION VIEW LOOKING WEST 374
FIGURE 14-18 PROSPECT DEPOSIT DOMAIN 400 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 374
FIGURE 14-19 PROSPECT DEPOSIT DOMAIN 300 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 375
FIGURE 14-20 PROSPECT DEPOSIT DOMAIN 30 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 376
FIGURE 14-21 PROSPECT DEPOSIT DOMAIN 400 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 376
FIGURE 14-22 PLAN SHOWING COVERAGE OF CROSSCOURSE DEPOSIT DRILLING 378
FIGURE 14-23 CROSS SECTION VIEW (LOOKING NORTH) OF CROSSCOURSE DEPOSIT E-LENS DRILLING – SECTION 6630MN 379
FIGURE 14-24 CROSS SECTION VIEW (LOOKING NORTH) OF CROSSCOURSE DEPOSIT E-LENS DRILLING – SECTION 6790MN 379
FIGURE 14-25 CROSS SECTION VIEW (LOOKING NORTH) OF UNION REEFS WEST DEPOSIT DRILLING – SECTION 6960MN 380
FIGURE 14-26 CROSSCOURSE AND UNION REEFS WEST DEPOSITS DRILLING AND MINERALIZED DOMAINS- OBLIQUE VIEW LOOKING NE 382
FIGURE 14-27 CROSSCOURSE LOG TRANSFORMED GOLD PROBABILITY PLOT FOR GOLD – E-LENS DOMAIN 100 COMPOSITES 384
FIGURE 14-28 CROSSCOURSE LOG TRANSFORMED GOLD PROBABILITY PLOT FOR GOLD – E-LENS DOMAIN 200 COMPOSITES 385
FIGURE 14-29 LOG TRANSFORMED GOLD PROBABILITY PLOT FOR GOLD – URW DOMAIN GEOLOGICAL COMPOSITES 385
FIGURE 14-30 URW 2D BLOCK MODEL AND COMPOSITE DATA – LONG SECTION VIEW LOOKING WEST 391
FIGURE 14-31 URW DOMAIN 1001 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 394
FIGURE 14-32 URW DEPOSIT COMPARISON BETWEEN THE 2D ACCUMULATION GOLD ESTIMATE AND OK 3D METHOD ESTIMATE 395
FIGURE 14-33 CROSSCOURSE DEPOSIT E-LENS DOMAIN 100 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 396
FIGURE 14-34 CROSSCOUSE DEPOSIT E-LENS DOMAIN 200 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 396
FIGURE 14-35 CROSSCOUSE DEPOIT E LENS DOMAIN 100 ESTIMATE METHODOLOGY COMPARISON. 397
FIGURE 14-36 CROSSCOURSE DEPOSIT E LENS DOMAIN 200 ESTIMATE METHODOLOGY COMPARISON 398
FIGURE 14-37 MINERAL RESOURCE CLASSIFICATIONS FOR URW LODE DOMAIN 1001 400
FIGURE 14-38 URW LODE DOMAIN 1001 MINERAL RESOURCE GRADE TONNAGE CURVE 401
FIGURE 14-39 CROSSCOURSE E LENS LODE DOMAIN 100 MINERAL RESOURCE GRADE TONNAGE CURVE 402

viii


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FIGURE 14-40 CROSSCOURSE E LENS LODE DOMAIN 200 MINERAL RESOURCE GRADE TONNAGE CURVE 402
FIGURE 14-41 ORINOCO DEPOSIT DRILLING AND MINERALIZATION WIREFRAMES PLAN VIEW LOOKING NORTHWEST 404
FIGURE 14-42 ORINOCO DEPOSIT DRILLING AND MINERALIZATION WIREFRAMES OBLIQUE VIEW LOOKING NORTHWEST 405
FIGURE 14-43 ORINOCO DEPOSIT DOMAIN 101 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 410
FIGURE 14-44 ORINOCO DEPOSIT DOMAIN 201 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 411
FIGURE 14-45 ORINOCO DEPOSIT DOMAIN 202 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 411
FIGURE 14-46 ORINOCO DEPOSIT DOMAIN 301 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 411
FIGURE 14-47 ORINOCO DEPOSIT TONNES AND GRADE CURVES COMPARISONS OF DIFFERENT ESTIMATION INPUT PARAMETERS 412
FIGURE 14-48 ESMERALDA PROSPECT - PLAN OF ALL DRILLING 416
FIGURE 14-49 ESMERALDA DEPOSIT PLAN OF MINERALIZED DOMAINS 417
FIGURE 14-50 ESMERALDA PROSPECT DOMAIN 2 HISTOGRAM OF GOLD GRADES, G/T 420
FIGURE 14-51 ESMERALDA DESPOST DOMAIN 4 MODEL AND COMPOSITE DATA – LONG SECTION VIEW LOOKING WEST 425
FIGURE 14-52 UNION REEFS RECONCILIATION OF MINED VALUES AGAINST MILL RECOVERY 429
FIGURE 14-53 UNION REEFS LOW GRADE WASTE DUMP LOCATION FROM CLOSURE FILES 430
FIGURE 14-54 LOCATION OF MINERAL RESOURCES AT PINE CREEK 433
FIGURE 14-55 PLAN OF INTERNATIONAL DEPOSIT DRILLING 436
FIGURE 14-56 CROSS SECTION VIEW (LOOKING NORTH) OF INTERNATIONAL DEPOSIT DRILLING 12550 MN 437
FIGURE 14-57 INTERNATIONAL DEPOSIT LOG TRANSFORMED GOLD PROBABILITY PLOT – GOLD BY OXIDATION 1 TO 5 439
FIGURE 14-58 INTERNATIONAL DEPOSIT DRILLING AND MINERALIZED WIREFRAMES INTERNATINAL DEPOSIT OBLIQUE VIEW LOOKING NORTHWEST 440
FIGURE 14-59 INTERNATIONAL DEPOSIT DOMAIN 100 BLOCK MODEL AND COMPOSITE DATA – LONG SECTION VIEW LOOKING WEST 446
FIGURE 14-60 INTERNATIONAL DEPOSIT DOMAIN 100 BLOCK MODEL AND COMPOSITE DATA – LONG SECTION VIEW LOOKING WEST 447
FIGURE 14-61 INTERNATIONAL DEPOSIT DOMAIN 100 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 447
FIGURE 14-62 INTERNATIONAL DEPOSIT DOMAIN 200 SWATH PLOTS BY NORTHING (LEFT) AND BY ELEVATION (RIGHT) 448
FIGURE 14-63 INTERNATIONAL DEPOSIT COMPARISON OF GRADE TONNAGE CURVES - OK AND MIK METHODS FOR ALL DOMAINS 449
FIGURE 14-64 INTERNATIONAL DEPOSIT COMPARISON OF METAL CURVES BETWEEN OK AND MIK METHODS FOR ALL MINERALISED DOMAINS 449
FIGURE 14-65 MAP SHOWING LOCATION OF BURNSIDE MINERAL RESOURCES (INCLUDING COSMO MINE) 456
FIGURE 14-66 PLAN OF RISING TIDE DEPOSIT DRILLING 458
FIGURE 14-67 CROSS SECTION VIEW (LOOKING NORTH) OF RISING TIDE DEPOSIT DRILLING 10080ME 459
FIGURE 14-68 RISING TIDE DEPOSIT DRILLING AND MINERALIZED WIREFRAMES OBLIQUE VIEW LOOKING NORTHEAST 461
FIGURE 14-69 RISING TIDE DEPOSIT DOMAIN 10 BLOCK MODEL AND COMPOSITE DATA – PLAN VIEW LOOKING DOWN DISPLAYING GOLD GRADE 467
FIGURE 14-70 RISING TIDE DEPOSIT DOMAIN 10 BLOCK MODEL AND COMPOSITE DATA – PLAN VIEW LOOKING DOWN DISPLAYING MINERAL RESOURCES CLASSIFICATION 467
FIGURE 14-71 RISING TIDE DEPOSIT DOMAIN 10 SWATH PLOTS BY NORTHING (LEFT) AND BY EASTING (RIGHT) 468
FIGURE 14-72 RISING TIDE DEPOSIT DOMAIN 80 SWATH PLOTS BY NORTHING (LEFT) AND BY EASTING (RIGHT) 468
FIGURE 14-73 WESTERN ARM DEPOSIT EASTING SWATH PLOT VALIDATION ALL DOMAINS 472
FIGURE 14-74 KAZI DEPOSIT BLOCK MODEL VALIDATION GRAPH, BY NORTHING - DOMAIN 90 477
FIGURE 14-75 KAZI DEPOSIT BLOCK MODEL VALIDATION GRAPH, BY NORTHING - DOMAIN 100 477
FIGURE 14-76 KAZI DEPOSIT BLOCK MODEL VALIDATION GRAPH, BY NORTHING - DOMAIN 110 478
FIGURE 14-77 KAZI DEPOSIT BLOCK MODEL VALIDATION GRAPH, BY NORTHING - DOMAIN 120 478
FIGURE 14-78 Q Q PLOT OF KAZI SAMPLE DUPLICATES 480
FIGURE 14-79 Q Q PLOT FOR KAZI PULP SAMPLING 480
FIGURE 14-80 HOWLEY DEPOSIT AU/NORTHING VALIDATION PLOT FOR HOWLEY DEPOSIT – ALL LODES 484
FIGURE 14-81 MOTTRAMS DEPOSIT LOG HISTOGRAM FOR LODE 100 485
FIGURE 14-82 MOTTRAMS DEPOSIT LOG HISTOGRAM FOR LODE 200 485
FIGURE 14-83 MOTTRAMS DEPOSIT LOG HISTOGRAM FOR LODE 300 485
FIGURE 14-84 MOTTRAMS DEPOSIT LOG HISTOGRAM FOR LODE 400 486
FIGURE 14-85 AU GRADE/NORTHING VALIDATION PLOT - 100 LODE, MOTTRAMS DEPOSIT 488
FIGURE 14-86 PRINCESS LOUISE DEPOSIT - AU GRADE/NORTHING VALIDATION PLOTS ZONE 100 494
FIGURE 14-87 PRINCESS LOUISE DEPOSIT - AU GRADE/NORTHING VALIDATION PLOTS ZONE 200 494
FIGURE 14-88 PRINCESS LOUISE DEPOSIT - AU GRADE/NORTHING VALIDATION PLOTS ZONE 300 495
FIGURE 14-89 PRINCESS LOUISE DEPOSIT - AU GRADE/NORTHING VALIDATION PLOTS ZONE 400 495
FIGURE 14-90 PRINCESS LOUISE DEPOSIT - AU GRADE/NORTHING VALIDATION PLOTS ZONE 500 496
FIGURE 14-91 FOUNTAIN HEAD LOG NORMAL HISTOGRAM PLOT L20 LODE 497
FIGURE 14-92 FOUNTAIN HEAD LOG NORMAL HISTOGRAM PLOT L30 LODE 497
FIGURE 14-93 FOUNTAIN HEAD LOG NORMAL HISTOGRAM PLOT L40 LODE 498
FIGURE 14-94 FOUNTAIN HEAD AU GRADE/DEPTH PLOT – L20 LODE 500
FIGURE 14-95 FOUNTAIN HEAD AU GRADE/DEPTH VALIDATION PLOT – L30 LODE 500
FIGURE 14-96 FOUNTAIN HEAD AU GRADE/EASTING VALIDATION PLOT – L40 LODE 501
FIGURE 14-97 FOUNTAIN HEAD AU GRADE/EASTING VALIDATION PLOT – L50 LODE 501
FIGURE 14-98 TALLY HO DEPOSIT AU GRADE/DEPTH VALIDATION PLOT – 100 LODE 506
FIGURE 14-99 TALLY HO DEPOSIT AU GRADE/DEPTH VALIDATION PLOT – 200 LODE 506
FIGURE 14-100 TALLY HO DEPOSIT AU GRADE/EASTING VALIDATION PLOT – 100 LODE 507
FIGURE 16-1 COSMO MINE CROSS SECTION OF LODES WITHIN THE MINERALIZATION ZONE LOOKING NORTH 514
FIGURE 16-2 COSMO MINE STOPE STABILITY CHART FOR THE EASTERN LODES 515
FIGURE 16-3 GROUND SUPPORT SELECTION CHART (A.M.C 2014) 516
FIGURE 16-4 COSMO MINE DECLINE LOCATION LOOKING WEST 518
FIGURE 16-5 COSMO STOPING BLOCKS LONGSECTION LOOKING WEST 519
FIGURE 16-6 DOWNHOLE STOPE AND FILL – STAGE 1 PRODUCTION 520
FIGURE 16-7 DOWNHOLE STOPE AND FILL – STAGE 2 PRODUCTION 520
FIGURE 16-8 DOWNHOLE & UPHOLE STOPE AND FILL – WITH CRF SILL PILLARS 521
FIGURE 16-9 COSMO VENTILATION CIRCUIT – CURRENT AT DEC 2015, LOOKING EAST 524
FIGURE 16-10 PROSPECT LODES AS SEEN FROM THE SOUTH 530
FIGURE 16-11 TOPING AND GEOTECHNICAL DOMAINS (AFTER (MCENHILL 2013)) 530
FIGURE 16-12 STOPING LAYOUT (BREMNER AND EDWARDS 2012) 533

ix


- x -

FIGURE 16-13 DEVELOPMENT COMMENCING FROM THE CURRENT LADY ALICE OPEN PIT 534
FIGURE 16-14 ESMERALDA A PIT DESIGN (3D VIEW) 540
FIGURE 16-15 ESMERALDA B PIT DESIGN (3D VIEW) 541
FIGURE 16-16 KOHINOOR PIT DESIGN - ISO VIEW 548
FIGURE 16-17 COX PIT DESIGN - ISO VIEW 549
FIGURE 16-18 INTERNATIONAL PIT DESIGN - ISO VIEW 550
FIGURE 16-19 SOUTH ENTERPRISE PIT DESIGN - ISO VIEW 551
FIGURE 17-1 MINE TO MILL ROAD MAP, COSMO MINE TO UNION REEFS MILL 553
FIGURE 17-2 UNION REEFS TAILINGS FACILITY 555
FIGURE 17-3 UNION REEFS PLANT LAYOUT 556
FIGURE 17-4 FLOW SHEET FOR PROCESSING AT UNION REEFS 557
FIGURE 20-1 FLOW CHART FROM NT EPA (NTEPA 2015)) SHOWING APPROVAL PROCESS FOR A PROJECT 568
FIGURE 20-2 HOLE CW92008 SAMPLES PLOTTED. COSMO MINE LOOKING SOUTH 570
FIGURE 20-3 NAF/PAF DRILLHOLES (IN YELLOW) WITHIN COSMO MINE LOOKING EAST. 570
FIGURE 20-4 REGISTERED ARCHAEOLOGICAL SITES FOR PINE CREEK 575
FIGURE 22-1 NT OPERATIONS MINING SCHEDULE 589
FIGURE 22-2 NT OPERATIONS PROCESSING SCHEDULE 589
FIGURE 22-3 BULLION PRODUCTION SCHEDULE 590
FIGURE 22-4 CASH FLOW AND NPV 591
FIGURE 22-5 POST-TAX CASH FLOW AND NPV 592
FIGURE 22-6 PRE- TAX NPV SENSITIVITY 595
FIGURE 23-1 ADJACENT PROPERTIES LOCATION MAP 596
FIGURE 23-2 SPRING HILL PROPERTY CONFIGURATION ON A GEOLOGY BASE 598
FIGURE 23-3 MOUNT BONNIE DEPOSIT DRILL PLAN 603
FIGURE 26-1 PLAN OF COSMO TARGETS 608

x



Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

1 EXECUTIVE SUMMARY

1.1 INTRODUCTION

This document has been prepared for Newmarket Gold Inc. (“Newmarket Gold”), the beneficial owners of the Northern Territory Operations (collectively, the “NT Operations”). Newmarket Gold is a Canadian, Toronto Stock Exchange listed (TSX) corporation, and the NT Operations comprise a group of mineral tenements totaling 2,030km2 in the Northern Territory, Australia which include an inventory of historical gold discoveries, historical and modern gold mines, and current mineral resources and mineral reserves.

In early July 2015, Newmarket Gold Inc. merged with Crocodile Gold Corp. (“Crocodile Gold”), to form a new Canadian gold mining company that has 100% ownership of the NT Operations including the producing Cosmo Mine.

This document provides a summary of the key changes in mineral resources and mineral reserves that have resulted from ongoing exploration and mineral resource definition drilling as well as ongoing mine design and evaluation up to December 31, 2015.

The NT Operations have previously been individually identified but frequently referred to as the Cosmo Mine, the Burnside Gold & Base Metals Project, the Union Reefs Gold Project and the Pine Creek Gold Project. Within each of these project areas are located numerous gold deposits with estimated mineral resources and mineral reserves. The processing facility at Union Reefs is factored into the economic evaluation of all of the Company’s mineral resources and mineral reserves in the NT Operations and as a result of the shared infrastructure and close proximity of the various projects Newmarket Gold has determined it is prudent to prepare one technical report and treat the NT Operations as a single project.

Since the publication of the last technical reports, Newmarket Gold has undertaken mining at the Cosmo Gold Mine and processed ore through the mill at Union Reefs. During the same period Newmarket Gold has completed exploration activities at the Esmeralda deposit to the south of the Union Reefs processing facility.

1.2 PROPERTY DESCRIPTION AND LOCATION

The NT Operations comprises a total of 141 mineral titles (including 133 granted and eight applications) covering an area of approximately 224km2. The NT Operations also comprises a total of 46 Exploration titles (all granted) that covers a total area of 1,806km2.

These tenements are generally 100% owned by Newmarket Gold as detailed in Table 1-1 (there are a two non-core titles operated by Newmarket Gold with less than 100% ownership):

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

License Type Number Area (km²)
Exploration Licence    
Exploration License (EL) 24 1,791.59
Exploration License Application (ELA) - -
Exploration License Retention (ELR) 2 14.83
Sub Total 26 1,806.41
Mineral Leases    
Mineral Claim (MC) 55 11.66
Mineral Lease (ML) 72 191.77
Mineral Lease Application (MLA) 8 20.11
Mineral Authority (MA) 6 0.82
Sub Total 141 224.36
Total   167 2,030.77 

TABLE 1-1 SUMMARY OF MINERAL TITLES FOR NEWMARKET GOLD NT OPERATIONS

Note*: Some areas of Exploration Licenses includes areas of Mineral Leases.

Geographically, the NT Operations are centered between the villages of Adelaide River to the north and Pine Creek to the south. The area was historically an important gold mining center, and is serviced by the Stuart Highway, 248km south-southeast of Darwin the capital city of the Northern Territory.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

1.3 GEOLOGY & MINERALIZATION

The NT Operations property falls within the Archaean to Palaeo-Proterozoic Pine Creek Orogen, one of the major mineral provinces of Australia. The Pine Creek Orogen is a deformed and metamorphosed sedimentary basin up to 14km maximum thickness covering an area of approximately 66,000km2 and extending from Katherine in the south to Darwin in the north. It hosts significant mineral resources of gold, uranium and platinum group metals (PGMs), as well as substantial base metals, silver, iron and tin-tantalum mineralization.

The Pine Creek Orogen comprises of a series of late Archaean granite-gneiss basement domes, which are overlain by a fluvial to marine sedimentary sequence. Several highly reactive rock units are included within this sedimentary sequence, including carbonaceous shale, iron stones, evaporite, carbonate and mafic to felsic volcanic units of the South Alligator and Finniss River Groups. This sequence has been subjected to regional greenschist facies metamorphism and multiphase deformation, which has resulted in the development of a northwest trending fabric. Subsequent widespread felsic volcanism and the intrusion of granitoids caused contact metamorphism, in aureoles between 500m and 2.0km wide, which overprint the earlier regional metamorphism. After the granitoid intrusions, during regional extensional deformation, an extensive array of northeast and northwest trending dolerite dykes intruded the metasedimentary sequence.

Gold mineralization within the Pine Creek Orogen is preferentially developed within strata of the South Alligator Group and lower parts of the Finniss River Group along anticlines, strike-slip shear zones and duplex thrusts located in proximity to the Cullen Granite Batholith. Of particular stratigraphic importance are the Wildman Siltstone, the Koolpin Formation, Gerowie Tuff, Mount Bonnie Formation and the Burrell Creek Formation.

The Cosmo Mine geology is made up of a series of distal cyclical marine depositional events contained in a sequence of inter-bedded siltstones, carbonaceous mudstones, banded ironstone, phyllites, dolerite sills and greywacke units.

Generally gold mineralization is associated with quartz veins that occur as stockwork veins, sheeted veins, and discordant quartz veins in faults and shear zones, and frequently as saddle-reefs. There is a common association with antiformal structures.

Gold occurs both as free gold, frequently associated with pyrite and arsenopyrite, and has been recorded as refractory in some deposits, but these are rare in the NT Operations project.

The Cosmo Mine mineralization lies within a marine siltstone package located between the Inner Zamu Dolerite sill and a +30m thick pyritic carbonaceous mudstone unit identified as the “Pmc” unit. Siltstones, near the Pmc contact often contain boudinaged chert lenses. These cherts are recrystallized to resemble the sucrosic texture of quartzite. The unit intercalates with massive and banded siltstones. The width of the gold hosting siltstones is 30 to 50m in the footwall of the F1 Fault and from several meters to 50+ meters in the hangingwall due to variably developed folding.

Four main lodes have been delineated in the Footwall Lodes and three in the Hangingwall Lodes in relation to the F1 Fault. These are the 100 Lode, 200 Lode, 300 Lode and the 400 Lode on the footwall of the Eastern Limb, with the 500 Lode, 600 Lode and 101 Lode in the hangingwall.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Gold mineralization is closely associated with arsenopyrite often seen within the boudinaged greywacke unit (Pgtb), especially in the 100 Lode. The mineralization styles, both on the hangingwall and footwall of the F1 Fault are very similar, with the main mineralization associated with, but not necessarily totally constrained within the Pgtb unit. The main sulphide minerals in the fresh rock are pyrite and arsenopyrite, with traces of sphalerite and chalcopyrite. Pyrrhotite occurs below depths of 300m and is predominantly seen in the Pmc unit.

The Union Reefs deposit model (including Esmeralda) generally conforms and supports the Pine Creek Orogen model as outlined in Section 8.1 . Gold mineralization has been focused within two zones, (Union and Lady Alice Line at Union Reefs and Zone “A“ and Zone “B“ at Esmeralda) in the sheared axial zones of two adjacent faulted antiforms that strike NNW-SSE. At Esmeralda the north eastern “Zone A” is within 300m of the contact of the Allamber Springs Granite of the Cullen Suite and lies within the outer metamorphic aureole of the granite. It dips steeply southwest and has been significantly silicified and brecciated. Chert facies rocks are reported to coincide with the mineralized zones, which locally contain visible gold.

Gold mineralization at Pine Creek is focused on the axial zones of parallel major upright folds. The most productive is termed the Enterprise Anticline; others include the less productive International-Czarina Anticline. The folds plunge shallowly towards 135 degrees at around 10 degrees and the limbs dip southwest and northeast at around 65 degrees. The fold axes are sub-vertical.

The Pine Creek Orogen also hosts some world-class uranium deposits, occasionally gold/PGM rich, and stratabound gold ± silver rich base metal deposits.

1.4 EXPLORATION, DEVELOPMENT AND OPERATIONS

The area currently covered by the NT Operations have undergone a lengthy exploration and development history that has spanned over 140 years of historical prospecting and mining and several waves of modern exploration and development in the 1980’s and 1990’s.

A total of over 3.7Moz of gold has reportable been produced from the Pine Creek Orogen and in excess of 3Moz have been produced in the past from deposits that are currently within the NT Operations Property.

It is estimated that over 750,000m of historical drilling have been completed within the land area covered by the Pine Creek Orogen. Since 2009, Crocodile Gold/Newmarket Gold has drilled roughly 220,000m of drilling across all NT Operations.

During the period between 2011 and 2015, Crocodile Gold/Newmarket Gold has drilled a total of 169,611m of diamond drilling into the Cosmo Mine. During the same period 2,969m of RC drilling was also completed within the same area.

At the Cosmo Mine, exploration efforts are centered on the definition of controls on gold mineralization to generate near mine exploration targets. This work resulted in four ‘in-mine’, and four ‘near-mine’, prioritized drill targets and recommendations to reprocess geophysical data and conduct additional targeted research projects around the mine.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Cosmo Mine exploration growth drill programs were conducted in six individual areas; and complimentary to the above mentioned exploration programs was the mining of a drive at the 640RL level with the purpose of providing optimal drill platforms to drill targets such as the Sliver, Hinge and Western Lodes to the deeper northern end of the underground mine.

The Cosmo Mine has been operating consistently since commercial production was declared with quarterly gold production ranging between 12,000 and 22,000oz.

At the Esmeralda deposit, located south of the Union Reefs mill, a mapping campaign was completed in 2014, which led to a series of drill holes being completed with the objective of improving the mineral resource classification from Inferred to Indicated.

Exploration activities at the Burnside area included mapping and sampling following up on targets generated by the 2011 airborne VTEM geophysical survey. This work has identified new targets that will require additional follow up work to determine the potential for future development.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

MINERAL RESOURCES AND MINERAL RESERVES

 6



Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Notes for Table 1-2:

1.

Mineral resources are stated as of December 31, 2015.

   
2.

Mineral resources are inclusive of mineral reserves, which are set out below.

   
3.

Mineral resources are calculated using these parameters.


  a.

Gold Price of $A1,500/oz, metallurgical recovery of 90-92.0% depending on mineral resource.

  b.

Lower cut-off of 2.0g/t Au is used to calculate the mineral resources for Underground deposit and 0.5g/t Au for open pit mineral resources at Pine Creek and Union Reefs and 0.7g/t Au for Burnside. A lower cut of 1.0 g/t Au for underground mineral resources at Crosscourse due to size of potential deposit.

  c.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces.

  d.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.


4.

The mineral resource estimates were prepared by Mark Edwards, B.Sc. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold who has over 18 years of relevant experience and is a qualified person for mineral resources as per the NI43- 101.

The mineral reserve estimate for the Cosmo Mine is summarized as follows:

Classification Tonnes (t) Gold (g/t) Gold (oz)
     Proven      
     Underground 479,000 3.50 53,800
     Stockpile 8,000 2.38 600
Proven Subtotal 487,000 3.47 54,400
     Probable      
     Underground 445,000 3.28 46,900
Total mineral reserve 932,000 3.38  101,300 

TABLE 1-3 COSMO MINE MINERAL RESERVE CLASSIFICATION AS AT DECEMBER 31, 2015

Notes on Table 1-3:

1.

The mineral reserve is stated as of December 31, 2015.

   
2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the National Instrument 43-101.

   
3.

mineral reserves were estimated using the following mining and economic factors:


  a.

14% dilution at 0.5g/t Au is added to all stopes, based on reconciled 2015 production.

  b.

Minimum stope width of 3.0m.

  c.

Stope recovery of 90%, based on reconciled 2015 production.

  d.

Crown pillar mining recovery of 50%.

  e.

15% dilution at the mineral resource 7 grade is added to all development.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

  f.

Mineralization development recovery of 100% is assumed.

  g.

A gold price of $A1,450/oz.

  h.

An overall processing recovery of 92.0% at a cost of $28.90/t.

  i.

Total mining cost used of $68.72/t.

  j.

Stockpiles include Cosmo material at the Mine and Union Reefs Processing facility.

  k.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100 oz.


4.

The cut-off grade for mineral reserves has been estimated at 2.3g/t Au.

   
5.

Mineral Reserve estimates were reviewed by Murray Smith who is a consultant with Mining Plus Pty Ltd. Mr. Smith is a Member and Chartered Engineer of the Australasian Institute of Mining and Metallurgy, has over 20 years of relevant engineering experience and is the Qualified Person for Mineral Reserves for Cosmo Mine.

The mineral reserve estimate for the Union Reefs Underground deposit at the Prospect Mine is based on bottom-up up-hole benching with backfill and longhole open stoping mining methods.

Classification Tonnes (t) Gold Grade (g/t) Gold (ozs)
Proven      
Probable 276,000 4.42 39,200
Total mineral reserve 276,000 4.42 39,200

TABLE 1-4 MINERAL RESERVE CLASSIFICATION PROSPECT DEPOSIT UNDERGROUND AS AT DECEMBER 31, 2015

Notes on Table 1-4:

1.

The mineral reserve is stated as of December 31, 2015.

   
2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the National Instrument 43-101.

   
3.

mineral reserves were estimated using the following mining and economic factors:


  a.

A 0.2m hangingwall and footwall skin has been added to the economic stope shape to allow for dilution.

  b.

Minimum stope width is 2m.

  c.

Stope recovery is 95% .

  d.

A gold price of $A1,450/ oz.

  e.

An overall processing recovery of 93% at a cost of $28.90/t.

  f.

Total mining cost of $87.10/t.

  g.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100 oz .


4.

The cut-off grade for mineral reserves has been estimated at 2.7g/t Au.

   
5.

Mineral Reserve estimates were reviewed by Murray Smith who is a consultant with Mining Plus Pty Ltd. Mr. Smith is a Member and Chartered Engineer of the Australasian Institute of Mining and Metallurgy, has over 20 years of relevant engineering experience and is the Qualified Person for Mineral Reserves at Prospect Underground .

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

The mineral reserves for the Union Reefs deposit for the Esmeralda open pit mine and are based on open pit mining techniques as at December 31, 2015.

Classification Tonnes (t) Gold Grade (g/t) Gold (ozs)
Proven      
Probable 244,000 1.61 12,700
Total mineral reserve 244,000 1.61 12,700

TABLE 1-5 MINERAL RESERVE CLASSIFICATION ESMERALDA OPEN PIT AS AT DECEMBER 31, 2015

Notes on Table 1-5:

1.

The mineral reserve is stated as of December 31, 2015

   
2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the National Instrument 43-101

   
3.

mineral reserves were estimated using the following mining and economic factors:


  a.

Dilution of 10% and mineralization loss of 5%

  b.

Mining costs of $4.50/t and processing costs of $26.00

  c.

A gold price of $A1,450/oz

  d.

An overall processing recovery of 90%

  e.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100 oz


4.

The cut-off grade for mineral reserves has been estimated at 0.7g/t Au

   
5.

Mineral reserve estimates were prepared by Mark Edwards who is a Member of the Australasian Institute of Mining and Metallurgy and has over 18 years of relevant experience and is the Qualified Person for mineral reserves for Esmeralda open pit as per the National Instrument 43-101.

The following is a summary of mineral reserves in the Pine Creek deposits.

Pit      Classification Tonnes (t) Gold Grade (g/t) Gold (ozs)
Cox Proven      
Probable 133,000 1.61 6,900
International Proven      
Probable 860,000 1.30 35,900
Kohinoor Proven      
Probable 129,000 2.39 9,900
South Enterprise Proven      
Probable 123,000 2.37 9,400
Total Mineral reserve 1,245,000 1.55 62,100

TABLE 1-6 MINERAL RESERVE CLASSIFICATION FOR PINE CREEK AS AT DECEMBER 31, 2015

Notes on Table 1-6:

1.

The mineral reserve is stated as of December 31, 2015

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the National Instrument 43-101

   
3.

mineral reserves were estimated using the following mining and economic factors:


  a.

Dilution of 15% and mineralization loss of 5% for all pits excluding International which used a mining dilution of 10%

  b.

Mining costs of $4.80/t and processing costs of $33.24

  c.

A gold price of $A1,450/oz

  d.

An overall processing recovery of 90% for all pits excluding International, which used a recovery of 85%

  e.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100oz


4.

The cut-off grade for mineral reserves has been estimated at 0.9g/t Au.

   
5.

Mineral reserve estimates were prepared by Mark Edwards who is a Member of the Australasian Institute of Mining and Metallurgy and has over 18 years of relevant experience and is the Qualified Person for mineral reserves at Pine Creek as per the National Instrument 43-101.

There are no known situations where the mineral reserves outlined above could be materially affected by environmental, permitting, legal, title, treatment, socio-economic or political issues. There is however some risk with any gold mineral reserve where the gold price may affect the overall economic viability of a mining operation.

1.6 CONCLUSIONS AND RECOMMENDATIONS

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations
1.6.1      COSMO MINE

Advancement in the geological and mineralization understanding for the Cosmo Mine over the past 12 months has resulted in the definition of exploration targets within the mine. It is recommended that this understanding continue to be developed and advanced through exploration drilling campaigns. Table 1-7 below covers the proposed exploration programs to continue the development of the Cosmo Mine.

Target Current
Exploration Status
Potential
outcome
Description Diamond
Drill Meters
Total Cost
Western Lodes Advanced Scoping Inferred Status Plunging mineralization system close to the 640 exploration drill drive 3,000 $350,000
Lantern 700 lode Project Scoping Inferred Status Detailed drilling of 700 lode material within Lantern lode close to current development 3,500 $350,000
Lantern Central Exploration Investigative Longer drilling testing the Central zone of the lantern target area 3,000 $400,000
Hinge Footwall Project Scoping Inferred Status Drill testing the hinge zone below the F1 fault, currently intersected with Sliver drilling 2,000 $250,000
Cosmo Deeps Project Scoping Investigative Investigative drilling of the 100- 300 lodes down plunge of current mineral resources 3,500 $420,000
Sliver Project Scoping Inferred Status Continue the development of the Sliver target down plunge of current mineral resources 5,800 $700,000
Cosmo Surface
2300mN
Exploration Investigative Test the down plunge extensions of the Sliver and Cosmo Deeps target 3,600 $900,000
Total Exploration       24,400 $3,370,000

TABLE 1-7 PROPOSED EXPLORATION PROGRAMS FOR COSMO MINE FOR 2016.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Figure 1-2 shows the approximate location of the Western Lode, Sliver Lode, Lantern Lode and the Footwall Hinge Lode. These areas have the potential to add additional ounces to the mineral resource and mineral reserve statement over the next 12 months. The 640 Drill Drive extension is currently underway and will form a good platform for underground drilling of most of these targets.

Infill drilling is also planned to continue in 2016 with the aim being to maintain at least 12-18 months of mineral reserves drilled out to 20m x 10m spacing. The possible status for the end of 2016 reflects the opinion of Mark Edwards, General Manager of Exploration for Newmarket Gold and Qualified Person for this technical report. This drilling is seen as a critical path to replacing mineral reserves mined each year. This is the highest priority drilling for the Cosmo Mine and is the focus for the geological team based at the mine. Each program is reviewed regularly and altered to provide the required outcomes for mine planning purposes.

1.6.2      UNION REEFS

Drilling completed at the Esmeralda deposit has demonstrated the potential for future mining activities. While an overall reduction in mineral inventory (when combining Indicated and Inferred inventories) has resulted, it has been recognized that one diamond hole performed badly with lower than expected core recoveries. This has resulted in the reduction of tonnes in the core of the Esmeralda A Deposit. It is recommended that a second RC hole be twinned with this diamond hole that returned questionable assay results to confirm the width of the mineralization. It would be estimated that an additional 100m of RC drilling be completed at a cost of $10,000 (excluding any potential mobilization costs).

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Work also needs to continue on developing the Mine Management Plan (MMP) for the Esmeralda deposit. This work advanced significantly over the past 12 months, with base line studies and a Notice of Intent supplied to the Northern Territory Environmental Protection Agency (NTEPA). This work will continue at an estimated cost of $80,000 to finalize the approval to mine.

With the potential to commence mining operations at the Esmeralda deposit, located to the south of the Union Reefs processing facility, it is recommended that all mineral resources around the Union Reefs processing facility be reviewed for mining potential. Some drilling was completed in 2011 around deposits such as Millars and Lady Alice. These drilling results should be used with the new understanding gained at Esmeralda and Prospect deposits, and new mineral resource estimations should be completed. These could then be optimized to identify the potential for open pit mining. The costs of this work would be captured within the current NT Operations staff budget.

1.6.3      PINE CREEK

In the Pine Creek area there are currently four different mineral reserve deposits reported, including International, Kohinoor, Cox and South Enterprise deposits. There is the potential to add one year’s additional processing material for the Union Reefs facility. Some work is required to further advance the permitting process for these operations. While the deposits are located on an active Mineral Lease, work is required on the development of a Mine Management Plan for operations. This will require $150,000 of test-work and reporting to be completed.

There is also the potential to identify additional mineral resources at Pine Creek, particularly around the Enterprise South and Gandy’s North deposits. It is estimated that 2,000m of RC drilling for Enterprise South would be required at a cost of $200,000. At Gandy’s North, a diamond drilling program of 1,500m is recommended at a cost of $400,000 in order to test the higher grade plunging structure. This would potentially be an underground target, but due to its proximity to the surface there would also be some open pit potential.

1.6.4      BURNSIDE AREA

Newmarket Gold has been active in the past 2 years in rationalizing land holdings and mineral resources within the Burnside area. During this period, the Iron Blow deposit has been divested to PNX Metals Ltd, the Bridge Creek deposit has been divested to a local quarry operation, and the Glencoe deposit has been divested to Ark Mines Ltd. Also during this period a series of smaller, non-core Mineral Leases have been sold to third parties. It is recommended that this divestment of non-core assets continue to rationalize holdings within the Company’s NT Operations.

The Western Arm, Kazi and Bon’s Rush deposits are located proximal to each other. These deposits contain Inferred mineral resources. None of these three deposits have been previously mined. It is interpreted that they contain significant amounts of oxide mineralization. The mineral resource estimates for these deposits were completed by previous owners and will require an update These estimates have been reviewed by the Author and are deemed to be suitable for reporting however, an update will allow for more modern techniques to be utilized. Investigations are underway to understand the amount and quality of diamond drilling that was previously completed, and the remaining drill core that is available for additional study and test work. This drill core could be analyzed for required QA/QC purposes. It is estimated that this work would cost in the order of $10,000.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Drilling would also be required to convert these Inferred mineral resources to a high category to be used in future mineral reserve estimations. Below is an estimation of the drilling requirements for each deposit to convert them into a suitable mine plan.

1.6.5      OTHER

Newmarket Gold will continue to review and make recommendations on the many mineral deposits contained within its NT Operations, in order to identify opportunities to expand its mineral resource base.

Farm-in agreements have been completed that allow third parties to carry out exploration on significant parts of the Company’s land position. It is anticipated that this allows for increased exploration expenditure that should identify opportunities for more focused work.

The Company should also regularly monitors local competitor activities in the area in order to quickly identify opportunities that may be potentially beneficial to Newmarket Gold, for example the opportunity to toll treat ore from deposits around the Union Reefs plant.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

2 INTRODUCTION AND TERMS OF REFERENCE

2.1 INTRODUCTION

The purpose of this technical report on the Northern Territory (NT) Operations is to support public disclosure of the mineral resource and mineral reserve estimates for the NT Operations, including the operating Cosmo Mine, as at December 31, 2015, and has been prepared for the use of Newmarket Gold to provide technical information to assist with business decisions and future project planning. This technical report conforms to National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101) in accordance with Form 43-101F1, Guidelines for Preparation of Technical Reports. Mineral resource and mineral reserve estimations are prepared in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards - On mineral resources and mineral reserves (May, 2014) as incorporated by reference in NI 43-101.

This technical report has been prepared at the request of Newmarket Gold by Mr. Mark Edwards of Newmarket Gold (the Author), with Sections 1-3, 15.1, 15.2, 15.5, 16.1, 16.2, 18.1, 18.2, 18.3, 21.1, 21.2, 22, 24, 25, 26 and 27 of this technical report reviewed by Murray Smith of Mining Plus Pty Ltd (the Independent Author) (collectively, the Authors).

Newmarket Gold is a Canadian TSX-listed gold mining and exploration company with three operating mines in Australia – the Fosterville and Stawell Gold Mines in the State of Victoria and the NT Operation’s Cosmo Mine in the Northern Territory.

This technical report includes a geological overview of the NT Operations, including a description of the geology, mineralization, key occurrences and deposits. It provides an update on mineral resources and mineral reserves, and makes recommendations on additional exploration and development drilling, which has the potential to upgrade mineral resource classifications and to augment the mineral reserve base.

The NT Operations have previously been individually identified but frequently referred to as the Cosmo Mine, the Burnside Gold & Base Metals Project, the Union Reefs Gold Project and the Pine Creek Gold Project. Within each of these project areas are located numerous gold deposits with estimated mineral resources and mineral reserves. The processing facility at Union Reefs is factored into the economic evaluation of all of the Company’s mineral resources and mineral reserves in the NT Operations and as a result of the shared infrastructure and close proximity of the various projects Newmarket Gold has determined it is prudent to prepare one technical report and treat the NT Operations as a single project.

Since the publication of the last technical reports, Newmarket Gold has undertaken mining at the Cosmo Mine and processed ore through the mill at Union Reefs. During the same period Newmarket Gold has completed exploration activities at the Esmeralda deposit located to the south of the Union Reefs processing facility.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

2.2 SCOPE OF WORK

The purpose of this technical report is to demonstrate the viability of the NT Operations (including the Cosmo Mine) through:

The Authors were engaged by Newmarket Gold to prepare this technical report in compliance with NI 43-101 and Form 43-101F1.

The Authors have relied upon information made available to them by Newmarket Gold, which has included, in part, access to historical electronic databases and files, internal technical memorandums and reports, drill logs, assay reports, etc.

The Authors have also relied upon the technical assistance of the consultancy group Cube Consulting, a Perth, Australia, based consultancy group specializing in the generation and review of mineral resource estimates. This assistance relates to the several mineral resource outlined in this report, where they assisted with some technical aspects of the estimate process. Where they have contributed has been noted in the text of this report. This work has been reviewed by the Authors and is included as required.

Additional information from public domain sources and the Authors’ files were utilized to prepare this technical report.

One of the Authors, Mark Edwards, is the General Manager for Exploration for Newmarket Gold and is a Qualified Person under the requirements as set out in NI 43-101 and is not independent. The other Author of this report, Murray Smith, is from Mining Plus Pty Ltd and for the purposes of this report is an Independent Author.

The Authors have reviewed all such information and determined it to be adequate for the purposes of this technical report. The Authors do not disclaim any responsibility for the above noted information.

2.3 AUTHORS, QUALIFICATIONS AND RESPONSIBILITIES

Responsibilities for the preparation of certain sections of this technical report have been assigned to individual authors as shown in Table 2-1.

Technical reporting responsibilities of this technical report, and such individual authors are not responsible for sections of this technical report other than those indicated in this table.

Technical Report Section Qualified Person Employer
1, 2, 3, 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14, 15.3, 15.4, 15.5, 16.3, 16.4, 17, 18.1, 18.4, 18.5, 19, 20, 21.3, 21.4, 22, 23, 24, 25, 26 & 27 Mark Edwards, BSc, MAusIMM (CP) MAIG Newmarket Gold
1, 2, 3, 15.1, 15.2, 15.5, 16.1, 16.2, 18.1, 18.2, 18.3, 21.1, 21.2, 22, 24, 25, 26 & 27 Murray Smith, B.Eng. (Mining), MAusIMM (CP) Mining Plus Pty Ltd

Table 2-1 Technical Reporting Responsibilities

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Mark Edwards is the General Manager for Exploration for Newmarket Gold. Mr. Edwards has 18 years’ experience working in gold operations and exploration in Western Australia, Botswana and the Northern Territory. Mr. Edwards has been involved in all aspects of the geological operations for Newmarket Gold (formerly Crocodile Gold) NT Operations Project (including the Cosmo Mine) since 2010. Mr. Edwards is a Chartered Geologist of the Australasian Institute of Mining and Metallurgy (member number 220787) as well as a Member of the Australian Institute of Geoscientists (member number 3655).

Murray Smith is a consultant with Mining Plus Pty Ltd. Mr. Smith is a Member and Chartered Engineer of the Australasian Institute of Mining and Metallurgy (member number 111064), has over 20 years of relevant engineering experience and is a Qualified Person under the requirements as set out in NI 43-101. Mr. Smith visited the property in March 2014 including touring the underground operation and meeting key staff.

2.4 DEFINITIONS

In this technical report, reference to the “NT Operations” of Newmarket Gold refers to all deposits and operations located in the Northern Territory. Additionally, reference to “Cosmo Mine” (Cosmo) of Newmarket Gold refers to the current mine area, which has been in operation since 2010. This deposit was previously reported in 2009, 2011, 2013, 2014 and again in 2015 with other deposits in the Pine Creek region, owned and operated by Crocodile Gold (now Newmarket Gold). Other deposits in the NT Operations area were reported previously in a single report in 2011 and separately in 2013.

The regional coordinate system utilized throughout the properties is the Universal Transverse Mercator System (UTM) projection. The Global Positioning System (GPS) datum is WGS-84, Zone 52L. Local mine grid conversion will be shown later in the report. Mineral resource estimates were carried out on the local grid corresponding to each individual mineral resource. All units, unless expressed otherwise, are in the Metric System. All gold assay grades are expressed as grams per metric tonne (g/t) unless otherwise specified, with tonnages stated in metric tonnes. Gold metal is reported in troy ounces.

Unless otherwise stated, monetary values are in Australian Dollars ($A).

Abbreviation Unit or Term
Historical Mineral resource

Non-compliant mineral resource as reported in publicly available documentation. In no terms is this type of mineral resource to be included or quantified but is noted in this technical report to reflect previous work that has been completed on deposits outside the current listing in this mineral resource statement

IRR

Internal Rate of Return

kg

Kilogram(s)

km

Kilometer(s)

m

Meter (s)

Mt

Million tonnes

Mtpa

Million tonnes per annum

MMP

Mine Management Plan

NPV

Net Present Value

NMI

Newmarket Gold Inc.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Abbreviation Unit or Term
NT Northern Territory, Australia
NTG Northern Territory Government
NTGS Northern Territory Geological Survey
Oz Troy ounces (31.1035 g)
Oz/an Ounce (gold) per annum
% Per cent by weight
PAF Potential Acid Forming
Pb Lead
PEA Preliminary Economic Assessment
ppb Parts Per Billion
ppm Parts Per Million
Pmc Graphitic Mudstone Unit at Cosmo
QA/QC “Quality Assurance – Quality Control”
QP, Qualified Person “Qualified Person” has the meaning as ascribed to such term in NI43-101
RAB Rotary Air Blast drill hole
RAR Return Airway Rise
RC Reverse Circulation Drill Hole
ROM Run of Mine mineralization pad
T or t Metric tonne (2,204lbs)
U3O8 Uranium Oxide
VTEM Versatile Time Domain Electromagnetic Surveying – Geophysical Surveying technique
WA State of Western Australia, Australia
$A Australian Dollar
$C Canadian Dollar
AusIMM Australasian Institute of Mining & Metallurgy
AIG Australian Institute of Geoscientists
Ag Silver
AMC Australian Mining Consultants
Au Gold
Azi Azimuth
BCM Bulk Cubic Meter
BLEG Bulk Leachable Gold analysis for soil sampling
BOPL Burnside Operations Pty Ltd
BJV Burnside Joint Venture
CIM Canadian Institute of Mining, Metallurgy & Petroleum
Cu Copper
CRF Cement Rock Fill
CRK, CGA, CGAO Crocodile Gold (now Newmarket Gold)
DME Northern Territory Department of Mines and Energy (Mines Department)

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Abbreviation Unit or Term
DD, DDH Diamond Drilling, Diamond Drill Hole
FAR Fresh Airway Rise
Fm Formation
g or gm Gram (s)
g/t Grams per tonne
ha Hectare (10,000m2)
GCPL Geotechnical Consultants Pty Ltd
$US United States Dollar
°C Degrees Celsis
WDL Water Discharge License
WRD Waste Rock Dump
Zn Zinc

TABLE 2-1 DEFINITIONS

2.5 MINERAL RESOURCE AND MINERAL RESERVE DEFINITIONS

The following definitions have been taken from the CIM definition standards for mineral resources and Reserves, prepared by the CIM Standing Committee on Reserve Definitions, and adopted by CIM Council on May 10, 2014.

CIM Definitions are underlined and defined terms referenced to NI 43-101 are double underlined.

2.5.1      MINERAL RESOURCES

Mineral resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured categories. An Inferred mineral resource has a lower level of confidence than that applied to an Indicated mineral resource. An Indicated mineral resource has a higher level of confidence than an Inferred mineral resource but has a lower level of confidence than a Measured mineral resource.

A mineral resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction.

The location, quantity, grade or quality, continuity and other geological characteristics of a mineral resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.

Material of economic interest refers to diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals.

The term mineral resource covers mineralization and natural material of intrinsic economic interest, which has been identified and estimated through exploration and sampling, and within which mineral reserves may subsequently be defined by the consideration and application of Modifying Factors. The phrase ‘reasonable prospects for eventual economic extraction’ implies a judgment by the Qualified Person(s) in respect of the technical and economic factors likely to influence the prospect of economic extraction. The Qualified Person(s) should consider and clearly state the basis for determining that the material has reasonable prospects for eventual economic extraction. Assumptions should include estimates of cutoff grade and geological continuity at the selected cut-off, metallurgical recovery, smelter payments, commodity price or product value, mining and processing method as well as mining, processing and general and administrative costs. The Qualified Person(s) should state if the assessment is based on any direct evidence and testing.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Interpretation of the word ‘eventual’ in this context may vary depending on the commodity or mineral involved. For example, for some coal, iron, potash deposits and other bulk minerals or commodities, it may be reasonable to envisage ‘eventual economic extraction’ as covering time periods in excess of 50 years. However, for many gold deposits, application of the concept would normally be restricted to perhaps 10 to 15 years, and frequently to much shorter periods of time.

2.5.1.1  Inferred mineral resource

An Inferred mineral resource is that part of a mineral resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply, but not verify, geological and grade or quality continuity.

An Inferred mineral resource has a lower level of confidence than that applying to an Indicated mineral resource, and must not be converted to a mineral reserve. It is reasonably expected that the majority of Inferred mineral resources could be upgraded to Indicated mineral resources with continued exploration.

An Inferred mineral resource is based on limited information and sampling gathered through appropriate sampling techniques from locations such as outcrops, trenches, pits, workings and drill holes. Inferred mineral resources must not be included in the economic analysis, production schedules, or estimated mine life in publicly disclosed Pre-Feasibility or Feasibility Studies, or in the Life of Mine plans and cash flow models of developed mines. Inferred mineral resources can only be used in economic studies as provided under NI 43-101.

There may be circumstances, where appropriate sampling, testing, and other measurements are sufficient to demonstrate data integrity, geological and grade/quality continuity of a Measured or Indicated mineral resource, however, quality assurance and quality control, or other information may not meet all industry norms for the disclosure of an Indicated or Measured mineral resource. Under these circumstances, it may be reasonable for the Qualified Person to report an Inferred mineral resource if the Qualified Person(s) has taken steps to verify the information meets the requirements of an Inferred mineral resource.

2.5.1.2  Indicated mineral resource

An Indicated mineral resource is that part of a mineral resource for which quantity, grade or quality, densities, shape and physical characteristics are estimated with sufficient confidence to allow the application of Modifying Factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit.

20


Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing, and is sufficient to assume geological and grade or quality continuity between points of observation.

An Indicated mineral resource has a lower level of confidence than that applying to a Measured mineral resource and may only be converted to a Probable mineral reserve.

Mineralization may be classified as an Indicated mineral resource by the Qualified Person(s) when the nature, quality, quantity and distribution of data are such as to allow confident interpretation of the geological framework, and to reasonably assume the continuity of mineralization. The Qualified Person(s) must recognize the importance of the Indicated mineral resource category to the advancement of the feasibility of the project. An Indicated mineral resource estimate is of sufficient quality to support a Pre-Feasibility Study, which can serve as the basis for major development decisions.

2.5.1.3  Measured mineral resource

A Measured mineral resource is that part of a mineral resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit.

Geological evidence is derived from detailed and reliable exploration, sampling and testing and is sufficient to confirm geological and grade or quality continuity between points of observation.

A Measured mineral resource has a higher level of confidence than that applying to either an Indicated mineral resource or an Inferred mineral resource. It may be converted to a Proven mineral reserve or to a Probable mineral reserve.

Mineralization or other natural material of economic interest may be classified as a Measured mineral resource by the Qualified Person(s) when the nature, quality, quantity and distribution of data are such that the tonnage and grade or quality of the mineralization can be estimated to within close limits and that variation from the estimate would not significantly affect potential economic viability of the deposit. This category requires a high level of confidence in, and understanding of, the geology and controls of the mineral deposit.

2.5.2      MODIFYING FACTORS

Modifying Factors are considerations used to convert mineral resources to mineral reserves. These include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

2.5.3      MINERAL RESERVES

A mineral reserve is the economically mineable part of a Measured and/or Indicated mineral resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at Pre-Feasibility or Feasibility level as appropriate that include application of Modifying Factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

The reference point at which mineral reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.

The public disclosure of a mineral reserve must be demonstrated by a Pre-Feasibility Study or Feasibility Study.

Mineral reserves are those parts of mineral resources which, after the application of all mining factors, result in an estimated tonnage and grade which, in the opinion of the Qualified Person(s) making the estimates, is the basis of an economically viable project after taking account of all relevant Modifying Factors. Mineral reserves are inclusive of diluting material that will be mined in conjunction with the mineral reserves and delivered to the treatment plant or equivalent facility. The term ‘mineral reserve’ need not necessarily signify that extraction facilities are in place or operative or that all governmental approvals have been received. It does signify that there are reasonable expectations of such approvals.

‘Reference point’ refers to the mining or process point at which the Qualified Person(s) prepares a mineral reserve. For example, most metal deposits disclose mineral reserves with a “mill feed” reference point. In these cases, reserves are reported as mined ore delivered to the plant and do not include reductions attributed to anticipated plant losses. In contrast, coal reserves have traditionally been reported as tonnes of “clean coal”. In this coal example, reserves are reported as a “saleable product” reference point and include reductions for plant yield (recovery). The Qualified Person(s) must clearly state the ‘reference point’ used in the mineral reserve estimate.

2.5.3.1  Probable mineral reserve

A Probable mineral reserve is the economically mineable part of an Indicated, and in some circumstances, a Measured mineral resource. The confidence in the Modifying Factors applying to a Probable mineral reserve is lower than that applying to a Proven mineral reserve.

The Qualified Person(s) may elect, to convert Measured mineral resources to Probable mineral reserves if the confidence in the Modifying Factors is lower than that applied to a Proven mineral reserve. Probable mineral reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study.

2.5.3.2  Proven mineral reserve

A Proven mineral reserve is the economically mineable part of a Measured mineral resource. A Proven mineral reserve implies a high degree of confidence in the Modifying Factors.

Application of the Proven mineral reserve category implies that the Qualified Person(s) has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report. The term should be restricted to that part of the deposit where production planning is taking place and for which any variation in the estimate would not significantly affect the potential economic viability of the deposit. Proven mineral reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study. Within the CIM Definition standards the term Proved mineral reserve is an equivalent term to a Proven mineral reserve.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations
 
2.5.4      FEASIBILITY STUDY

A Feasibility Study is a comprehensive technical and economic study of the selected development option for a mineral project that includes appropriately detailed assessments of applicable Modifying Factors together with any other relevant operational factors and detailed financial analysis that are necessary to demonstrate, at the time of reporting, that extraction is reasonably justified (economically mineable). The results of the study may reasonably serve as the basis for a final decision by a proponent or financial institution to proceed with, or finance, the development of the project. The confidence level of the study will be higher than that of a Pre-Feasibility Study.

2.5.5      PRELIMINARY FEASIBILITY STUDY

A Pre-Feasibility Study is a comprehensive study of a range of options for the technical and economic viability of a mineral project that has advanced to a stage where a preferred mining method, in the case of underground mining, or the pit configuration, in the case of an open pit, is established and an effective method of mineral processing is determined. It includes a financial analysis based on reasonable assumptions on the Modifying Factors and the evaluation of any other relevant factors which are sufficient for a Qualified Person, acting reasonably, to determine if all or part of the mineral resource may be converted to a mineral reserve at the time of reporting. A Pre-Feasibility Study is at a lower confidence level than a Feasibility Study.

2.5.6      PRELIMINARY ECONOMIC ASSESSMENT

A Preliminary Economic Assessment (PEA) is a study, other than a pre-feasibility study or feasibility study, which includes an economic analysis of the potential viability of mineral resources. It can only demonstrate the potential viability of mineral resources, not the technical or economic viability of a project.

2.5.7      MINERAL RESOURCE AND MINERAL RESERVE CLASSIFICATION

The CIM Definition Standards provide for a direct relationship between Indicated mineral resources and Probable mineral reserves and between Measured mineral resources and Proven mineral reserves. In other words, the level of geoscientific confidence for Probable mineral reserves is the same as that required for the in situ determination of Indicated mineral resources and for Proven mineral reserves is the same as that required for the in situ determination of Measured mineral resources. Figure 2-1, displays the relationship between the mineral resource and mineral reserve categories.

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Technical Report Newmarket Gold Inc.
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Technical Report Newmarket Gold Inc.
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3 RELIANCE ON OTHER EXPERTS AND DISCLAIMER

This technical report has been prepared by the Authors for Newmarket Gold and is based, in part, as specifically set forth below, on the review, analysis, interpretation and conclusions derived from information which has been provided or made available to the Authors by Newmarket Gold, augmented by direct field examination and discussion with former employees, current employees of Newmarket Gold and consultants who have previously worked for past operators or are currently working for Newmarket Gold.

Mr. Smith has not performed any sampling or assaying, performed any detailed geological mapping, excavated any trenches, drilled any holes or carried out any independent exploration work.

Newmarket Gold used the assistance of internal employees to assist with the generation of this technical report. Table 3-1 is a summary of those roles and the areas they were responsible for within this technical report.

Area of Contribution Site Expert Sections
Metallurgy and Recovery Chris Buda
Union Reefs Plant Manager
13 & 17
Geological Review Wess Edgar
Senior Exploration Geologist
7, 8, 9, 10, 11, 12 & 14
Mineral Resource Review Andrew Lindsay
Geological Superintendent
11 & 14
Environmental Studies Paul McHugh
Environmental Manager
4 & 20
Contracts and Financial Considerations Adam Barnett
Finance Manager
19, 21 & 22
Mining Economics and Costs Wayne Chapman
Technical Services Manager - Stawell
15 & 16
Mining Economics and Costs Dan Hennessey
Technical Services Superintendent
15 & 16

TABLE 3-1 SITE EXPERTS WHO CONTRIBUTED TO THE TECHNICAL REPORTS

The Authors have reviewed all such information provided by the internal employees and determined it to be adequate for the purposes of this technical report. The Authors do not disclaim responsibility for this information.

3.1 LEGAL ISSUES – AGREEMENTS, LAND TENURE, SURFACE RIGHTS, ACCESS & PERMITS

With respect to Sections 4 and 20, the Independent Author (Murray Smith) has not researched property ownership information such as tenement ownership or status, joint venture agreements, surface access or mineral rights and has not independently verified the legal status or ownership of the Property. With respect to Sections 4.2 and 4.4 of this technical report, the Authors have previously relied upon tenement information and legal opinions provided to Newmarket Gold by their independent Tenement Management Consultants based in Darwin. The consultancy group is called Complete Tenement Management but the information in Sections 4.2 and 4.4 was prepared by individuals who are not Qualified Persons as defined by National Instrument 43-101. Advice was given on these sections by Complete Tenement Management in March 2013 for a previous technical report. There have not been any significant changes to tenement regulations since 2013 so the informations has been deemed by the Author as being current.

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Technical Report Newmarket Gold Inc.
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Wherever possible, Newmarket Gold gains the assistance of legal counsel on matters requiring expert opinions. This is generally done using legal counsel based in Darwin, who has a sound and working knowledge of all local and federal legislation. These legal groups are also used to assist with generating agreement and contacts whenever required.

3.2 HISTORICAL INFORMATION

Information relating to historical exploration, production and mineral resources and mineral reserves, mining and metallurgy has in part been sourced from summary documentation prepared by past operators and Newmarket Gold, from previously filed NI 43-101 technical reports and corporate filings and press releases available on the System for Electronic Document Analysis and Retrieval (SEDAR) website: www.SEDAR.com and from other public sources. Where required the source of this information has been noted in this technical report. These historic reports would be previously lodged under Crocodile Gold.

Interpretations and conclusions contained herein reflect the detail and accuracy of historical exploration data available for review. Given the nature of mineral exploration, and with more detailed modern exploration work and new exploration and mining technology, more precise methods of analysis and advances in understanding of local and regional geology and mineral deposit models over time, the interpretations and conclusions contained herein are likely to change and may be found to be in error or be obsolete. As part of Newmarket Gold’s ongoing process to improve mineral resource estimates, all mining information is reconciled against the models to ensure accuracy; this assists in improving the accuracy of the models.

3.3 ENVIRONMENTAL ISSUES

The Authors are not experts in the assessment of potential environmental liabilities associated with these properties and no opinion is expressed regarding the environmental aspects of these properties. Liabilities for this project are summarized in Section 20.5 of this report, including an estimation of the closure costs associated with current mining activities.

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Technical Report Newmarket Gold Inc.
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4 PROPERTY DESCRIPTION AND LOCATION

The Northern Territory Property described within this technical report is located within the Pine Creek region of the Northern Territory of Australia (Figure 4-1).

4.1 LOCATION

The Northern Territory Property is comprised of 141 mineral titles (133 granted and eight applications covering 22,400Ha) and 23 exploration titles covering a total area of approximately 1,806.41km 2, as follows:

License Type Number Area (km²)
Exploration License
Exploration License (EL) 24 1,791.58
Exploration License Application (ELA) - -
Exploration License Retention (ELR) 2 14.83
Sub Total 26 1,806.41
Mineral Leases
Mineral Claim (MC) 55 11.66
Mineral Lease (ML) 72 191.77
Mineral Lease Application (MLA) 8 20.11
Mineral Authority (MA) 6 0.82
Sub Total 141 224.36
Total 167 2,030.77

TABLE 4-1 SUMMARY OF MINERAL TITLES NEWMARKET GOLD NT OPERATIONS

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Technical Report Newmarket Gold Inc.
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Geographically, the Property is centered north of Pine Creek, a small village which historically was an important gold mining center, on the Stuart Highway, 248km south-southeast of Darwin, the capital city of the Northern Territory (population 136,200), at Latitude 13°49’24”S, Longitude 131°50’05”E and UTM (“AMG”) coordinates (WGS-84, Zone 52L) 806,474mE and 8,469,997mS, elevation 208mASL.

The Property is located between Adelaide River (population 237) and Pine Creek (population 380), 125–248km respectively south-southeast of Darwin (Figure 4-2).

The Union Reefs mineral resources and processing facility is located on MLN1109. This title was granted on the December 16, 1993 for a period of 23 years. It was renewed in 2015 for a period of 19 years.

The Cosmo Mine is located with a converted local grid (mine grid). All drill collars are stored within the drillhole database within the mine grid co-ordinates. The mine grid is rotated approximately 45o to the UTM grid. The conversion from the local mine grid to UTM co-ordinates can be seen in Section 10.

4.2 MINERAL RIGHTS, MINING LAWS AND REGULATIONS

Mineral Rights in the Northern Territory of Australia are governed by the Mineral Titles Act 2015 (the Act).

Exploration for minerals and the extraction of minerals and extractive minerals (sand, gravel, rocks, peat and soil) may only occur by title holders who are authorized to do so under the Act by the grant of a mineral title. 

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The Mining Management Act 2015 provides for the management of operational activities on exploration and mining sites.

Exploration and mineral titles on Aboriginal freehold land are subject to the provisions of Part IV of the Commonwealth Aboriginal Land Rights (Northern Territory) Act (ALRA); other parts of the Territory may be subject to the provisions of the Commonwealth Native Title Act (Native Title Act).

Administration of these acts is the responsibility of the Minerals Titles Group of DME.

4.2.1      MINERAL RIGHTS

The Minister for Mines and Energy is responsible for the Mineral Titles Act, which is administered on his behalf by the DME.

The Act provides a legislative framework for the management of the application, granting and maintenance of mineral exploration and mineral titles in the Northern Territory. The primary function of the Act is the administration and regulation of exploration and mineral titles. Originally the Act also contained provisions for the management of operational activities on exploration and mining sites; however, these provisions were more recently incorporated into the Mining Management Act.

Both the Act and the Mining Management Act are supported by regulatory legislation, the NT Mineral Titles Regulations and NT Mining Management Regulations, respectively, which enable the DME to administer the industry.

Other relevant legislation applicable to the exploration and mining industry operating in the Northern Territory includes:

A major policy objective of the Northern Territory Government is to ensure that the maximum amount of land is being actively explored and mined at any one time. The Mineral Titles Act includes a number of provisions that attempt to encourage the active exploration and mining of commodities as well as providing equitable opportunities to access land for large, medium and small enterprises.

The primary vehicle for mineral exploration in the Act is the exploration license, a title that provides for systematic exploration and regular reductions of the title area so as to provide for a turnover of land available for exploration purposes.

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Technical Report Newmarket Gold Inc.
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The Act also contains provisions relating to mineral leases for conducting mining activities. Where the mining of material is required for construction materials, such as road works or material for concrete manufacture, the Act provides for titles such as extractive mineral permits and extractive mineral leases to be issued.

Apart from the Act, the form and operations of the mining industry are shaped by a variety of legislation and policies such as the Territory’s Parks and Wildlife legislation and Sacred Site legislation, the ALRA and the Native Title Act. One of the objectives of the Act is to ensure that it is compatible with these and other relevant legislation.

Central to the establishment and operation of a mineral resource industry in the Northern Territory is the ability to access land in a transparent, equitable, timely and cost effective manner. The Northern Territory Government’s Multiple Land Use Policy means that all land is potentially available for exploration and mineral production.

The ability to deal in exploration and mineral titles is integral to a successful mining industry. The Act provides for these transactions and for a register of legal transactions to be maintained as a matter of public record.

Additionally, the Act also provides mechanisms for the consideration of submissions and objections to the grant of titles and the resolution of disputes through the Lands Planning and Mining Tribunal.

The Mineral Titles Act is also able to exclude land from the general provisions of the Act for the purposes of either temporarily or permanently prohibiting exploration and/or mining on a particular area or to provide for controlled development of that area.

4.2.2      TITLES

Exploration and mineral titles in the Northern Territory are administered in accordance with the provisions of the Act and the Mineral Titles Regulations. Applications for mineral titles are made in accordance with the Act and where the underlying land is Aboriginal Freehold land or land that is subject to native title, the applicant must also follow additional processes.

The Act has a variety of title categories to provide for a range of activities from low level and non-intrusive exploration to major mining projects.

The principal forms of mineral tenure that are issued under the Act are summarized below and elaborated upon in the section following (Table 4-2):

Exploration License (EL): Provides exclusive rights for the holder to undertake exploration activities within the license area and to apply for a mineral title.

Exploration License in Retention (ELR): Grants the holder the right to retain an area of land under title where there is evidence of a mineralization body or anomalous zone of possible economic potential, which requires further assessment. This assessment may involve the conduct of further exploration, feasibility studies or waiting for market and economic conditions to change before production commences.

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Mineral Lease (ML): Provides legal title for the mining of minerals. Generally used for substantial mining operations and may be used for constructing related mining infrastructure. (Previously known as MLN).

Mineral Claim (MC): Provides legal title for the mining of minerals, principally for small miners. (Previously known as MCN). These titles are classified under the Mineral Titles Act (2015) as non-compliant titles, the process to convert these titles to compliant titles is outlined within the act and has progressed for Newmarket Gold over the past 12 months.

Mineral Authority (MA): The Minister may declare mineral reserves in the Territory that exclude certain land from particular exploration or mining activities. Exploration for, or mining of, a particular mineral may be excluded for example. The declaration of a mineral reserve, however, does not necessarily mean that the land is completely excluded from exploration and mining activities. An (MA) may be declared by the Minister in respect of general mineral reserved land. An MA is a mineral title that corresponds to a mineral title that may be otherwise granted under the Mineral Titles Act.

Extractive Mineral Lease: Provides title for the larger scale mining of extractive minerals by quarrying or other means.

Extractive Mineral Permit: Provides title for shorter term or smaller extractive operations.

Extractive Mineral Exploration License: Provides title for shorter term or smaller extractive operations.

Exploration licenses and mineral leases are the predominant titles in the Northern Territory.

4.2.3      EXPLORATION LICENSES

An exploration license can be granted to explore up to 250 graticular blocks of land, or approximately 805km2. Exploration licenses may be granted for periods of up to six years and may be renewed for periods, of two year terms. Applications for renewal must be made prior to the expiry of the exploration license.

Exploration licenses are subject to regular size reductions. Those reductions occur at the end of years two, four and six. The license area must be reduced by half its previous size on each reduction, or a waiver may be requested. As part of the annual review process, explorers are required to report both technically and expenditure annually on their exploration programs.

Prior to commencing any substantial disturbance, explorers are required to obtain an authorization under the Mining Management Act.

4.2.4      MINERAL LEASES

A mineral lease can be unlimited in area (recorded as hectares) and may be granted for the period of the mine with renewal options. The lessee is authorized to explore and mine for minerals on the lease area subject to other legislation such as the Mining Management Act. A mineral lease may be issued for other purposes as specified in a lease document such as constructing related infrastructure. A mineral lease may also be issued for ancillary purposes in conjunction with mining of minerals.

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Technical Report Newmarket Gold Inc.
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The Mining Management Regulations provide administrative procedures for management of the Act.

4.3 ADMINISTRATION

Exploration and mineral titles in the Northern Territory are administered by the Titles Division within the Minerals and Energy Group of the Northern Territory Department of Mines and Energy (DME). In addition to administering the Act and Regulations the Division manages the procedures associated with the ALRA and the Native Title Act.

The Authorizations and Evaluations and Compliance Divisions of the Minerals and Energy Group also administer the requirements of the Mining Management Act and Regulations for operational activities within the Northern Territory. These activities include exploration activities such as drilling and bulk sampling which are defined as causing substantial disturbance. These divisions grant authorizations for operational activities, manage rehabilitation securities, regulatory reporting and audit environmental performance. They also conduct mine audits and inspections to ensure compliance with Mining Management Plans and relevant standards. Compliance issues relating to occupational health and safety are administered by Northern Territory Worksafe.

4.4 MINERAL TENURE

4.4.1      MINERAL TENURE - BURNSIDE (INCLUDING COSMO MINE)

The Burnside area contains several listed mineral resources and the Cosmo Mine. This project area consists of 107 Mineral Leases in the forms of ML’s, MLN’s and MCN’s, all of which are described in more detail in Section 4.2.2 above. The Cosmo Mine Mineral Lease is MLN993, which was recently renewed in 2012 for a period of 10 years. In the Burnside deposits there are also 15 Exploration Licenses covering a combined area of 1,012km2.

The Mineral Lease title allows the owner to conduct mining activities once the Government has approved a Mine Management Plan (MMP). During the assessment of the MMP, the Government will request that a security bond be paid to cover the cost of future mining disturbance. The MMP will also set in place the requirements and scope of work to be completed. These MMP’s are updated annually for all mining activities. More details for the MMP process is outlined in Sections 4.2 and 20.

Under the terms of the Act, NT Mining Operations Pty Ltd is the nominated operator of the mining operation. The Mineral and Exploration Leases are owned by Newmarket Gold NT Holdings Pty Ltd. Both companies are 100% owned by Newmarket Gold Inc.

At the Cosmo Mine there is another lease, which forms part of the Cosmo Mine area (MLA27938) this is a Mineral Lease under application with the DME. Newmarket Gold believes (but cannot guarantee) that this lease would be granted within the next 12 months. The lease was applied to encompass the Cosmo Village mine camp. Currently the camp is located on Exploration License EL25748, which is owned by Newmarket Gold.

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Technical Report Newmarket Gold Inc.
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Also within the Burnside deposits area are 17 Exploration Licenses, which include both EL’s and one ELR for the Western Arm Deposit. All EL’s are owned by Newmarket Gold but are included in the PNX Metals Farm-in agreement, which is described in more detail in Section 4.5.3.

Licence Type Number Area (km²)
Exploration Licence
Exploration Licence (EL) 16 1,005.64
Exploration Licence Retention (ELR) 1 6.52
Sub Total 17 1,012.16
Mineral Leases
Mineral Claim (MC) 44 9.02
Mineral Lease (ML) 56 87.53
Mineral Lease Application (MLA) 7 19.53
Sub Total 107 116.08
Total 124 1,128.24

TABLE 4-2 SUMMARY OF MINERAL TITLES BURNSIDE (* MINERAL LEASES ARE INCLUDED IN EXPLORATION LICENSES)

4.4.2      MINERAL TENURE - UNION REEFS

The Union Reefs deposit consists of 18 Mineral Leases including the newly granted ML27999, which covers the Esmeralda mineral resource and mineral reserve. The Union Reefs processing facility is located on MLN1109, which was renewed in 2015 for a period of 19 years.

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Technical Report Newmarket Gold Inc.
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Also within the Union Reefs deposit are four Exploration Licenses covering 79.92km 2. These leases are all owned and operated by Newmarket Gold. These titles are not part of any Farm-In agreement and are maintained by Newmarket Gold.

Licence Type Number Area (km²)
Exploration Licences
Exploration Licences (EL) 4 79.92
Sub Total 4 79.92
Mineral Leases
Mineral Claim (MC) 5 0.26
Mineral Lease (ML) 8 47.8
Mineral Authority (MA) 5 0.79
Sub Total 18 48.85
Total 22 128.77

TABLE 4-3 SUMMARY OF MINERAL TITLES - UNION REEFS (* MINERAL LEASES ARE INCLUDED IN
EXPLORATION LICENSES)

4.4.3      MINERAL TENURE - PINE CREEK

The Pine Creek area of Newmarket Gold comprises a total of seven mineral titles (all granted) and one Exploration title (granted) covering a total area of approximately 172.47km 2, as outlined below. The mineral resources and mineral reserves for Pine Creek are located on MLN13 and MLN1103, which are due to expire in 2030.

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Technical Report Newmarket Gold Inc.
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Licence Type Number Area (km²)
Exploration Licence
Exploration Licence (EL) 1 163.86
Sub Total 1 163.86
Mineral Leases
Mineral Lease (ML) 3 8.58
Mineral Authority (MA) 1 0.03
Sub Total 4 8.61
Total 5 172.47

TABLE 4-4 SUMMARY OF MINERAL TITLES PINE CREEK (* MINERAL LEASES ARE INCLUDED IN EXPLORATION LICENSES)

4.4.4      MINERAL TENURE - OTHER PROJECTS

Other areas owned by Newmarket Gold include the Maud Creek project, which is currently part of a standalone PEA review completed by SRK (Australia) Pty. Ltd., which is due for release in 2016. Other projects include the Moline area, which is located to the east of Pine Creek and the Yeuralba area, which is located northeast of the Maud Creek project. These two areas do not contain any mineral resources or mineral reserves.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

The Moline are is include in the PNX Metals Farm-In agreement and is currently managed by their team. Parts of the Maud Creek project are part of the same agreement (excluding the Mineral Leases containing the mineral deposit).

The other projects outside of the reported mineral resources and mineral reserves in this technical report consist of 11 Mineral Leases (ML’s, MLN’s MCN’s and MA’s) covering 50.24Ha, and three Exploration Leases covering 542.16km 2.

Licence Type Number Area (km²)
Exploration Licence
Exploration Licence (EL) 3 542.16
Sub Total 3 542.16
Mineral Leases
Mineral Claim (MC) 6 2.38
Mineral Lease (ML) 5 7.86
Sub Total 11 50.24
Total 14 592.4

TABLE 4-5 SUMMARY OF MINERAL TITLES FOR NT OPERATIONS OUTSIDE BURNSIDE, UNION REEFS AND
PINE CREEK (* MINERAL LEASES ARE INCLUDED IN EXPLOR ATION LICENSES)

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

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Technical Report Newmarket Gold Inc.
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4.5 AGREEMENTS

4.5.1      JOINT VENTURES & OPTION AGREEMENTS

The NT Operations project is subject to a number of “farm-out” Joint Venture Agreements on certain areas (see Figure 4-6), the terms of which are summarized in Section 4.5.2 below.

4.5.2      FARM OUT AGREEMENTS

“Farm-out” agreements provide for third parties to explore on mineral titles, which are not owned 100% or substantially controlled by Newmarket Gold.

On November 6, 2013, Thundelarra Exploration Limited Uranium Exploration (Thundelarra) withdrew from a joint venture agreement with Crocodile Gold. Thundelarra was replaced by Rockland Resources Pty Ltd (Rockland) as party to the joint venture agreement; a 100% owned subsidiary of Oz Uranium Pty Ltd. Rockland was then replaced as a party to the agreement with Oz Uranium Exploration Agreement for the Pine Creek Tenements. Rockland Resources Pty Ltd (Rockland), a wholly-owned subsidiary of Oz Uranium, and Crocodile Gold (now Newmarket Gold) formed a joint venture on November 6, 2013, in regards to uranium exploration and development on the Maud Creek, Burnside, Cosmo, Pine Creek, Union Reefs and Moline projects. Rockland has a minimum expenditure commitment of $1 million over the next four years. Rockland has the rights to apply for a mineral tenement in its own right as long as it does not conflict with Newmarket Gold’s operations.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Over the past 24 months Rockland has been active in the Pine Creek region. They have conducted regional scale geophysical surveys and reviews (VTEM) as well as geochemical analysis, structural mapping and drilling in and around their currently identified uranium deposits. While one prospect is close to the Cosmo Mine (Fleur de Lys) no work has been conducted by Rockland on MLN993.

4.5.3      FARM IN AGREEMENTS

In 2014, Phoenix Copper Pty Ltd (now PNX Metals) entered into a “Farm-in” agreement with Crocodile Gold. The Heads of Agreement was signed in August 2014 and was completed in December 2014. The “farm-in” agreement relates to exploration activities on the Burnside Exploration Licenses as well as at the Chessman (close to Maud Creek) and Moline projects.

The Farm-in Tenements include the Burnside Exploration Titles (ELR97 and exploration licenses EL10012, EL10347, EL23270, EL23431, EL23536, EL23540, EL23541, EL24018, EL24051, EL24058, EL24351, EL24405, EL24409, EL24715, EL25295, EL25748, and EL9608), the Maud Creek Project including exploration licenses EL25054 and EL28902, and mineral lease ML30293 and the Moline Project including exploration license EL28616, and mineral leases ML24173 and MLN’s1059 and 41. In 2015 the title ELR97 was removed from the agreement by mutual consent of Newmarket and PNX Metals.

The “Farm-in” agreement will allow Phoenix to earn up to 51% through the spending of $A2 million on exploration activities over a two year period. They can then earn a further 39% by spending an additional $A2 million for another two year period. This will potentially take their ownership of these projects to 90%. Newmarket Gold retains a claw-back right to precious metal discoveries. While this agreement is not over the Cosmo Mine area, it covers the exploration licenses that surround the Cosmo Deposit.

PNX Metals has been active since signing the Heads of Agreement in August 2014.

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Technical Report Newmarket Gold Inc.
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4.5.4      OTHER AGREEMENTS

Within the past year the Company has divested tenements in the Glencoe, Redbank, Iron Blow-Mt Bonnie areas as well as the tenements that host the Bridge Creek deposit in the Burnside area.

The Bridge Creek tenements included MLN’s 1060 & 766, MCN’s 4293-429 and MCN’s 4956-4958. The agreement only applies to non-alluvial mining operations. Newmarket Gold retains a 1% NSR on any mineral production from the leases.

Newmarket Gold divested tenements in the Iron Blow – Mt Bonnie area to Phoenix Copper Limited (now PNX Metals). Tenements included MCN’s 3161, 504 and 505, MLN’s 1033, 1039, 214, 341, 342, 343, 346, 349, 40, 459, 811 and 816. Newmarket Gold retains a 2% NSR on any gold and silver produced from the leases.

4.6 SURFACE RIGHTS – LAND ACCESS

4.6.1      PASTORAL LEASES

The Northern Territory Operations area is located on a pastoral lease. Holders of mineral and exploration tenements have rights of access to their tenements, including access through neighboring pastoral leases, and are not obligated to remunerate pastoral leaseholders for recovered minerals because by law they do not have any title to the minerals on the tenements. However, as a matter of commercial practice, mining companies and pastoral leaseholders often reach access agreements governing their activities and relationships.

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Technical Report Newmarket Gold Inc.
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Tenements comprising the Cosmo Mine are located on a single pastoral lease, as follows:

Tenements comprising the Union Reefs area are located on various pastoral and Crown leases, as follows:

Tenements comprising the Pine Creek area are located on various pastoral and Crown leases, as follows:

Relations with each pastoral lease owner and/or operator are reported to have been harmonious and regular communications are reportedly maintained with the lease operators for the active mining and exploration areas. No formal agreements exist with the pastoral leaseholders, although historically past operators prepared draft agreements for submission to the pastoralists at various points in time.

4.7 OPERATING AUTHORIZATIONS

The mining regime in the Northern Territory is governed by the Mining Management Act (2015) (the Mining Act). In accordance with this legislation, the owner of an exploration license or mineral lease is required to submit MMP to DME. This plan, covering key aspects of mine operations and exploration activities, health and safety, environmental management and mine closure is assessed and audited by DME. Upon approval of the MMP, an Authorization to Operate for a 12-48 month period is issued to the mining operation. Depending on the status of operating or exploration activities involved, the MMP can be a relatively simple or detailed document akin to a Notice of Intent (NOI).

Newmarket Gold develops MMPs for all Mineral Leases as required under the Mining Management Act, this process is ongoing and is required before mining, or exploration such as drilling, can commence. In places where the MMP has expired, the DME can allow mining to continue while the new MMP is approved. Newmarket Gold currently has an MMP in place for the Cosmo Mine, which is updated annually. More details on this process can be found in Sections 4.11and 20.

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Technical Report Newmarket Gold Inc.
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4.8 MISCELLANEOUS LICENSES & ACCESS

Access to the Northern Territories Property is generally through Newmarket Gold-managed tenements and/or access roads.

Several of the properties are subject to infrastructure easements; including the Darwin and other gas supply pipelines, the Adelaide-Darwin Transcontinental Railway line, telecommunication towers and overhead electricity power supply lines and equipment. These easements are located throughout the tenement holdings but generally do not impact directly on any mineral resources or mineral reserves reported in this technical report. The only exception is the Amadeus Gas Pipeline, which is located in close proximity to the Esmeralda deposit south of Union Reefs. There is an exclusion zone around the pipeline and restrictions to mining around the pipeline. These factors have been included in the mineral reserve estimations for the deposit and Newmarket Gold continues to work with the owners of the pipeline to ensure all regulations are complied with.

4.9 NATIVE TITLE

The majority, i.e. over 96%, of the current mineral resource and mineral reserve base within the Northern Territory Gold Properties lies within granted mineral leases for which Native Title has been extinguished. Hence, Native Title issues will not affect the development and operation of mining operations within these project tenements. This excludes the Esmeralda deposit which has an active land use agreement in place.

The Cosmo Mine is located on a Mineral Title where Native Title has been extinguished. Native Title is an issue for MLA27938, however, this is a Mineral Lease marked for infrastructure and not mining so any agreement reached will not affect mining activities at Cosmo Mine. This agreement is required before the title can be granted and Newmarket Gold is working with the Northern Land Council in conjunction with the Traditional Owners to finalize the agreement. The Authors see no significant issues to having this agreement finalized in the coming 12 months, which will allow the DME to continue with the approval process for the Mineral Lease application. This process has no influence over MLN993, which was granted prior to Native Title being granted. Details of this process are outlined below.

No sites of Aboriginal or other historical significance have been located or documented for the project area.

4.9.1      NATIVE TITLE PROCESS - SUMMARY

Native Title is a complicated issue and the Authors are not experts in this area; the information below, which is extracted from Northern Territory Government websites and is provided for information purposes, may not be complete, accurate or current, and is presented subject to the disclaimer provided in Section 3, above.

4.9.2      EXPLORATION & MINING ON NATIVE TITLE AFFECTED LAND

Application for exploration and mineral title may, depending on the underlying land tenure, be required to comply with the Native Title Act prior to the grant of a title.

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Typically, compliance with the Native Title Act is required where an application is over Pastoral Lease or Crown Lease land.

The Native Title Act provides an option of seeking an expedited right to negotiate procedure for the grant of exploration licenses compared to the right to negotiate procedure that applies to mining tenure. Applicants may also enter into Indigenous Land Use Agreements (ILUA) with Native Title parties to facilitate tenure grant.

4.9.3      EXPEDITED PROCEDURE

In the Northern Territory, applications for the grant of an exploration license are generally required to comply with the expedited right to negotiate procedure, which provides a faster route for the grant of exploration title that have lower impact.

The Native Title Act defines an act attracting the expedited procedure, as one that is not likely to interfere with Indigenous community or their social activities, significant sites, or involve major disturbance to land or waters.

The expedited procedure is activated when the notification process includes a statement that the government “considers the act of granting the exploration license is an act attracting the expedited procedure”.

Registered Native Title claimants may object to the inclusion of this statement within the four month notification period. If the objection is not withdrawn after a period of negotiation the matter is required to proceed to arbitration. In the Northern Territory, the National Native Title Tribunal (NNTT) is the arbitral body, which handles the expedited procedure objection inquiry.

Agreements, which allow the objection to be withdrawn, may be reached at any stage during the expedited procedure. The Northern Territory Government encourages such agreements.

The Northern Territory Government has successfully used the expedited procedure for the grant of exploration licenses. This success is largely the result of additional conditions placed on exploration holders to further protect the rights and interests of Native Title holders and the requirement for exploration license holders to comply with Northern Territory’s Aboriginal Sacred Sites Act.

4.9.4      RIGHT TO NEGOTIATE PROCEDURE

Applications for all forms of mining tenure, on which development may occur, are required to comply with the right to negotiate procedure.

This procedure is commenced by a public notification process in which details of the mineral tenement applications are placed in a Northern Territory and an Indigenous newspaper.

If a Native Title claim is lodged and registered within four months of the notification date, it is a requirement of s31(1)(b) of the Native Title Act that an agreement be reached, formalized by the execution of a Tripartite Deed prior to the grant of title.

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This Deed between the Northern Territory Government, the Native Title parties and the applicant will generally be supported by an Ancillary Agreement between the title applicant and the Native Title parties.

A feature of the Native Title Act is the requirement for negotiation to be carried out in good faith.

The Northern Territory Government, through the DME Titles Division, plays an active part in managing the right to negotiate procedure.

If the negotiating parties are unable to reach agreement the matter may be referred to the NNTT for either mediation or arbitration.

4.9.5      INDIGENOUS LAND USE AGREEMENT (ILUA)

Applications for exploration and mineral tenure may also be granted where the applicant/s and the Native Title Representative Body enters into an ILUA.

There are a number of ILUAs registered under the Native Title Act for exploration and mining within the Northern Territory. These are flexible agreements that can provide for various activities including exploration and mining activities, suitable for small exploration or large mining projects.

4.9.6      EXPLORATION & MINING ON ABORIGINAL FREEHOLD LAND

Some 44% of the Northern Territory is Aboriginal freehold land and subject to the ALRA. Under that Act, Land Councils administer this land on behalf of the traditional owners. There are four Land Councils in the Northern Territory, the Northern Land Council, the Central Land Council, the Tiwi Land Council and the Anindilyakwa Land Council.

4.9.7      THE MINING ACT PROCESS

Applications for exploration and subsequent mining titles on Aboriginal freehold land are required to comply with the Northern Territory Mining Act. It is a requirement of the ALRA that a miner seeking to explore on Aboriginal freehold land initially applies for an exploration license.

Applications for exploration licenses must be made through the Department of Regional Development, Primary Industry, Fisheries and Resources (DME) Titles Division office. Following receipt of an exploration license application the Department will assess the application to ensure legislative compliance and the adequacy of the exploration proposals. The application is also subject to a public notification process. On completion of this initial review process, the Northern Territory Minister for DME may issue consent to negotiate. This consent activates the mining processes under Part IV of the ALRA.

4.9.8      THE ALRA PROCESS

Once consent is issued, the applicant is required, to develop and lodge an exploration proposal with the relevant Land Council within three months. These proposals must contain details of proposed exploration activities and details of the method of extraction and treatment of any commodity that may be discovered, as required by s41(6) of the ALRA.

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Guidelines on developing proposals are available from the relevant Land Council. The DME has also published a booklet titled Exploring Country to assist with exploration and mining agreements. This can be accessed on line.

Once the proposal is accepted by the Land Council, the parties, the applicant and the Land Council, have an initial 22 month negotiation period in which to reach an agreement. During the process, the Land Council is required to consult with the traditional owners.

The consultation process may include the convening of one or more meetings with traditional owners.

Applicants are entitled to attend certain meetings for the purpose of explaining and discussing the proposed exploration activities.

The Department of Business, Economic and Regional Development’s Indigenous Business Industry Services Branch is available for guidance on how to best present this material in a culturally sensitive manner.

4.9.9      REACHING AGREEMENT

When an agreement is reached, it is a requirement of the ALRA that consent is given by the Land Council and the Federal Minister for Families, Community Services and Indigenous Affairs. The agreement and the consents are required to be submitted to DME, following which the exploration license can be granted.

Following grant, the Department administers the exploration license in accordance with the Mineral Titles Act.

4.9.10      NEGOTIATING TIMEFRAMES

Under the ALRA, negotiation towards agreement is to be carried out within prescribed timeframes.

If an agreement is not reached within the standard negotiating period, there is provision for extension to the negotiation period by agreement between the Land Council and the applicant. The first extension is for a two year period, followed by periods of one year. The standard negotiation period commences upon lodgment of the proposals with the Land Council and ending 22 months from the January 1st following the date of lodgment.

4.9.11      CURRENT NATIVE TITLE AGREEMENTS

4.9.11.1  Esmeralda Land Use Agreement

This Agreement has been negotiated between Crocodile Gold, The Northern Land Council (NLC) and representatives of the Wagiman, Warai and Jawoyn peoples. This agreement was required for the granting of Mineral Lease ML27999, which covers the Esmeralda deposit to the south of Union Reefs processing plant. This agreement was signed in 2015 with the terms now active after the granting of ML27999.

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Technical Report Newmarket Gold Inc.
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4.9.11.2  Kazi Land Use Agreement

The Agreement is being negotiated between Newmarket Gold, The Northern Land Council (NLC) and representatives of the Warai, Kungarakan, Wagiman and Kamu peoples. This agreement is required for the granting of several Mineral Leases;

This agreement remains in negotiation between Newmarket Gold, the NLC and traditional owners. More work is required to complete this agreement. Newmarket continues to hold discussions with the NLC over this agreement.

4.10 ROYALTIES

All tenements within the Northern Territory, Australia are subject to a Northern Territory Government Minerals Royalty in accordance with the Northern Territory Mineral Royalty Act 1982 (as amended) (Mineral Royalty Act). This royalty is calculated as 20% of the “Net Value” of mine production, where “Net Value” equals the gross revenue from the relevant production unit less the operating costs of the production unit for the year, a capital allowance on eligible capital assets expenditure, eligible exploration expenditure and additional deductions as approved by the Northern Territory Minister for Mines.

Royalty calculations are detailed below:

From 1 July 2010 the Mineral Royalty Act levies royalty at a rate of 20 per cent (prior to 1 July 2010 the rate was 18 per cent) of the Net Value of mineral commodities sold or removed from a production unit, regardless of the type of mineral commodity or whether the mine is situated on Crown, freehold, leasehold or aboriginal land. Net Value is calculated as follows:

Net Value = GR – (OC + CRD + EEE + AD)

Where: –

GR is the Gross Realization from the production unit;
OC represents the Operating Costs of the production unit for the royalty year;
CRD is the Capital Recognition Deduction on eligible capital assets expenditure;
EEE is any Eligible Exploration Expenditure; and
AD represents Additional Deduction as approved by the Minister.

A "production unit" is a mineral tenement of two or more mining tenements operating as part of an integrated operation. It also extends to other facilities (whether or not adjacent to the mineral tenements) that are essential for the production of a saleable mineral commodity.

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Technical Report Newmarket Gold Inc.
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Net value for royalty is thus defined as the value of minerals sold or removed without sale plus an adjustment for assets disposed of, less

Furthermore, the first $50,000 of Net Value is not liable to royalty. This exempts a number of small mines from royalty payment entirely.

Royalty is payable by six monthly provisional payments. An annual return detailing the actual royalty payable together with payment for any additional liability must be lodged within three months after the end of each royalty year. Interest applies for late payment and further penalties may apply if the sum of the provisional payments is less than 80 per cent of the actual royalty payable.

The Authors have not reviewed or investigated individual title information and this information, which may not be complete, accurate or current, and is presented subject to the disclaimer provided in Section 3, above.

4.10.1      UNION REEFS AREA ROYALTIES

A vendor royalty of 1.5% if mining for the purposes of commercial production of gold commences on 10 tenements held by Newmarket Gold in the Union Reefs area is payable to the estate of Robert Michael Biddlecombe. The Royalty is not payable on gold mined from the main Union Reefs mineral lease (MLN1109) but leases to the north around the historical Elizabeth Mine. The tenements that this royalty apply to are MCN’s 734, 506, 507, 735, 738, MLN’s 779, 135, 779, 780, 882 and 856.

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Technical Report Newmarket Gold Inc.
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Project Parties Involved Royalty
Commitment
Tenements Comments Paid to
date
Union
Reefs
Newmarket
Gold
The estate of
Biddlecombe
1.5% Gross Royalty
for all minerals
MCN's 734, 506-7 MCN's
735, 738 MLN's 135, 779-
780 MLN's 856, 882
Elizabeth
north of
Union Reefs
No
Esmeralda Newmarket
Gold
NLC 2.0% Gross Royalty
for all Minerals
ML27999 Required for land use agreement No

TABLE 4-7 LIST OF UNION REEFS ROYALTY’ S CURRENTLY REQUIRED BY NEWMARKET GOLD

As part of the land use agreement negotiated between the NLC, the Traditional owners and Newmarket Gold a 2.0% NSR royalty is payable on all minerals produced from ML27999, which contains the Esmeralda deposit.

4.10.2      PINE CREEK AREA ROYALTIES

Vendor royalty of $A4 per ounce of gold produced from certain Pine Creek tenements, payable to a privately owned company, Silver Coin Mining and Prospecting Pty Ltd. Silver Coin Mining is a company formed from local members of the Pine Creek Community.

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Technical Report Newmarket Gold Inc.
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Project Parties Involved Royalty
Commitment
Tenements Comments Paid to
date
Pine Creek Newmarket
Gold
Silver Coin
Mining
$4/ounce produced MLN's 13, 1130
MCN's 317, 4072,
4074 MCN's 523,
1054-1055 EL23583
Pine Creek
deposit
Royalty
No

TABLE 4-8 LIST OF PINE CREEK ROYALTIES CURRENTLY REQUIRED BY NEWMARKET GOLD

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Technical Report Newmarket Gold Inc.
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4.10.3      BURNSIDE AREA ROYALTIES

Franco-Nevada Australia Pty Ltd has a vendor royalty of $A20 per ounce of gold produced and sold from the Brocks Creek underground mine. Royalty payments have been made under this Royalty agreement.

Freeport-McMoran Australia Inc. has a vendor royalty of 1% of gold produced from certain tenements in the Brocks Creek area, which includes the Brocks Creek underground mine; the royalty becomes payable only after recovery of all operating and capital costs involved with the post-1995 development of the Brocks Creek tenements. In late 2012 Crocodile Gold was contacted by a party representing Cyprus. Crocodile Gold and Cyprus as they were looking to sell the royalty for Brocks Creek. Cyprus subsequently agreed to sell the royalty to Freeport-McMoran to which Newmarket Gold is now dealing with for any future royalties. This royalty is payable on mining activities at both Brocks Creek and Rising Tide deposits.

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Technical Report Newmarket Gold Inc.
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Karen On (formally Ben Hall) and Mary and Joseph Groves have a vendor royalty of 3% of gross product from any mining operation on four tenements held by Newmarket Gold in the North Point area.

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Technical Report Newmarket Gold Inc.
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Project Parties Involved Royalty
Commitment
Tenements Comments Paid to
date
Burnside Newmarket Gold Franco-Nevada Australia Pty Ltd $20/ounce gold produced MLN1139 Payable on mining from Brocks Creek Yes
Newmarket Gold Freeport- McMoran Australia Inc 1% NSR on Gold MLN's 1139, 176 MCN's 4689-4697 MCN's 4701- 4703 Payable on mining from Brocks Creek and Rising Tide in 2013
Newmarket
Gold
On and Groves 3% on Gross Product MCN's 46, 47,
49 & 50
Royalty on
Temperance
No

TABLE 4-9 LIST OF ALL ROYALTYIES CURRENTLY REQUIRED BY NEWMARKET GOLD

4.10.4      ENVIRONMENTAL CONSIDERATIONS

For more details on environmental matters, please refer to section 20 of this technical report.

The Northern Territory Operations lies within areas, which have been subject to significant historical mining and mineralization processing activities for over 100 years. This historical activity, like many mining areas worldwide, has left permanent evidence of this activity on the physical landscape and the natural environmental balance may also have affected.

Location of the Operations lies within an environment characterized by low relief, abundant ephemeral and permanent drainage and, particularly closer to the coast, sizeable billabongs and wetlands and a monsoonal wet season with heavy rainfall requires careful management of water, particularly discharge water from mining and milling operations.

Acid rock drainage is an issue at several locations and various systems have been developed to carefully manage this issue.

Newmarket Gold has included environmental management as an integral part of its operations. All exploration activities and mining operations have been performed in compliance with all environmental regulations within a defined environmental management plans. Past operators reported that environmental assessments and project reviews have been completed as required and were thoroughly scrutinized before commencement of operations.

Site rehabilitation and reclamation has also been completed in a number of locations. This is currently an active part of the mining operations with waste dump rehabilitation a part of the daily mining activities. Site rehabilitation is factored into the operation costs for the earth-moving contractor and is required to be completed as soon as areas become available.

All recent mining operations have operated in accordance to MMPs submitted to DME, with various environmental permits in place, particularly including Waste Discharge Licenses (WDLs).

Since Crocodile Gold took over the responsibility of the tenements in November 2009 several steps have been taken to ensure the environmental sustainability of the project. Several historical issues have been noted and Newmarket Gold is in the process of ensuring these legacy issues are managed. An example of this is the activity where Crocodile Gold has treated a legacy stockpile and rehabilitated the area at the Golden Dyke deposit to reduce the impact of weathering of this material on the local environment.

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Technical Report Newmarket Gold Inc.
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There are currently no investigations of breaches of any regulatory regime or are there any current sanctions or restrictions imposed by Government Departments. The Northern Territory Government has a constant review process including site visits. On these visits they inspect current and past mining areas to ensure Newmarket Gold is compliant with the MMPs approved as well as the relevant legislation. To date no major issues have been identified or recorded against the Company.

The Authors are not expert in the assessment of potential environmental liabilities associated with mineral properties.

4.11 ENVIRONMENTAL MANAGEMENT PLAN

Under the terms of the Northern Territory Mining Management Act 20153, existing mining operations in the Northern Territory are required to submit an annual MMP to DME. This plan covers key aspects of mine operation, Occupational Health and Safety, environmental management and mine closure. This plan is then assessed and audited by DME. Upon approval of the MMP, an Authorization to Operate is issued to the mining operation.

Newmarket Gold has submitted annual MMPs for all of its operating and exploration activities, and provided required annual reports to DME and other relevant departments.

Newmarket Gold has MMP’s in place with DME for the Cosmo Mine area under Authorization numbers 0546-03, for the Union Reefs area under Authorization numbers 0539-03.

Unconditional performance bonds totaling $12,221,052 (Table 4-10) for all of the NT Operations area have been lodged with the Northern Territory Government to cover the anticipated cost of the rehabilitation commitments associated with the Project. This was reduced by 10% and replaced with a yearly 1% charge to be used by the Northern Territory Government on rehabilitation of legacy sites in 2015.

Project/Site Authorization
No
Tenements Bonds
Maud Creek 0524-02 EL25054; EL28902; ML30260; ML30293 $107,984
Moline 0525-02 MLN1059; ML24173; EL28616; MLN41 $288,438
Fountain Head 0526-01 MLN4; MLN206; MLN1020; MLN1034; MCN1172; MCN4785 $984,816
Brocks Creek 0528-01 MLN1139 $1,264,915
North Point & Princess Louise 0530-01 MLN823; MLN824; MLN825; MLN826; MLN827; MLN828; MLN829; MLN830; MLN831; MLN832; MLN858; MLN859; MLN860; MLN861; MLN862; MLN863; MLN940; MLN1112; MCN46; MCN47; MCN49; MCN50; MCN624; MCN625; MCN898; MCN899; MCN4432; MCN4434 $1,148,871
Pine Creek 0538-01 MLN13; MLN1130; MCN523; MCN1054; MCN1055; MA416 $538,738
 
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Project/Site Authorization
No
Tenements Bonds
Union Reefs 0539-03 ELR130; MA398; MA399; MA400; MA401; MA402; MCN506; MCN507; MCN734; MCN735; MCN738; MLN135; MLN779; MLN780; MLN822; MLN833; MLN856; MLN1109 $1,511,369
Cosmo Howley 0546-03 MCN377; MCN378; MCN379; MCN380; MCN852; MCN853; MCN854; MCN855; MCN856; MCN857; MCN1014; MCN1015; MCN1035; MCN1231; MCN1232; MCN3099; MCN3100; MCN3101; MCN3102; MCN3103; MCN3104; MCN3105; MCN3106; MCN3107; MCN3108; MCN3109; MCN3110; MCN3111; MCN3112; MCN3113; MCN3114; MCN3115; MCN3117; ML30892; ML30887; MLN809; MLN884; MLN885; MLN886; MLN887; MLN888; MLN889; MLN890; MLN891; MLN892; MLN993; MLN1000; MLN1027; MLN1053; MLN1060; MLN1062; MLN1129; ML27938 $6,375,921
Total     $12,221,052

TABLE 4-10 NEWMARKET GOLD PERFORMANCE BONDS – 2016

4.11.1      MMP - COSMO HOWLEY 0546-03

The Cosmo MMP contains bonding to the value of $6,375,921. The current MMP that is used for the Operations is 0546-03 and is valid for the period 2013-2017. This MMP covers the Cosmo Underground Mine as well as the Howley, Mottrams and Western Arm mineral resources.

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4.11.2      MMP - UNION REEFS 0539-03

The Union Reefs MMP contains bonding to the value of $1,511,369. The current MMP that is used for the Processing Facility is 0539-03 and is valid for the period 2013-2017. This MMP covers all of the Union Reefs mineral resources and mineral reserves such as the Prospect and Esmeralda deposits.

4.11.3      MMP - PINE CREEK 0539-01

The Pine Creek MMP contains bonding to the value of $538,738. The current MMP that is used for the care and maintenance of the Pine Creek deposit is 0538-01 and is valid for the period 2013-2017. This MMP covers all of the Pine Creek mineral resources and mineral reserves such as the International Deposit.

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4.11.4      OTHER NT OPERATIONS MMPS

There are several other MMP’s that are managed and updated by Newmarket Gold, these are summarised below;

0525-02 – Moline. Bonding of $288,438 covering Care and Maintenance. Valid 2014-2018. There are no reported mineral resources or mineral reserves on the Moline Project.

0526-01 – Fountain Head. Bonding $984,816 covering Care and Maintenance activities. Valid 2014-2018. This MMP covers the Fountain Head and Tally Ho mineral resources.

0528-01 – Brocks Creek. Bonding of $1,264,915 covering Care and Maintenance activities. Valid 2013-2017. This MMP covers the Rising Tide mineral resources.

0530-01 – North Point/Princess Louise. Bonding of $1,148,871 covering Care and Maintenance activities. Valid 2013-17. This MMP covers the North Point and Princess Louise mineral resources.

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4.12 WASTE DISCHARGE LICENSE

A Waste Discharge License (WDL) is the formal approval under Section 74 of the Northern Territory Water Act that authorizes and regulates the release of potential contaminants to water in the Northern Territory to ensure environmental protection objectives are met.

The WDL controls the type, quality and quantity of the release and ensures that monitoring and reporting occur on a regular basis. Newmarket Gold currently has 3 active WDL’s for the NT Operations. There is one active for the Cosmo Operation (WDL 180-03), Pine Creek (WDL166-03) and Union Reefs (WDL138-03). The Cosmo license is valid until December 2017, while the Union Reefs and Pine Creek licenses are valid until December 2016.

Waste discharge licenses are not required on any other projects as there is no active discharge from the remaining care and maintenance sites.

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTURE AND PHYSIOGRAPHY

The Northern Territory is the least populated of all areas in Australia. It encapsulates a total area of 1.35 million square kilometers and accounts for 20% of the whole country; however, just 245,100 (Australian Bureau of Statistics 2014) or 1% of Australia's population reside there.

The Territory varies considerably in topography, climate, and infrastructure. The “Top End”, where the Northern Territory Gold Properties are located, is home to the vast Aboriginal Arnhem Land, which includes the Kakadu National Park. The region is dry between April and September, and wet between October and March. During the wet season everything is green and there is no dust; however, the humidity and temperatures are high and access “off road” is difficult.

The center is extremely arid, with greatly varying temperatures and is known as the “Red Center” named because red is the predominant color found in the soil.

Darwin, Capital of the Northern Territory, lies on the coast to the north and provides the majority of infrastructure support and services for the mining industry. The Stuart Highway, which virtually bisects the country, is the main road that leads from Darwin to Alice Springs then on to Adelaide in South Australia.

5.1 TOPOGRAPHY

Generally the topography of the Property area is flat, locally gently undulating near the coast and slightly more elevated and locally rugged towards Katherine at the southern extremity of the Northern Territory Properties.

In the vicinity of Union Reefs, elevations range from 35m to 50m above mean sea level. Drainage is generally to the north to the Timor Sea via ephemeral creeks, streams and gullied tributaries to Mary and Alligator Rivers, two major rivers running north to the coast.

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Further north, the Burnside area is made up of a complex of landforms, which include plains, peneplains, rises and low hills that are built of undifferentiated Paleoproterozoic metasedimentary units. A series of east-west trending hills comprising granite pavements punctuate the plains and are characterized by rocky outcrops and sandy gravely soils. The topography of the area varies from 35m to 300m above sea level.

Numerous ephemeral watercourses including the Adelaide, McKinley and Margaret Rivers drain northwards to the Timor Sea across Newmarket Gold’s tenements.

5.2 ACCESS

Access to the Property is from Darwin, capital of the Northern Territory, which is an important communication and transportation center, with a busy port and international airport providing daily services to other Australian capital cities and several Asia-Pacific destinations.

The Stuart Highway, the area’s major thoroughfare, and the Adelaide-to-Darwin transcontinental railway line bisect Australia in a north-south sense and provide access to Newmarket Gold’s Northern Territory Projects. The Cosmo Mine and Union Reefs plant sites are easily accessed via good all-weather roads and there is excellent road, rail, water and electric power infrastructure available in the region. A major gas pipeline is also located in close proximity to operations.

The Northern Territory Properties lies between the towns of Pine Creek and Adelaide River to the southeast of Darwin. Access is gained to the Cosmo Mine from Darwin by travelling for some 160km along the sealed Stuart Highway, then turning southwest onto Fountain Head West Road for around 2km.

The Union Reefs processing facility is located approximately 185km southeast of Darwin, 15km north of the town of Pine Creek. All Newmarket Gold projects are located in close proximity to the Stuart Highway and can generally be accessed through the use of sealed roads, government operated gravel roads or other minor farm tracks.

5.3 CLIMATE AND VEGETATION

The “Top End” of the Northern Territory has a tropical monsoon climate characterized by two distinct seasonal patterns: the ‘wet’ monsoon and the ‘dry’ seasons. The wet season generally occurs from November through to April and the dry season between May and October. Almost all rainfall occurs during the wet season, mostly between December and March, and the total rainfall decreases with distance from the coast. Mining operations are largely unaffected by normal seasonal conditions. Average annual rainfalls for Darwin average at about 1,713mm, however, this can be quite variable with an extreme in 2011 of 2,680mm.

The mean daily maximum temperature, as recorded at Darwin on the northern coastline, is 31°C in the coolest months of June to August and 33°C in the hottest months of October and November. The mean daily minimum temperature in Darwin range from approximately 19°C (dry season) to 25°C (wet season).

During the wet season, high intensity rainfall events are common, resulting in local flash flooding of ephemeral streams and watercourses. Mining operations are continuous throughout the year; however, during open pit mining activities increased ore stockpiling is undertaken in the lead up to the wet season thereby offsetting the reduced mining movements over that period. Experience has shown that it is best to shut down mineralization hauling during periods of extreme rainfall as damage to haul roads by large trucks may occur quickly.

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The annual evaporation rate remains high throughout most of the Northern Territory, ranging from 2,400mm to 4,000mm per annum. Monthly evaporation exceeds rainfall for eight months of the year at the coast increasing to the whole year inland. It remains relatively high even during the wet season.

Climate gradually moves from seasonally wet tropical in the north to arid in the south, with corresponding changes in landscape, with areas of rocky escarpment and plateau which break a low relief in the north and rocky ridges in the south.

The Northern Territory has a diversity of vegetation that is maintained by its variety of climate and soils. Natural vegetation of the Properties is typical of savannahs of the northern part of Australia, dominated by Eucalypt species with a grassy understory dominated by sorghum species. The Northern Territory is the only area in Australia that does not have conspicuous temperate flora.

In the north, the vegetation is typically tropical savannah (eucalypt woodland and eucalypt open woodland with a grassy understory). This landscape experiences dramatic seasonal changes with intense growth in the wet season (summer) and widespread fires in the dry season (winter). Internationally famous for the tropical wetlands and rugged sandstone escarpments of Kakadu National Park, the wetlands are of importance for conservation, providing breeding areas, habitat and refuge for important wildlife populations.

From the north, a transition area moves from eucalypt woodlands into areas of melaleuca and acacia forests and woodlands and south into the spinifex (hummock grasslands), Mitchell grass (tussock grasslands) and acacia woodlands and shrublands. The vegetation increases in diversity around Alice Springs with areas of mulga, mallee, chenopods, hummock grasslands, small pockets of eucalypt woodlands and salt lakes.

Major land uses are traditional Indigenous uses, nature conservation (including parts of Kakadu National Park and World Heritage Area and Litchfield National Park), urban and other intensive uses and grazing. Approximately 85,000ha have been cleared.

The Property lies within the Pine Creek and Daly River Bioregions. The Pine Creek Bioregion consists of hilly to rugged terrain and is within the tropical monsoonal belt of northern Australia. Dominant vegetation is tropical eucalypt woodlands/grasslands with some eucalypt open forests, melaleuca forests and woodlands and rainforest and vine thickets.

The region has undergone some localized clearing and the major land uses are grazing, nature conservation (including parts of Kakadu National Park and World Heritage Area and Litchfield National Park), traditional Indigenous uses and other intensive uses including horticulture.

The Daly Basin Bioregion consists of gently undulating plains and scattered low plateau remnants and has a tropical monsoonal climate with distinct wet and dry seasons and high temperatures throughout the year. Dominant vegetation is tropical eucalypt woodlands/grasslands and eucalypt open forests. Smaller patches of eucalypt woodlands and melaleuca forests and woodlands are present.

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Technical Report Newmarket Gold Inc.
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The major land use is grazing on native pastures and traditional Indigenous uses with some horticulture, grazing on modified pastures and nature conservation. The region has undergone some clearing (approximately 167,000ha) for these developments.

The Northern Territory Property is characterized by tall, open eucalypt forests, typically dominated by Darwin woollybutt (Eucalyptus miniata) and Darwin stringybark (E. tetrodonta), and woodlands (dominated by a range of species including E. grandifolia, E. latifolia, E. tintinnans, E. confertiflora and E. tectifica), with smaller areas of monsoon rainforest patches, Melaleuca woodlands, riparian vegetation and tussock grasslands.

5.4 LOCAL RESOURCES AND INFRASTRUCTURE

Darwin has a population in excess of 150,245 (Australian Bureau of Statistics 2014) and is the capital city of the Northern Territory. It is the administrative center of the Northern Territory Government and a major transportation hub, with an international airport and deep-water port and the Adelaide to Darwin transcontinental railway terminating at the East Arm port. As it is the largest city in the Northern Territory, Darwin also has excellent schools, hospitals, and retail, commercial and light industrial services.

Darwin is rapidly developing into a significant freight interchange for trade with southeastern Asia. A considerable proportion of consumer and other goods reaching the Northern Territory are brought by road from Queensland or South Australia. The Stuart, Arnhem, Kakadu, Barkley and Victoria Highways ensure high service levels to the Darwin region from the Australian capitals and other regional centers.

Despite its low population, the area between Darwin and Katherine in the Northern Territory is well serviced with infrastructure. Significant mining operations have been developed in the area over the past 30 years, with gold mining and processing operations conducted within or in close proximity to the project areas.

The regional mining communities of Pine Creek, with a population of 380 (Australian Bureau of Statistics 2014) and Adelaide River (population of 237 (Australian Bureau of Statistics 2014)) support the Northern Territory Property of Newmarket Gold.

The Arnhem Highway to the east-southeast of Darwin provides a communication link to the Kakadu National Park and Jabiru, a town of 1,135, which provides accommodation for the uranium mines in the vicinity. Accommodation and services are available along the highway, primarily for the tourist trade.

5.5 POWER

Power in the Northern Territory is generated and distributed by the Northern Territory Power and Water Authority (NTPWA). The NTPWA’s main gas turbine power station is located at Channel Island in Darwin, which is capable of producing 254 megawatts (MW). A 19.5 MW power station exists at Pine Creek and is interconnected to the 132 kilovolts (kV) line from Darwin to Katherine. A 66kV line connects the Union Reefs processing facility, Brocks Creek, Cosmo Howley and the Cosmo Village to the Pine Creek Township.

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Gas is supplied to the area via the Amadeus Basin to Darwin pipeline. Spurs off this pipeline service Katherine, Pine Creek and the Cosmo site. The Bonaparte gas pipeline also runs through the area, connecting with the Amadeus pipeline near the Fountain Head/Tally Ho deposits area.

5.6 WATER

The Union Reefs processing plant sources its water from two main storage dams off the McKinley River with total storage capacity of some 1,970 million liters. In addition, the Union Reefs deposit has on-site water storage capabilities within the former Crosscourse open pit, which is currently partially flooded and is used for tailing storage. Pit dimensions are 1,300m by 600m by 240m deep. The processing plant does re-cycle most of its water from the Crosscourse pit via a return water system.

The Property area receives approximately 1,500 mm in rainfall each year.

The Cosmo accommodation camp has capacity for 280 people and has its own potable water bore field and treatment plant, which softens and chlorinates the water supplied to the camp.

5.7 COMMUNICATIONS

The project areas have landline telephone communications (Telstra) as well as satellite and microwave communication systems. Mobile telephone coverage under the Telstra Next-G network exists throughout a large area of the mining and plant sites and standard VHF radio communications are used for operational purposes. Work was completed in 2014 to upgrade the mobile coverage in the Cosmo Village, which has shown a great increase in the signal strength.

5.8 MINING PERSONNEL

The Property is located within an area that has a strong mining tradition and, as a result, the mining industry within the region is well understood and supported by the surrounding centers. Mining activities have a direct impact on the manufacturing, service and hospitality sectors of the local economies immediately surrounding the Property area, with past mining operations at Tom’s Gully, Mount Todd, Maud Creek, Cosmo Howley, Brocks Creek, Enterprise and Union Reefs gold mines previously employing significant numbers of local residents.

The surrounding provincial centers of Pine Creek and Adelaide River, which occur within an 80km radius of the projects, provide good support facilities including housing, law enforcement, basic medical and community facilities.

Further afield, the Darwin central business district is the administrative center for the Northern Territory and lays one to three hours’ drive from the project tenements. Mining, engineering and consulting firms as well as commercial assay laboratories are based in Darwin and other population centers in the Northern Territory such as Pine Creek.

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Newmarket Gold is involved with safety, management, mining engineering, geology and exploration, survey, stores control, processing and maintenance, environmental and permitting, and administration functions. Contractor personnel are involved primarily in the mining and haulage functions.

Underground mining at Cosmo is conducted by Downer. This contract was placed out for tender and won by Downer during the Q4, 2013. Downer commenced operation at Cosmo in Q1, 2014 and opted to continue the contract for an additional 12 months in 2016.

Haulage of mineralization from the mining centers to the Union Reefs mill is completed by Fawcett’s; they have supplied this service since the start of 2015, taking over from Ostojic’s.

A number of other contracting groups were engaged for maintenance services, labor hire services, road repairs, drilling activities and other typical contracted activities. These service providers are covered in more detail in Section 18 of this technical report.

The Union Reefs Operations Center (i.e. processing facility) has a workforce of approximately 60 people including 45 employees and 15 contractors. Newmarket Gold recruits people by offering residential accommodation opportunities and good camp facilities.

Mining operations run 24 hours a day, each day of the year primarily based on two 12 hour shifts working a range of rosters. Milling operations at the Union Reefs mill currently run on a 9 days on, 5 days off roster with two 12 hour shifts. During the off days, maintenance is conducted on the plant.

5.9 ACCOMMODATION

The majority of Newmarket Gold personnel live in the local towns or the Cosmo Village. The Pine Creek accommodation village was placed on care and maintenance in December 2013.

The 120-person Pine Creek camp, located in the town of Pine Creek, was used to house the processing and administration personnel, however, since being placed on care and maintenance all these personnel have been housed at the Cosmo Village.

The 280-person Cosmo Village, located adjacent to the Cosmo-Howley mining areas, is used to house all personnel. The camp is an excellent facility with full kitchens, recreation facilities and single ensuite rooms. The camp provides two hot meals per day, pack lunches, hospitality, and laundry service and also has entertainment facilities such as tennis courts, swimming pool and pool tables.

Rented housing accommodation is also available for site personnel in the local communities of Adelaide River and Pine Creek, if required.

5.10 PROCESSING FACILITIES

All mineral processing is conducted at the Union Reefs site. More details of the processing facility can be found in Section 17 of this technical report.

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6 HISTORY

Darwin (previously named Palmerston) was initially settled in 1869. In the following year, a decision was made to build a telegraph line overland between Port Augusta in the state of South Australia and Port Darwin to link Australia with the rest of the world. It was by chance that a significant discovery was made when hole diggers found gold in the Pine Creek area (Jones 1987).

The discovery of gold in the Northern Territory was followed up by exploration hoping that the area would establish itself on gold mining. Senior Surveyor George McLachlan was appointed on April 29, 1871 to the new post of Warden of Pine Creek Gold Find. McLachlan and his members organized a number of wells along the course of Pine Creek in an effort to discover the extent of the gold find. The results were promising as traces of gold were found in every case.

The gold rush began shortly after the discovery. A group of people from Adelaide migrated up to join the mining boom. In 1873, the suggestion of employing Chinese ‘coolie’ labor was raised and Douglas Bloomfield took the responsibility of recruiting laborers from Singapore. He secured 176 Chinese and 10 Malays who arrived at Port Darwin in 1874. There were also independent Chinese gold seekers who took part in the mining activities. The first group of Chinese gold seekers reached Port Darwin in October 1877.

Since the first discovery of gold in 1870, the Northern Territory has produced approximately 14.9Moz of gold. Of this total, an estimated 3.7 million ounces have been produced from the Pine Creek Orogen (Ahmad, Wygralak and Ferenczi 2009). There are about eight hundred documented gold occurrences of potential economic significance and a mineral resources inventory of a further 17 million ounces gold.

LOCATION ESTIMATED PRODUCTION
Union Reefs 1.0moz
Pine Creek 750,000oz
Cosmo, Howley / Woolwonga 650,000oz
Brocks Creek 230,000oz
Tom’s Gully 200,000oz
Rustler’s Roost 113,000oz
Alluvial Deposits ~1.0moz

TABLE 6-1 HISTORICAL GOLD PRODUCTION – PINE CREEK OROGEN

There have been three significant periods of exploration and gold production. Early mining (1870-1915) was selective and concentrated on high-grade (several 10’s of g/t Au) veins, mainly in the Pine Creek Orogen. During this period, mining was from highly selective shallow pits, shafts and narrow adits that systematically followed the auriferous lodes. These old mines generally were confined to the oxide zone and stopped at the water table.

The next significant phase commenced with the discovery of medium-high grade (15-20g/t Au) ironstone hosted deposits in the Tennant Creek inlier in 1936, with production peaking during 1971-75.

The current phase of gold exploration and production commenced in 1987 and concentrated on bulk open cut mining of relatively low-grade (2-3g/t Au) mineralization.

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6.1 COSMO MINE AND SURROUNDING AREAS

Gold was discovered at Brocks Creek in 1871 and at Cosmo Howley in 1873 during the construction of the Adelaide – Darwin overland telegraph line. This was the prelude to a long period of alluvial working by Chinese miners and lode mining by English companies and Chinese tribute miners until 1914.

Balfour (Balfour 1981) estimates that approximately 1,028kg of gold was produced from the Brocks Creek line of workings from discovery to 1914, with some 823kg produced from high grade gold reefs at Zapopan; the name Brocks Creek and Zapopan seem to have been used interchangeably in the early years. Blanchard and Hall (1937) (Blanchard 1937) estimate that about 71,000t of mineralization was mined along the Howley Line during this period, producing about 24,000oz of gold at a recovered grade of about 10g/t Au.

The field subsequently declined in the mid-1910s and there was little activity in the area for the next 60 years apart from limited small scale mining and minor exploration. The Brocks Creek and Howley areas were explored for gold and base metals during the 1970’s and 1980’s by numerous mining and exploration companies, often in joint venture. Most of these companies carried out extensive drill testing of various costean (trench) intercepts, IP/resistivity geophysical surveys and soil geochemical anomalies with mixed results.

In 1975, Dampier Mining acquired an exploration license over the Howley line, and under a joint venture with Dampier Mining, Homestake Gold Mining conducted the first major assessment in 1977 to 1978.

Several small alluvial shows in the Chinese Howley area were mined between 1986 and 1990. Mining operations commenced at Cosmo Howley in 1987 with gold-bearing mineralization mined from the Cosmo Howley, Phantom, Chinese South, Chinese Howley and Big Howley pits. All mining ceased prior to April 1995.

Deposit Years Tonnes (Mt) Grade (g/t Au) Ounces (Moz)
Cosmo, Howley 1987 – 1993 6.9 2.1 0.47
Woolwonga 1991 – 1995 2.9 2.5 0.23
Howley-Big Howley 1992 – 1995 1.1 1.8 0.06
Total   10.9 2.2 0.76

TABLE 6-2 ESTIMATED HISTORICAL GOLD MINED. COSMO HOWLEY GOLD PROJECT

During 2002, exploration activities focused on compilation and validation of data in relation to the Cosmo underground mineral resource and the re-evaluation and reassessment of the near surface mineral resources at Howley, Western Arm and Yam Creek. Acquisition of detailed airborne magnetic geophysical data, Landsat and SPOT remote sensing imagery and GIS topographic data provided the basis for a structural interpretation and targeting definition over the Brocks Creek/Zapopan, Yam Creek and Woolwonga areas. Infill and extensional mineral resource definition drilling at Brocks Creek/Zapopan, the commencement of the Brocks Creek/Zapopan underground decline, mineral resource modeling of the Cosmo mineralization and initial Scoping Studies on the Cosmo underground mineral resource were all undertaken.

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In March 2003, pre-production activities within the Burnside Gold Project were postponed, pending an improvement in the gold price and the decline into the Brocks Creek/Zapopan underground deposit was suspended at a vertical depth of approximately 125m below surface.

In 2004, infill drilling was undertaken to define high-grade mineralized zones within the Cosmo Deposit and a Scoping Study into the potential for Cosmo to be mined as an underground gold operation was initiated.

In early 2005, further mineral resource definition drilling of eight RC/diamond drillholes was completed at the Cosmo Deposit with the aim of infilling the upper levels of the deposit to bring the mineral resource to a mineral reserve category. Samples from the Cosmo Deposit were sent for metallurgical test work at Amtec Laboratories in Perth, Western Australia. This test work demonstrated that the Cosmo mineralization is amenable to good recoveries through the Union Reefs processing facility.

From late-2005, past operators carried out exploration and mineral resource definition drilling programs, which have led to the updating of several mineral resource models and optimized mine designs. These programs were mainly focused on prospects within the western and southern portions of the Burnside deposits, principally along the Howley and Brocks Creek-Zapopan Anticlines.

During 2009 to 2012 Crocodile Gold mined from the open pits at Howley, Mottrams, Princess Louise, North Point, Rising Tide deposits as well as obtaining ore from the Brocks Creek and Cosmo underground mines. A total of 6.98Mt of mineralization was mined since production re-commenced in November 2009 until the end of December 2015. The overall reconciliation with the mill has performed well with grade control estimating a grade of around 2.07g/t Au and the mill averaging around 1.97g/t Au (see Table 6-4).

During the past three years some additional material has been milled from historic stockpiles located around the Pine Creek region. A low-grade stockpile of around 80,000t was transported from Moline Mine area to the Union Reefs mill for processing in 2011 and 2012. It is estimated that the grade was around 0.8g/t Au. Also in the period around 23,000t at 1.0g/t Au was hauled from the Glencoe Deposit. This is included in the Milling Figures but is not included in the Mining Figures. A third stockpile from the Golden Dyke pit, which was rehabilitated during 2014, with around 58,000t of material at an average grade of around 0.92g/t Au was also processed at the Union Reefs processing facility.

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Period                                                                                              Company
1975-1977 Dampier Mining
1977-1982 Homestake Gold Mining
1982-1984 Golden Dyke Joint Venture (Geopeko-Anaconda)
1984-1987 Regent- Southern Goldfields JV
1987-1995 Dominion Mining
1995-2003 Territory Goldfields NL
2003-2005 Burnside Joint Venture (Buffalo Creek Mines Pty Ltd and Territory Goldfields)
2005-2008 GBS Gold Australia
2008-2009 Receivership of GBS Gold
2009-2015 Crocodile Gold Australia/Newmarket Gold NT Operations

TABLE 6-3 SUMMARY OF HISTORIC OWNERSHIP OF COSMO HOWLEY MINING AREA

  Milled Mined
Year Tonnes Grade
g/t Au
Ounces Tonnes Grade
g/t Au
Ounces
2009 29,000 1.72 1,600 199,000 1.56 10,000
2010 1,855,000 1.55 92,300 1,967,000 1.62 102,200
2011 1,886,000 1.21 73,100 1,854,000 1.22 72,700
2012 916,000 1.38 40,700 705,000 1.72 38,900
2013 720,000 3.55 82,200 741,000 3.63 86,500
2014 809,000 3.30 85,900 791,000 3.30 83,900
2015 725,000 2.99 63,300 727,000 3.00 70,100
Total 6,940,000 1.97 439,100 6,984,000 2.07 464,300

TABLE 6-4 RECONCILIATION FIGURES FOR CROCODILE GOLD/NEWMARKET GOLD MILLING - 2009- 2015

6.1.1      RISING TIDE

From 1996 to 1998, Acacia Resources undertook several RC and diamond drilling programs to complete definition of the Rising Tide Deposit, with 506 drill holes for a total of 26,637m.

In 2006, a total of 21 hole RC and two hole diamond drill holes for a total of 994m were completed by GBS Gold to define known geological structures and extend existing mineralization.

A subsequent 2006 mineral resource calculation defined an Indicated mineral resource above a cut-off of 0.7g/t Au of 2.06Mt grading 1.62g/t Au for a total of 107,200oz Au. At the same lower cut-off grade, the Inferred mineral resource was 0.92Mt at 1.33g/t Au.(1)

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves. In additional a mineral resource estimate that is more current is included in Section 14, and should be used for all reviews of the Rising Tide deposit.

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In 2006 and 2007 GBS Gold mined by open pit from the Rising Tide Deposit 533,930t grading 1.47g/t Au for 25,282oz.

6.1.2      BON'S RUSH

At Bon’s Rush Western Mining Corporation (W.M.C.) as part of their Mount Ringwood Joint Venture previously explored the area of the deposit in the late 1980’s. WMC undertook extensive regional exploration including an aeromagnetic survey, ground magnetic survey, rock chip sampling, stream sediment sampling, soil sampling, costeaning, RAB drilling, RC drilling, and diamond drilling programs. They identified numerous areas of anomalous gold mineralization, including the Bon’s Rush area.

In the late 1990’s to early 2000’s, Northern Gold explored the Bon’s Rush area as part of their Mt. Paqualin Project, completing regional BLEG soil surveys, detailed regional geological mapping and a detailed aeromagnetic survey. Prospective targets generated from this initial work were followed up with infill LLFA soil sampling and/or RAB drilling, which resulted in the discovery of the Bon’s Rush deposit.

Northern Gold completed at least 234 RAB holes (4,102.5m) at Bon’s Rush, which along with the soil sampling results identified five main target areas. Detailed 1:5000 geological mapping and a structural interpretation of the area accompanied this RAB drilling. Follow up of limited RC drilling (45 drill holes (3,372.1m) was completed on four of these targets and two diamond drill holes (150.3m) were completed on the Bon’s Rush Main Zone.

Western Mining Corporation as part of their regional exploration completed three RC holes over the Bon’s Rush South target (FSDC 57, 58 and 59).

6.1.3      GOODALL

The Goodall Mine, located 30km to the east of the town of Adelaide River, was discovered in 1981 as a result of a helicopter supported rock chip sampling program. This was followed up with soil sampling and trenching and in 1982 with diamond drilling.

In 1985 Western Mining became involved with the program and was known as the Mt Ringwood Joint Venture. By 1986 a total of 44 diamond drill holes, 135 RC holes, two water wells and six water monitoring holes had been established along with over 7,500m of trenching. Gold was first poured in 1988 and by September 1993 had produced a total of 4.095Mt grading 1.99g/t Au for a total of 228,400oz.

6.2 UNION REEFS AREA

Gold was discovered by prospectors at Union Reefs in December 1873 and since then approximately 1,600 pits, shafts, adits and open cuts have been worked to a depth of approximately 60m (Shields et al, 1967). Chinese miners held most of the claims until 1894. Much of the gold recovered during this time was not recorded but data for the period 1884 to 1910 show production of 48,000oz of gold from 58,000t of mineralization for a recovered grade of 26g/t Au (Hossfeld 1936).

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Very little work other than small-scale mining and exploration was conducted over the next 50 years.

Drilling during the 1960’s, by the Bureau of mineral resources, identified a gold mineral resource at the Crosscourse deposit. This was followed by further drilling in 1969 and 1970 at the Crosscourse and Lady Alice North deposits. Between 1984 and 1988, 25 exploration holes were drilled at Ping Que and Crosscourse deposits, which gave encouraging results and led to a historical estimate. In 1988, 68 percussion holes were drilled along the northern half of the Union and Lady Alice lines of mineralization. No further work was carried out until 1991 when detailed soil sampling, geophysical surveying and percussion and diamond drilling was carried out within the Union Reefs leases. By 1994 a mineral resource had been defined at Crosscourse and Union North deposits that led to the completion of a detailed feasibility study, open cut mining and the construction of a 1.7Mtpa CIL processing plant.

In 1997, Union Reefs was expanded to process harder fresh rock and increase throughput to 2.5Mtpa and gold production from 87,000 to 120,000oz per annum. The expansion included adding two leach tanks and a tertiary crusher. Mine production was primarily from the Crosscourse, Union North, Prospect Claim, Lady Alice and Ping Que deposits.

In 1998, exploration drilling to the north of the Crosscourse open pit identified shallow mineralization over a 1.2km strike length. As drilling continued in this area, several shallow and continuous zones of mineralization varying from 5 to 15m in width were defined over a 700m strike length at Alta and Orinoco. A new mineralized zone, Dam A, was identified 400m northeast of the Union North Deposit and extended for more than 300m. In late 1998 and early 1999, head grades declined as mining progressed in lower grade areas on the eastern margins of the Crosscourse pit and at Union North deposit.

By 2003, mining at the project was moving towards its final stages and was directed towards small, dispersed remnant mineral resources in proximity to the plant. Total gold production from the commencement of operations at Union Reefs in 1994 to June, 2003 was estimated by AngloGold to be 879,824oz.

GBS Gold’s Union Reefs Operations Centre was officially opened on November 10, 2006. Gold production commenced in late 2006 with mineralization being sourced from historical stockpiles, the Brocks Creek/Zapopan underground mine and the Rising Tide and Fountain Head open pit mines. The Fountain Head mine has extracted mineralization from both the Fountain Head and the Tally Ho lodes. Total production from the Union Reefs operations until December 31, 2007 was 1,084,000t at an average head grade of 2.43g/t Au producing 80,092oz Au. In 2008, total production was 1,660,496t at an average grade of 1.64g/t Au to produce 87,538oz gold.

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Project Years Tonnes
(Mt)
Grade
(g/t Au)
Ounces
(Moz)
Burnside (Zapopan) Gold Project 1987-2000 16.5 2.0 1.05
Union Reefs Gold Project 1884-1910 0.06 26.0 0.05
1994-2003     0.88
2006-2007* 1.08 2.43 0.08
Pine Creek Gold Project 1894-1915 0.01 18.39 0.07
1984-1993 12.3 2.37 0.77
Maud Creek Gold Project 1988-1991 Tailings ? 0.04
2000 0.18 3.5 0.02
Moline Gold Project 1954-1957 0.03 31 0.04
1987-1991 1.60 2.1 0.1
Total       3.10

(*)Includes production from outside the Union Reefs deposit area

TABLE 6-5 ESTIMATED HISTORIC GOLD PRODUCTION PINE CREEK REGION 1985- 2007

6.2.1      ESMERALDA

In 1990-1991 Zones “A” and “B” were defined by Cyprus within EL6880 by a soil geochemical survey. Gold and arsenic were determined. Cyprus was earning equity from registered owners, Astron-Solpac, within the Esmeralda Joint Venture.

In 1991-1992 Cyprus Gold drilled 25 RC drill holes into the prospect (ERC0001-ERC0025). The holes were allocated to Zone “A” (ERC0001-ERC0010) and to Zone “B” (ERC0011-16). This drilling program was completed in two phases: a 16 hole/1,110m phase followed by a nine hole/740m program. The initial phase was targeted on soil and rock anomalies, the second phase providing selective down dip testing of phase one intersections. Phase two drilling was allocated to Zone “B” (ERC0017-ERC0019) and to Zone “A” (ERC0020-ERC0025). The best result from Zone “A” was 12m @ 3.03g/t Au from 22m in ERC0002. The best result from Zone “B” was 13m @ 2.33g/t Au from 37m in ERC0023.

In 1994 Billiton Australia reviewed the Cyprus data and drilled 15RC holes (EAP0001-0015) into Zone “A” for a total of 938m and a diamond tail of 21m on EAP0015 (renamed EAD0015).

In 1995 Acacia drilled 40 RC holes (ERC0041-0080) into Zone “A” and “B”, for a total of 2,573m. In August 1995, a manual mineral resource calculation was completed with the available data. Bulk densities of 2.52g/cm 3, weathered, and 2.74g/cm 3, fresh were used. This uncut geological mineral resource estimates using a 0.7g/t Au lower cut-off gave a combined inferred mineral resource of 879,000 tonnes @ 2.0g/t Au.

In 1996 Acacia completed 27 RC holes for 1,794.5m and 4 diamond drill holes for 155.5m. Twenty three of the holes were drilled on Zone “A”.

In 1997 50 RC holes and one re-entry were completed for 4,495m. All holes were surveyed with Eastman single shot. At Zone “A” the deposit was tested to 100m vertical depth. A new lens 100m west of Zone “A” was discovered on four sections. Further drilling to extend the southern limits was unsuccessful. Also during the year a structural analysis of the deposits was commissioned, with a further eight costeans dug for 514m. An airborne radiometric/magnetic survey was completed by UTS. (50m line spacing, 60 degree orientation, 20m terrain clearance, 127km2 total area.) Aerial photography and digital terrain modeling were undertaken.

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A mineral resource estimate was completed using all data. M&RT Consultancy defined an inferred mineral resource of 1.26Mt @ 1.62g/t AuNB1.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves. In additional a mineral resource estimate that is more current is included in Section 14, and should be used for all reviews of the Esmeralda deposit.

6.2.2      PROSPECT DEPOSIT

In 2005 Bill Makar (B. Makar 2005b), Chief Mine Geologist for AngloGold at Union Reefs wrote a report on the Prospect deposit. A summation of his comments and findings follows:

 

Prospect deposit was originally mined in the late 1800’s, mainly by underground means but also a few shallow pits and gouges. It was recorded as one of the richest prospects in the Union Reefs mining district.

 

In the mid- 1990’s. Prospect was mined as an open pit by Acacia /AngloGold in two stages.

 

Initial stage was in 1997. The mining was strongly influenced by the old working, being largely a remnant mining exercise of the remaining low-grade mineralization and remaining pillars. The old workings continued below the 1207.5RL and were found to extent below the 1170RL final pit floor along the main ramp access.

 

The second phase was the mining of the pit from the 1207.5RL to the final depth of 1150RL.

 

Total mined by Acacia/ AngloGold was 443,886t @ 1.55g/t for 22,090oz gold.

 

Gold recoveries were in excess of 93% with nearly 50% recovered by gravity means.

 

High-grade gold values are associated with intense quartz veining and stockworks within a near vertical shear zone. These quartz veins/stockworks pinch and swell from <1m up to 5-10m both along strike and down dip. The lode appears to plunge to the north.

 

A preliminary underground mineral resource evaluation was carried out on the Main Lode to assess if it was a viable underground target. The Main Lode is the down dip extension of the main zone mined by the open pit. The maximum depth that the pit was mined down to was the 1151RL.

 

A preliminary mineral resource of 103,000t @ 8.04g/t Au (1) (applying a 45g/t top-cut) was identified. The mineral resource extent is between 7160N to 7600N and from 1150RL to 1060RL. The mineral resource is relatively well drilled from below the mined pit to the 1120RL and poorly drilled to the 1050RL. It was open down dip.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves. In additional a mineral resource estimate that is more current is included in Section 14, and should be used for all reviews of the Prospect deposit.

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Makar subsequently calculated a mineral resource for the Prospect Deposit. In summary:

cut-off g/t Au Surface to 1150RL 1150RL to 1060RL
Tonnes g/t
Au
Ounces Tonnes g/t
Au
Ounces
30 23,342 5.85 4,390 103,903 7.2 24,052
45 23,342 6.29 4,720 103,903 8.04 26,858
60 23,342 6.62 4,968 103,903 8.58 28,662
Un-cut 23,342 7.26 5,448 103,903 10.12 33,806

Notes:
Material above 1150RL mainly transitional and oxide
Material below 1150RL is all fresh, SG of 2.7g/cm 3
Average Oxide surface ~1170RL; SG 2.5g/cm 3
Transitional horizon between 1170RL to 1155RL; SG 2.6g/cm 3
Deepest Prospect pit was mined was ~1151RL (bottom of Good-bye cut in the north pit)

TABLE 6-6 HISTORIC GRADE COMPARISON OF PROSPECT DEPOSIT MAIN LODE AT VARIOUS AU CUT- OFF GRADES (NB1)

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves. In additional a mineral resource estimate that is more current is included in Section 14, and should be used for all reviews of the Prospect deposit.

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Makar’s conclusions and recommendations included:

6.3 PINE CREEK GOLD PROJECT

6.3.1      RENISON GOLDFIELDS

1980-1984 - Renison Goldfields Consolidated (RGC) and Enterprise Gold Mines (formerly Jingellic Minerals NT P/L) had a joint venture with the leases. RGC managed exploration on their sector of Gandy’s Hill. The original Gandy’s Hill leases were GML163A-166A inclusive. Color aerial photography at 1:100,000 was flown over 752km2, centered on Pine Creek. Geospectrum scaled the topographic maps to local grid at scales of 1:500, 1:1000 and 1:5000.

1986 - 278m of RC drilling was completed on MLN785 at North Gandy’s Deposit.

1987 - Ten diamond core holes were completed over Gandy’s North and a seven hole RC program was completed over MCN157 and MCN969 on South Gandy’s.

1988 - A program of 445 vertical RC holes was drilled on MLN 786, 785, and MCNs 969, 1056 and 157. An oxide mineral resource estimate was carried out.

1989 - Nine diamond core holes were completed for 845m as well as 5,000m of RC drilling. On MCN1058 (Nth International) and MCN1230 (south of Gandy’s Hill) a program of RC drilling was completed for 1,169m.

1990 - Five costeans were dug over North Gandy’s and North International.

1991 - Pine Creek Goldfields obtained the title of eleven leases northeast of Gandy’s Hill Deposit rom Australian Energy. Eight costeans were dug on the northern extensions of the Enterprise Anticline. The northern part of Arimco's MLN39 was swapped for a part of PCG's MCN1058 for the re-location of the Stuart Highway.

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1992 - Leases over Gandy’s Hill and International deposits were handed to Pine Creek Goldfields. The negotiation and final acquisitions occurred.

1993 - After the acquisition, the grid was extended and RC drilling was undertaken into infill the original work to 25m line spacing. Holes were drilled to 30m on a 50m by 20m pattern. A waste dump sterilization was carried out north of the International Deposit comprising 21 holes vertical RC holes. In mid-1993, the Gandy’s Hill North, Gandy’s Hill South and International deposits were mined by open pit methods. Mining at North Gandy’s was completed in November.

1994 - Mining at Gandy’s Hill South and International deposits was completed.

1995 - A program of RC drilling comprising 14 holes for 420m was completed at the Gully Project between the original Gandy’s and International ridges. No significant gold values were met with in what was interpreted to be a synclinal structure. The mined areas were rehabilitated, the Enterprise treatment plant was sold off and the area became inactive.

6.3.2      CYPRUS-AMOCO-ARIMCO

1983 – The Gandy’s Hill property was optioned by Tasbax P/L to Amoco Minerals. The company carried out surface rock-chip, dump sampling and mapping followed by 15 RC holes for 1,163m. The known association between the quartz sulphide bodies and gold mineralization was tested by an induced polarization survey.

1984-86 - Amoco dealt their tenement to two companies. Amoco executed a farm out agreement with Lightning Ridge Mining and later joint ventured 80% of its interest to TERREX Resources NL. Terrex carried out further exploration. Mapping and drilling, which included both diamond and RC were undertaken. Some Terrex data was 'lost' during the time when Tasbax put the managers in default and Cyprus (formerly Amoco) renegotiated the agreement.

1987 - Cyprus and Hudspeth & Co. entered into an option with Frith, an adjacent lease owner, covering MLN790 over the International deposit. At this time Tasbax also added the house and building on site to the agreement.

1988 - The house and all the leases including late MLN39 in the "Carlton Project area" (also known as Gandy’s Hill) were purchased by Cyprus from Tasbax.

1989 - Cyprus Gold Australia transferred their interest in the project to JV partner Arimco NL.

1990 - Arimco undertook an evaluation of the Carlton Project. The extra data received from Pine Creek Goldfields was included into the database. It was concluded that the Gandy’s Hill deposits amounted to 759,257 @ 2.41g/t Au (inferred global mineral resources) and the International Deposit amounted to 1,607,821t @ 2.59g/t Au (inferred global mineral resources) NB(1).

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves. In additional a mineral resource estimate that is more current is included in Section 14, and should be used for all reviews of the Gandy’s prospects.

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1992 - Arimco's entire "Carlton Project" leases were transferred to Pine Creek Goldfields. Pine Creek Goldfields NL also held leases that surrounded the Arimco Group and simultaneously did exploration activity on their sectors.

6.3.2      ENTERPRISE MINE

The Enterprise shaft preceded the establishment of the Enterprise pit, which was sunk to 79m in 1915 following a government sponsored diamond drilling program. Multiple drives led off the shaft targeting mineralization. Operations at the mine closed down with the outbreak of WWI and R. and M. Blake did no work until the 1960s when the shaft was worked intermittently for a period of about 20 years.

In late 1980 Goldfields Exploration Limited commenced a program of rock chip, road cut and underground channel sampling as well as surface geological mapping over all the areas of old workings. An RC and diamond drilling programs commenced in May 1981 and by 1984 a total of 13,232m had been completed in the Enterprise Mine area. Further geological mapping was carried out in 1983 and 1994.

A local mine grid was established that paralleled the axis of the Enterprise Anticline. The bearing of the grid was 41deg 29min 20sec west of true north. This grid has been applied to the whole of the Pine Creek Goldfields. Easting 11200E passed along the axis of the fold and 11000N was just beyond the southern crest of the Enterprise pit.

Mining of the Enterprise pit commenced in September 1985 following a drill-out on sections 50m apart and holes on section 20m apart, and continued until closure in August 1993. A mineral reserve of 6.7Mt @ 3.33g/t Au was used prior to mining (0.7g/t Au oxide lower cutoff, 1.0g/t Au primary lower cutoff).

Following the confirmation of geostatistically predicted mineralization by diamond drilling in 1986 mineable mineral reserves at January 1987 were 9.2Mt @ 2.7g/t. AuNB1. The combined Enterprise-Czarina pits ended up producing 9Mt of mineralization and this was treated for the recovery of 600,000oz at an average head grade of 2.59g/t Au. Average mill recovery was 79%.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves. In additional a mineral resource estimate that is more current is included in Section 14, and should be used for all reviews of the Enterprise prospects.

It was observed early in mining the Enterprise pit that blast hole and RC drilling gave more accurate assay results than diamond drilling by a factor of +1.35.

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The down plunge extension of the Enterprise mineralization below the pit and beyond the southern mine limits was tested by drilling. There had been some suggestions of a possible steepening of plunge at the south end to around 45 degrees, or of cross faulting moving the axis up or down.

Diamond drilling programs were completed that had the objective of targeting the 40m to 60m wide mineralized zone that dipped west at 75 to 80 degrees within the southern limits of the pit.

High grades were known to be associated with linear/planar structures 100-150m in strike extent and relatively continuous down dip and also the intersection of planar structures that form pods of variable orientations having average dimensions of 15m x 15m.

Half a dozen deep holes were targeted beneath the pit, however, these only met with narrow modest gold grades. There is up to 100m of untested zone beneath the pit floor to the RL of the deep holes. Testing of this zone would require large step backs and shallow dips.

6.3.4      SOUTH ENTERPRISE DEPOSIT

The area just to the south of the Enterprise Pit has been an area of exploration interest since the 1980s. Significant amounts of drilling have been conducted and a mineral resource has been outlined at South Enterprise.

Makar and Muller (Makar and Muller 2006) has identified nine different loads in the South Enterprise prospect ranging in size from 17,000t to over 100,000t of mineralization grade material.

Drilling and rock chip sampling has been done in the area along strike to the south of the Enterprise pit. The old Stewart Highway cuts diagonally across the eastern side of the tenement as a road cutting. This cutting was excavated in 1988-1989 and mineralization grade material was extracted from it.

The Burnside Joint Venture conducted drilling at the South Enterprise Prospect in two phases; one phase of diamond and RC drilling in the early 1990’s and a second phase of RC drilling in 2004-2005. Ninety five diamond holes for a total of 12,048m were drilled prior to 1994; this was done alongside 28 RC holes for a total of 1,846m. The 2004 drilling consisted of five RC holes and a total of 652m. Another 30 RC holes, with a total of 2,099m, were drilled in 2005.

6.3.5      INTERNATIONAL & GANDY'S HILL DEPOSITS

Tasbax P/L optioned the Gandy’s Hill property to Amoco Minerals in October 1983. Amoco carried out surface rock-chip, dump sampling, and mapping followed by 15 RC holes for 1,163m. An induced polarization survey was completed to test the known association between the quartz sulphide bodies and gold mineralization. Amoco concluded a farm-out agreement with Lightning Ridge Mining NL and subsequently joint ventured 80% of its interest to Terrex Resources NL. Terrex carried out further mapping and drilling which included both diamond and RC work. By September 1986 Tasbax put the managers in default and Cyprus (formerly Amoco) renegotiated the agreement. Some Terrex data was ‘lost’ during this change of interests. During March 1987, Frith, an adjacent lease owner, entered into an option with Cyprus and Hudspeth & Co. covering MLN790 over the International Deposit. At this time Tasbax also added the house and buildings on site to the agreement. By July 1988 Cyprus had purchased from Tasbax the house and all the leases including MLN39 in the “Carlton” project area.

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Between 1987 and 1989 a number of RC drilling programs were completed on Gandy’s Hill and International deposits along with geological mapping. An agreement for the transfer of exploration data with PCG was completed in November 1989. By this time, a total of 8,932.1m of drilling had been completed at Gandy’s Hill and 8,581m on International Deposit.

On December 29, 1989, Cyprus Gold Australia transferred its interests in the Carlton Project to joint venture partner Arimco NL, for other considerations as part of an Australia-wide redistribution of assets on dissolution of the JV.

In January and February 1990 Arimco undertook an evaluation of the Carlton Project. The extra data from Pine Creek Goldfields was incorporated into the database. Geological modeling and geostatistical work was carried out by Guibal (Guibal 1990). The work was completed in June 1990 and inferred global mineral resources were reported at the Gandy’s Hill and the International line.

Negotiation and final acquisition by PCG of all Arimco leases over Gandy’s Hill and International deposits occurred in July-August 1992. The Gandy’s Hill North, Gandy’s Hill South and International deposits were mined by open pit methods commencing in mid-1993. North Gandy’s was completed in November 1993 and the others were completed in 1994.

Mining ceased at Pine Creek on November 2, 1994 when the last truckload of ore was hauled from the South Gandy’s pit. Milling of approximately 1Mt of low-grade stockpiled ore continued for a further eight months and the mine was officially closed in mid-1995.

All pits on MLN1130, except South Gandy’s, were backfilled and rehabilitated during operations. At this stage it is difficult to detect the location of the backfilled pits because of the advanced rehabilitation.

6.3.6      CZARINA AND SOUTH CZARINA DEPOSITS

Czarina deposit is a zone of mineralization that was mined along with the Enterprise deposit in the early 1990’s by Pine Creek Gold Fields Ltd. The pit was back filled and rehabilitated when mining was finished. A mineral resource potentially remains under the old pit floor at Czarina.

Pine Creek Goldfields operated the Czarina pit as a satellite pit to the Enterprise pit between January 1992 and September 1993. A total of 738,047t of ore was extracted at this time at an average grade of 1.67g/t Au. This can be divided into 593,617t of oxide material at 1.61g/t Au and 144,430t of primary material at 1.92g/t Au. The pit was roughly 175m by 685m at the time of its completion.

A total of 169 drill holes have been drilled at Czarina, 153 diamond and reverse circulation drill holes before 1994 by Pine Creek Goldfields and 16 reverse circulation drill holes in 2004 by previous owners.

6.3.7      MONARCH DEPOSIT

The Monarch pit sits between the Enterprise and Gandy’s pits. This small pit provided low-grade mineralization for the Pine Creek mill when the other larger open cut operations were taking place. It was a low-grade, small operation. The Monarch open pit has also been backfilled and is also difficult to locate due to rehabilitation.

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6.3.8      COX DEPOSIT

The Cox deposit is located to the north of the Bashi Bazook deposit. It is located in rugged terrain covered in numerous old workings. As with Bashi Bazook the orientation of the mineralized zone is oblique to the anticline structures that dominate the Pine Creek area.

Diamond drilling programs were completed that had the objective of targeting the 40m to 60m wide mineralized zone that dipped west at 75 to 80 degrees within the southern limits of the pit.

High gold grades were known to be associated with linear/planar structures 100-150m in strike extent and relatively continuous down dip. The intersections of planar structures form pods of variable orientations with average dimensions of 15m x 15m.

Half a dozen deep holes were targeted beneath the pit, however, these only met with narrow modest gold grades. There is up to 100m of untested zone beneath the pit floor to the RL of the deep holes. Testing of this zone would require large step backs and shallow dips. Mineralization is hosted in the Cox’s Shear, which strikes at roughly 335°, sub-parallel to Bashi Bazook. The shear dips between 65° and 80° to the west. Marjoribanks concluded that the Chinaman’s deposit is hosted in the same shear structure at its southern extent. The Cox’s Shear cuts across the Czarina and Kohinoor anticlines and this is where the highest gold grades have been observed.

Exploration work done on the Cox deposit was all conducted prior to 1994 and the end of mining at Pine Creek. Costeaning of the poorly exposed mineralized zone at Cox gave interesting results including 13.5m @ 7.32g/t Au. Follow up 25m spaced RC drill program gave inconclusive results possibly due to poor understanding of the controlling structures.

Further drilling is required at Cox to understand the extent of the mineralized zone. Mineralization may extend along strike to Chinaman’s Shear as well as linking to the Kohinoor deposit.

6.3.9      BASHI BAZOOK DEPOSIT

Battery Shear/ Bashi Bazook is a zone of mineralization located between Cox’s Shear and Chinaman’s Shear and is probably on a common structural set. The deposit has been referred to as Battery Shear, Bashi Bazook and sometimes as Battery Shear in the south and Bashi Bazook in the north. The area is covered with numerous shafts and historic workings.

Pine Creek Goldfields drilled 20, mostly vertical, RC holes at Bashi Bazook for a total of 1,637m in 1990-1991. Data from these holes was interpreted by Schofield (Schofield 1991)) and it was determined that the mineralization was low grade, and discontinuous. No work has been conducted since on the deposit.

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6.3.10      KOHINOOR DEPOSIT

The Kohinoor deposit sits directly south of the Cox deposit in the mining lease MCN523. It is a long zone of mineralization sub-parallel to the anticline structures that dominate the Pine Creek Goldfields. The Kohinoor deposit has been divided up into several deposits in the past, the northern portion being known as Henry George, the central portion known as Kohinoor and Jensen’s in the south.

The area is covered in old workings and is within the Pine Creek Heritage Zone, which requires minimal disturbance. A colony of ghost bats exists in the Kohinoor adit, which is protected and they would need to be relocated if mining operations were to commence.

The Kohinoor deposit has been drilled extensively. All drilling was completed prior to 1994 and consisted of reverse circulation, RAB and diamond drilling. The results obtained were irregular but some high-grade gold values were observed. They included; 16.5m at 9.72g/t Au in the saddle reefs on the western limb of the anticline, and 9m at 7.7g/t Au in a crosscutting fault across the anticline. Drill spacing varies from about 20-50 meters in the mineralized zone.

6.3.11      ELEANOR DEPOSIT

The Eleanor deposit is situated on a collection of shallow workings and underground workings in the southeastern leg of MLN13. Roughly 200m true north is the Jensen’s adit, at the Kohinoor deposit and workings in the Eleanor prospect are orientated sub parallel to workings at Kohinoor. It has been recorded that the Eleanor Mine was a relatively high-grade producer in in the Pine Creek Goldfields, though tonnage was relatively low. Jones (Jones 1987)reported Olaf Jensen produced a total of 1,652oz of gold from 783t of mineralization in 1887 from the Eleanor Mine.

Relatively little drilling has been conducted at the Eleanor prospect. Holes have been mostly vertical and failed to intercept the high-grade shoots at depth. Trench 10 has been placed across strike at the Eleanor prospect and rock chip samples taken along its length.

6.3.12      ELSINORE DEPOSIT

The Elsinore deposit sits in MCN523 between 9,100N and 9,250N local grid. This area is generally a lower grade zone of mineralization. However, this old mine was one of the first hard rock workings developed in the Pine Creek Goldfields. The prospect has had extensive historic work done, in the form of both surficial workings and underground workings. Heritage listed sites relating to Chinese mining exist around the prospect and require minimal disturbance.

The Elsinore deposit was extensively worked from the 1880s until the early 20th century by mostly Chinese miners. It then remained dormant until Pine Creek Goldfields conducted drilling in the late 1980s and early 1990s.

A series of vertical percussion (25-45ms deep) holes in lines spaced 50m apart were drilled along with eight angled diamond holes (37-122m deep).

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Costeaning has been done 130m along strike of the deposit and returned interesting results and an average grade of 2.7g/t Au.

6.3.13      SURROUNDING EXPLORATION TITLE EL30419

Large portions of MCN523, 1055, 1054 and a small part of MLN13 cover a portion of EL30419. The main purpose of the exploration lease is to cover the zone along strike to the south of the mining leases. The only mineral known occurrence, outside of the mining leases, is a base metal anomaly, termed Lucknow, in the southeastern portion of the exploration lease. Very little work has been done south of the mining leases. Rock chip and soil sampling produced discouraging results and no drilling has been conducted. Magnetic and gravity images of the lease show a structure running from the center of the lease to the southeast, close to Lucknow, that looks similar to the central zone of the Pine Creek Leases

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7 GEOLOGICAL SETTING AND MINERALIZATION

The Precambrian rocks of the Northern Territory have been subdivided into two principal orogenic provinces: the North Australian Craton; and, the Central Australian Mobile Belt. Orogenic domains within the North Australian Craton include the Pine Creek Orogen, the Tanami region, the Murphy, Tennant and Arnhem Inliers and the Northern Arunta Province. These underwent extensive orogenic movements and regional metamorphism between 1870 and 1830ma (Barramundi Orogeny) followed by variably developed transitional tectonics and igneous activity from 1850ma to 1800ma (Ahmad, Wygralak and Ferenczi 2009).

7.1 REGIONAL GEOLOGY

The Northern Territory Properties, including the Cosmo Deeps Mine and surrounding projects, fall within the Archaean to Paleoproterozoic Pine Creek Orogen (PCO), one of the major mineral provinces of Australia (Figure 7-1, Figure 7-2 and Figure 7-3). The PCO is a deformed and metamorphosed sedimentary basin up to 14km maximum thickness covering an area of approximately 66,000km2 and extending from the Katherine area in the south to Darwin in the north. It hosts significant mineral resources of gold, uranium and platinum group elements (PGEs), as well as substantial base metals, silver, iron and tin-tantalum mineralization.

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The PCO comprises a series of late Archaean granite-gneiss basement domes, which are overlain by a fluvial to marine sedimentary sequence. Several highly reactive rock units are included within this sedimentary sequence including carbonaceous shale, iron-stones, evaporite, carbonate and mafic to felsic volcanic units of the South Alligator and Finniss River Groups. This sequence has been subjected to regional greenschist facies metamorphism and multiphase deformation, which has resulted in the development of a northwest trending structural fabric. Subsequent widespread felsic volcanism and the intrusion of granitoids caused contact metamorphism, in aureoles between 500m and 2km wide that overprint the earlier regional metamorphism. After the granitoid intrusions an extensive array of northeast and northwest trending dolerite dykes intruded the metasedimentary sequence during regional extensional deformation.

Gold mineralization within the Pine Creek Orogen is preferentially developed within strata of the South Alligator Group and lower parts of the Finniss River Group along anticlines, strike-slip shear zones and duplex thrusts located in proximity to the Cullen Granite Batholith. Of particular stratigraphic importance are the Wildman Siltstone, the Koolpin Formation, Gerowie Tuff, Mount Bonnie Formation and the Burrell Creek Formation.

The Wildman Siltstone consists of medium to thinly bedded, to laminated fine grained pyritic carbonaceous sediments with minor sandstone and tuff beds, with an overall thickness of approximately 1,000m.

The Koolpin Formation consists of sulphidic and carbonaceous argillite, ferruginous chert, ironstone, silicified dolomites and phyllitic mudstones, which were deposited in a low energy environment. The contact between the Wildman Siltstone and the overlying Koolpin Formation is partially conformable and partially an angular unconformity. The Koolpin Formation varies in thickness from less than 300m to in excess of 1,000m, but its overall thickness is difficult to determine due to the presence of several intrusive sills of Zamu Dolerite, which vary from several meters to a few hundred meters in thickness.

The Burrell Creek Formation comprises a 1,500m thick sequence of turbiditic sediments including greywackes, siltstones and mudstones. The Mount Bonnie Formation is a transitional unit between the Koolpin and Burrell Creek Formations, comprising greywacke, carbonaceous siltstone, chert, tuff and ironstone and with a variable thickness between 150m and 400m.

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The Gerowie Tuff (GTF) is up to 400m thick and consists of tuff, tuffaceous chert and tuffaceous siltstone, with subordinate amounts of laminated cherts and carbonaceous siltstones.

Numerous semi-conformable sills of pre-orogenic Zamu Dolerite intrude the Koolpin Formation and the Gerowie Tuff. The post mineralization Burnside Granite and Mount Goyder Syenite intrude the sedimentary sequence.

The Northern Territory Operations area lies in the central part of the Pine Creek Geosyncline. Proterozoic rock units in the Burnside area comprise the Mt Partridge Group of the South Alligator Group and the overlying Finniss River Group.

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7.2

LOCAL COSMO MINE GEOLOGY

The Cosmo deposit geology is made up of a series of distal cyclical marine depositional events contained in a sequence referred to as the Upper to Middle Koolpin Formation. This formation consists of interbedded siltstones, carbonaceous mudstones, banded ironstone, phyllites, dolerite sills and greywacke units (Alexander, Kavanagh and Rolfe 1990).

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7.2.1      FOLDING

These units have been folded and faulted by a series of regional structural events during the Paleoproterozoic Pine Creek Orogen. This period of deformation lead to the formation of the local Burnside intrusion and the Cosmo Anticline, which is the main mineralization control at the Cosmo Mine. The Cosmo Anticline is a kilometer-scale tight, gently inclined fold that plunges between 50o to 75o northwards. The eastern fold limb is slightly overturned (east-verging). Numerous parasitic folds are present within the Cosmo Anticline, evident down to millimeter scale. Flexural slip and layer-parallel decoupling are apparent indicating significant local tectonic activation of bedding planes and contacts. The Cosmo Anticline and the smaller-scale folds, appear to be critical controls on localising gold mineralization. The variety in fold styles is illustrated in Figure 7-6.

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Figure 7-6 shows (a) isoclinal folds of early silica-pyrite veins localized in the hangingwall of an open fold with a faulted-out short limb; late quartz-pyrite infill of irregular veins is also evident. (b) Open fold with bedding parallel pyrrhotite locally reoriented to define a penetrative axial planar fabric. (c) disharmonic folds of bedding influenced by silica nodule; thinner layers show shorter fold wavelengths; carbonate-quartz veins crosscut bedding and folds. (d) Gentle to open folds of bedding in a parasitic fold hinge in siltstone; note sulfide remobilization into fold hinges (Beeson 2015)

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Note in Figure 7-7 sulfide remobilization in hinges and along limbs. (b) Closely to tightly folded siltstone with sulfide remobilization into hinges. (c) Parasitic folded chlorite-cordierite-garnet siltstone; note thickened fold hinges and thinned limbs.

Fold style is significantly influenced by host rock, and varies from open to isoclinal with attenuation, boudinage and shearing commonly evident along the short limbs of asymmetric folds. The carbonaceous sulphidic mudstone unit shows the most variety in fold style and orientation. This relatively incompetent unit appears to have localized strain resulting in the formation of mostly close to isoclinal folds (elastica folds are also locally evident), many of which are noticeably non-cylindrical. Folds in the carbonaceous graphitic mudstone unit commonly form at wavelengths of meters to centimeters. By contrast, folding in the underlying nodular greywacke and siltstone units is typically developed at much broader wavelength (meter to decameter scale), although domains of tighter folding at centimeter-scale are evident locally, particularly in the vicinity of the gold-mineralized zone.

7.2.2      HOST ROCK TYPES

There are four main dolerite sills in the area. Two outer dolerites that have intruded the upper younger sedimentary sequence; the Zamu Dolerite and an inner dolerite, now termed the “Phantom Dolerite”, which was mapped in the south end of the Phantom open pit (Figure 7-8).

The sedimentary package between the Zamu Dolerite and the outer dolerites is a series of carbonaceous sulfidic (pyrite+pyrrhotite) mudstones and banded siltstone units some of which contain boudinaged quartzite units, massive siltstones and non-graphitic mudstones.

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The Zamu Dolerite appears to have intruded a dolomite unit resulting in a thin quartz-carbonate vein marking the boundary of the Zamu Dolerite and the mineralization hosting siltstones. A completely recrystallized dolomite unit that averages approximately 7m in true width marks the inner contact of the Zamu Dolerite.

Within the dolomite are more siltstone horizons, many with chert interbeds up to 1cm in width with many in the early stages of being strained into boudinage. These inner siltstones are similar in appearance to those that host the mineralization lodes.

The carbonaceous mudstone is characterized by abundant (up to 70%) pyrite and pyrrhotite bands that have been deformed and attenuated in an erratic ductile manner. This unit is extremely carbonaceous and is used as a stratigraphic marker.

Mineralization at Cosmo Mine generally occurs within a package of pervasively metamorphosed siltstones between the Zamu Dolerite sill and the thick carbonaceous mudstone unit (Figure 7-10). The carbonaceous mudstone is identified as the “Pmc” unit. The contact of the Pmc is a useful stratigraphic marker with the siltstones and is a major control on mineralization in the 100 Lode on the Eastern Cosmo Anticline Limb.

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The Section in Figure 7-10 is based on Hole CE1055011 Collared in The Zamu Dolerite (West) and terminated in The ‘Pmc’ Carbonaceous Mudstone Unit.

Adjacent to the carbonaceous Pmc unit is a 10-20m wide banded coarser meta-siltstone / greywacke (“Pgt”) unit. Within the Pgt unit are irregular intervals containing commonly boudinaged and metamorphosed interbedded chert nodules, which are termed Boudinaged Greywacke and coded as “Pgtb”. The Pgtb unit host the main mineralization in the Cosmo deposit with over 85% of the contained gold ounces mined to date. The Pgt and Pgtb units appear to intercalate, but much of this is due to fold repetition on the limbs of the Cosmo Anticline. The Pgt/Pgtb units have the appearance of a greywacke and are logged as such. Past petrographic studies have indicated that the unit is more likely to be a recrystallized siltstone with regular compositional banding.

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Some of the finer mudstones are distinguished by the presence of 1-2mm chloritoid blebs and 1mm garnet psuedomorphs. Mineralization often occurs in association with the highly strained Pgtb unit, but it also occurs in banded siltstones.

Alteration/metamorphism is of the lower greenschist facies with some structurally related phyllite units and narrow aureoles of higher grade contact alteration related to the emplacement of the dolerite sills.

In Figure 7-12 above the Mineralization in:

  a.

Pgt Metasediment (Greywacke) unit containing quartz-carbonate-pyrite veins with ankerite- pyrite selvedges overprinted by chloritic shear vein associated with abundant euhedral arsenopyrite.

  b.

Pgtb (Nodular Greywacke) Unit containing silica nodules with euhedral arsenopyrite overgrowths after fine-grained pyrite localized along silica nodule margins.

  c.

Pmc Carbonaceous Mudstone Unit containing pyrrhotite and pyrite replacements oriented sub-parallel to bedding and within incipient fracture mesh.

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7.2.3      FAULTS

There are a series of crosscutting faults that displace the stratigraphy, with several faults forming key domain boundaries. The predominant fault is the F1 Fault, which crosscuts the entire Cosmo mineralized body in a broadly east-west direction. Additional knowledge gained in 2015 reveals the F1 Fault to be a curvi-planar structure, which splays out of the Western Limb of the Cosmo Anticline from west to east (Beeson 2015) along the major F2 Fault.

The F2 Fault is another significant structure, known only from drilling data, and is evident within the carbonaceous sulphidic mudstone sequence located along the Western Limb of the Cosmo Anticline. It has a north to locally north-northeast trending orientation, and is interpreted to generate a number of northeast migrating splays in the Western Lodes area and flatten slightly at shallower levels. This later aspect may suggest the F2 Fault becomes similar to the F1 Fault in orientation with depth and distance to the north away from the present mine development.

The F1 Fault is interpreted as an oblique-slip, planar thrust zone that shows up to 100m of displacement (Smith and Pridmore 2014) This fault separates the Cosmo Eastern Lode mineralization into a hangingwall and footwall zone and appears to have a significant impact on the distribution and tenor of gold mineralization along the Eastern Limb of the Cosmo Anticline. Other crosscutting faults also play a significant role in localising gold mineralization.

The F10 splay fault has been shown to have a very close spatial association with mineralization grade gold mineralization in the Eastern Lode system ( (J. Miller 2014). This fault reflects tectonic activation of the F10 marker unit, resulting in gross decoupling of its hangingwall and generating a series of tight to isoclinal folds and fault-bound slivers of the nodular greywacke host unit towards the hinge of the Cosmo Anticline; mineralization grade gold mineralization is located in the hinges of the decoupled folded host sequence as well as within planar segments of the host sequence preserved in proximity to the F10 splay fault. Miller (J. Miller 2014) also proposes a key control on high-grade mineralization imparted by steeply dipping northwest trending faults that crosscut the Eastern Lode system.

Current thoughts are that the F9 Fault at least pre-dates the F1 Fault, as the F9 Fault isn’t seen in the hangingwall where it would be otherwise. Displacement could be anywhere from a few meters to ~40m+, similar to the F9 Fault.

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7.2.4      METAMORPHISM

The sediments at Cosmo Mine have been metamorphosed up to amphibolite facies with common garnet and cordierite spotting with presence of hornblende, biotite, K-feldspar and quartz found throughout the mine and surrounds. A variety of retrograde mineral assemblages are present with preserved compositional banding reflecting original sedimentary bedding. Often apart from the preserved compositional bands, all other components of the protoliths were completely recrystallized to fine to medium grained, and locally porphyroblastic, metamorphic assemblages.

7.3 UNION REEFS GEOLOGY

Consulting geologist, Paul Karpeta (Karpeta 2011)was retained by Crocodile Gold to study the Union Reefs geology and structures. His comments, which the authors have reviewed and agree with, are as follows:

The deposits occur along a major NNW-SSE striking shear zone, the Pine Creek Shear Zone and are largely hosted by the Burrell Creek Formation slates and greywackes.

Bedding (S0) in the Union Reefs area is unimodal and subvertical averaging 87° to 255°, though way-up structures indicate the presence of tight, upright, isoclinal folds. Bedding strike varies along the Pine Creek Shear from 350° in Union North through 355° at Lady Alice and 335° at Crosscourse to 352° in Union South indicating a 20° swing in strike to the east at Crosscourse Pit corresponding to the swing in shear direction.

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Foliation (S1) in the Union Reefs area generally dips 80° to 270° and appears to be axial planar to the isoclinal folding (F1). Shallow plunging minor fold axial planes dip 80° to 280° and represent folds parasitic to the major 1st order folding. The plunge of these shallowly plunging folds varies between 33° to 114° at Lady Alice through 18° to 325° at Crosscourse to 15° to 145° at Union South indicating reversals in plunge from north to south related to cross folding.

Steeply plunging minor fold axial planes also dip 80° to the west but have plunges varying between 85° to 354° at Union North and Lady Alice through vertical (90°) at Crosscourse back to 85° to 353° at Union South. These steeply plunging minor folds (F3) appear to be later than the shallow plunging folds (F2) occurring in shear zones with a sinistral sense of movement.

At least two sets of bedding plane parallel lineations have been observed in the Union Reefs area, an earlier steeply plunging set (L1) and a later shallowly plunging set (L2). The earlier set is probably related to bedding plane movement during F1 flexural slip isoclinal folding. Plunge of L1 varies from 80° to 340° at Union North through 84° to 346° at Lady Alice and 80° to 142° at Crosscourse to 82° to 332° at Union South, again indicating a change in plunge along the Pine Creek Shear. The later shallowly plunging lineations (L2) appear to be related to the sinistral shearing event and plunge again shows a north-to-south variation. At Union North they plunge 7° to 170°, at Lady Alice 9° to 160° whereas at Crosscourse this changes to 15° to 330° reverting to 10° to 150° at Union Reefs South. Total quartz vein orientation data for Union Reefs shows an overwhelming sub-vertical north-south striking population, which may represent bedding plane and shear zone parallel veins.

The bedding plane parallel veins (QV1) are usually boudinaged parallel to the fold axes by bedding-plane movement associated with the flexural slip folding. The shallow plunges of these boudins show a north-to-south variation plunging 15° to 155° at Union North, at Lady Alice, 7° to 345°, at Crosscourse, 10° to 334° and 8° to 160° at Union South. The later shear zone parallel veins (QV2) plunge steeply varying between 80° to 340° at Union North through 85° to 345° at Lady Alice and 80° to 140° at Crosscourse to 80° to 335° at Union South. These variations in plunge of various structures are attributed to low amplitude E-W striking cross folding. Late north-south striking, E-over-W and W-over-E brittle thrusts are also recognized dipping 30° to 245°, flat and 45° to 070°.

The structural evolution of the Union Reefs area involved initially horizontal E-W compression (D1) and the formation of tight, upright, N-S striking, isoclinal flexural slip folding (F1) accompanied by bedding plane slip and boudinaged bedding plane parallel quartz veins (QV1). This folding would have been buttressed against the Pine Creek Fault Zone, which was originally a normal fault.

The Zamu dolerite sills were also folded by D1. Subsequently the horizontal compression direction rotated clockwise to NW-SE producing a sinistral shear couple on the Pine Creek Fault, which then became reactivated as a sinistral shear zone (D2). Subordinate sinistral shears formed on optimally oriented bedding planes either side of the main shear. However, it appears that at the aptly named Crosscourse Pit, the Pine Creek Fault has a left-hand extensional stepover, which forms an area of dilation. Vertically plunging quartz veins were formed on the hinges of sinistral shear folds within these shear zones (QV2). The horizontal compression direction then rotated clockwise to approximately north-south (D3), producing E-W striking, open (30° interlimb angle), long wavelength (~1,000m) folds (F3), which tilted previously formed lineations, boudins and mineralization bodies to the north or south depending on which limb of the fold they were on.

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Subsequently the Pine Creek Shear Zone appears to have been reactivated as a dextral shear though this was not directly observed in the field (D4) but has been documented elsewhere. The last deformation event was the conjugate, E-over-W and W-over-E, brittle thrusting (D5) possibly a result of horizontal E-W compression.

Gold mineralization at Union Reefs appears to be related to the two major sets of quartz veins and lodes (QV1 & QV2). The first set comprises bedding plane parallel veins having a shallow plunge parallel to the F1 upright fold hinges and produced by bedding plane movement on the hinges of these flexural slip folds. Gold mineralized quartz veins are associated with folding and are probably therefore similar to the saddle reefs reported worldwide from slate belts. The second set is found on the D2 sinistral shear zones but plunging steeply down them. Since the D2 shears are usually bedding plane parallel, both sets of veins are effectively bedding-plane parallel.

The intensity of gold mineralization appears to be highest in the Crosscourse Pit, where the Pine Creek Shear Zone undergoes a left handed extensional stepover. Such an extensional stepover would produce an area of better permeability to allow the migration of mineralizing fluids and is therefore similar to the model proposed for shear zone hosted gold deposits.

The turbiditic Burrell Creek Formation of Union Reefs has thick, more competent beds of greywacke in mudstones producing a series of large amplitude, long wavelength folds. These folds would have been initiated against a normal fault and propagated backwards (westwards). Therefore at Union Reefs the first and biggest fold to form would have been immediately to the west of the Pine Creek Fault (the Lady Alice Anticline). Subsequent rotation of the compression direction to NW-SE would result in a sinistral shear fault reactivation of both the Howley Structure and the Pine Creek Fault, and thrusting/folding in the Rising Tide and Hayes Creek Faults.

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Jaques Stacy (Stacy 2011) a consulting geologist with Taiga Consulting was contracted to reconstruct the Crosscourse Deposit using an extensive drill database which included AngloGold’s entire grade control drilling data set. His extensive comments on the Union Reefs area and more specifically the Crosscourse Deposit, which the Authors have reviewed and agree with, are as follows:

Mineralization in the Union Reefs area is confined to a 300m-wide section of the Pine Creek Shear Zone (PCSZ), a 300km-long, NNW- trending regional shear with an overall sinistral sense of displacement. Gold occurrences in the region are broadly classified as “Orogenic-type” deposits and occur in a wide variety of lithological and structural settings. In the Pine Creek/Union Reefs area, the PCSZ is located in a narrow embayment of sedimentary rocks sandwiched between two lobes of the Cullen Igneous Complex (CIC). Field evidence suggests that the PCSZ was active before, during, and after intrusion of the CIC. Many gold deposits in the area, such as Cosmo Howley and Enterprise, are intimately associated with fold axes in their sedimentary host rocks. Union Reefs is one of the few deposits, which does not display an obvious fold association, and mineralization is instead hosted by faults and shear zones oriented sub-parallel to bedding.

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Mineralization at Union Reefs seems to have been predominantly of the stockwork and sheeted-vein type, with lode-style veins comprising a lesser proportion of the deposit. Photos and descriptions of the E-Lens itself suggest that it is composed mainly of stockwork-type veining in a greywacke host and that elevated gold grades within the ore shoot occur due to the overlap of multiple generations of gold-bearing veins. The steeply plunging aspect of the E-Lens suggests that ore shoot location and morphology is strongly controlled by structural intersections.

Geological mapping at the mine site has identified three sets of steeply dipping shears, oriented 010°, 330°, and 355°, which both crosscut and run sub-parallel to NNW-trending bedding. When plotted on a stereonet, the intersection lineation established by the confluence of these structures plunges north to NNW at angles of 50-60°, depending on the dip of the intersecting structures. This is consistent with the observed orientation of the E-Lens ore shoot, suggesting that the intersection of structures was the primary controlling factor in the formation of this mineralized zone.

The planar distribution of the East and West lodes suggests that fault zones host them, but controls on the morphology of the E-Lens lode are not well understood. One possibility is that the E-Lens is hosted by a NW(?)-trending “transfer fault” that links the structures hosting the East and West lodes. This type of structural host may explain the narrow, plunging, pipe-like distribution of the E-Lens, and may have implications for future exploration in the Union Reefs area. If high-grade ore shoots are controlled by transfer faults between major shears, then these structures become high-priority exploration targets, especially if they are located proximal to known mineralized trends within the environment of the regional Pine Creek shear zone.

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A block model (above) illustrates the principle of a transverse fault. The combination of dip slip and strike slip movement opens up near vertical, pipe-shaped open spaces. In consideration of the fact that gold mineralizing system needs to access and permeate host rock this provides a possible explanation for the distribution of the E-Lens shape mineralization. The higher gold grades will deposit in the spaces with the most dilation while lower grades will form where mineralizing fluids permeate the surrounding less permeable host rock. In real life large voids or open spaces do not stay open like in the cartoon above but the effect on permeability will be the same as if they were open.

The pipe-like morphology of the E-Lens becomes sharply evident at a cut-off grade of 1.65g/t Au at which point the widely distributed, diffuse low-grade planar mineralization becomes focused into the main mass of the E-Lens. The transition from diffuse to focused mineralization is abrupt, occurring at cut-off grades between 1.63g/t and 1.65g/t Au.

Mineralization in the upper E-Lens remains as a contiguous pipe up to cut-off grades of 3.0g/t Au. With each successive increase in cut-off grade, the volume of mineralized material retreats toward the center of the E-Lens. This reflects a significant grade zonation, with the central core of the mineralization shoot containing the highest grade material. The cause of this zonation is unclear, but may be related to the intersection of fault structures in the core of the E-Lens. There is a gap in the 3.0g/t Au grade shell between the 1,025 and 975m levels of the Crosscourse pit but high-grade material reappears beneath this level and seems to continue below the pit floor as the lower extension of the E-Lens. The discontinuous nature of the 3.0g/t Au grade shell may reflect a propensity for pod-shaped to cigar-shaped mineralization shoots in the Union Reefs system. If this is the case, then the lower extension of the E-Lens may represent the upper portion of a second high-grade mineralization “pod” underlying the mined-out segment of the E-Lens within the pit. Before mining, the upper 3.0g/t Au pod had a plunge extent of about 250m, and similar dimensions are expected to be encountered in the lower section. Currently, the lower extension of the E-Lens is defined

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over a plunge extent of ca. 170m but more drilling is needed to determine in detail the true extent of the zone.

 

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7.3.1      UNION REEFS DEPOSITS

The Union Reefs deposits occur along the northwest striking Pine Creek Shear Zone about 15km north along strike from Pine Creek (Figure 7-21) and like these deposits is closely flanked to both the east and west by the Cullen Batholith. The area contains numerous historical workings developed over an area 3.5km in length, 400m in width and a vertical depth of 300m. The Union Reefs mill, where all Newmarket Gold’s mineralization is processed, is located here and during mining from 1995-2003 a total of 202Mt of mineralization from the Union Reef deposits at an average grade of 1.47g/t Au were estimated to be have been processed (Ahmad, Wygralak and Ferenczi 2009)).

These workings follow the NW trend of the Pine Creek Shear Zone and are located on two smaller sub parallel NW trends within tightly folded inter-bedded metamorphosed greywackes and shales of the Burrell Creek Fm. Folds are tight, upright, recumbent and asymmetrical due to several folding events. The eastern trend is known as the ‘Lady Alice Line’ and the western trend is the ‘Union Line’ (Figure 7-22).

The Lady Alice Line mineralization occurs in a sub vertical shears on the western limb of the large Lady Alice anticline and is parallel to the axial plane; deposits include Millars, Ping Que, Lady Alice and Lady Alice North.

The Union Line is a steeply east dipping shear on which the Union South, Union Central, Prospect and Union North are located. The Crosscourse Deposit is the largest known deposit within the Union Reefs area and occurs where the two trends are proximal to each other. A cross cutting mineralization zone spans the shorter distance between the two structures and is interpreted to be a shear jog dilatational system later exploited by mineralizing fluids (Hellsten 2001)

The position of the pits along strike is believed to correlate with a series of crests along the anticlinal trace, which occurred during horizontal NNW-SSE shortening. The biggest of these being at Crosscourse; this is supported by changes in plunge of the axial trace observed by Hewson (Hewson 1997).

Currently the most prospective deposits within the Union Reefs area are the Esmeralda and Prospect deposits.

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7.3.2      ESMERALDA PROSPECT

The Esmeralda deposit has three main areas of gold mineralization, Esmeralda A, B and C, together with a single known base metal occurrence, Caroline.

The local geology of the Esmeralda and Caroline areas comprises an NNW-striking tightly folded inlier of Mount Bonnie Formation slates, greywackes and cherts, conformably overlain to the west by the younger Burrell Creek slates that host the Union Reefs Mines and to the east by the Allamber Springs lobe of the Cullen Batholith (Figure 7-25).

The stratigraphy of the Adelaide River-Pine Creek area (Figure 7-3) comprises Lower Proterozoic sedimentary and volcanic rocks of the South Alligator group and the Finniss River Group. Locally, underlying the South Alligator Group are the conglomerates, arkoses, siltstones and graphitic shales of the Mount Partridge Group. The South Alligator Group consists of the basal Koolpin Formation overlain by the Gerowie Formation and capped by the Mount Bonnie Formation. The overlying Finniss River Group in the area comprises the Burrell Creek Formation. The entire sequence is intruded by the Zamu Dolerites and the Cullen Batholith granites (Glass 2010).

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Gold mineralization was detected in four areas, the already drilled Esmeralda A and B areas, an area of shearing south of Caroline (Esmeralda C), indicated by geochemical anomalies and rock chip samples and an area to the NW of Esmeralda A and the NE of Esmeralda B (here termed Esmeralda D). Base metal mineralization has been noted at Caroline prospect south of Esmeralda B.

Esmeralda A: The gold mineralization at Esmeralda A occurs in a series of NNW-SSE striking, bedding plane parallel quartz-tourmaline veins associated with pyrite-sericite alteration in a sequence of alternating slates and greywackes. These veins are thought to have formed during an episode of dextral strike-slip movement between a series of right-lateral step-overs. However, evidence of sinistral movement was also seen on these faults and the mineralization could instead have been formed during left-lateral movement similar to Esmeralda B. The extent of this gold mineralized vein system is governed by a WNW-ESE striking cross fault to the north and the hornfelsed aureole of the Allamber Granite to the south.

Esmeralda B: The gold mineralization at Esmeralda B occurs in a series of NNW-SSE striking, bedding-plane parallel quartz veins in an alternating slate-greywacke sequence. Gold mineralization also occurs in NE-SW striking sinistral cross fractures and in the culminations of parasitic folds as pyrite-sericite alteration. Little or no tourmaline appears to be present. This mineralization appears to have formed during an episode of sinistral strike-slip movement between a series of left-lateral step-overs (Figure 7-15). The extent of this mineralization appears to be cut off to the north by the same WNW-ESE striking cross fault as Esmeralda A. The southern end of Esmeralda B is not constrained but disappears under a cover of siliceous rubble towards Caroline Hill.

Esmeralda C: The gold mineralization at Esmeralda C occurs in a NNW-SSE striking sinistral 5m wide shear zone cutting through a 20m thick greywacke and is associated with pyrite-sericite alteration that was picked up by soil and chip sampling. This mineralization appears to be limited by crossfaulting to the north but is unconstrained to the south, an area covered by siliceous rubble.

Esmeralda D: The gold mineralization at Esmeralda D was located by chip sampling (up to 0.3g/t Au) and comprises pyrite-sericite alteration in the culmination of a major NNW-SSE striking anticline. It is cut off to the south by the same WNW-ESE striking cross fault, as Esmeralda A but is unconstrained to the north.

7.4 PINE CREEK GEOLOGY

Gold mineralization at Pine Creek occurs in two distinct domains; within shear structures and along the axial plane of close anticline structures. Mineralization is hosted in clastic meta-sedimentary wall rocks. Quartz sulphide veins, with varying sulphide content, host the majority of the gold with smaller quantities occurring disseminated in sulphides in the wall rock. In primary mineralization between 2 and 50% of gold occurs as free gold, the rest is occurring as 2 to 30µm inclusions within sulphides, primarily arsenopyrite and pyrrhotite (partially refractory).

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Gold grades from mineralization hosted in the shear structures tend to be lower and patchier than the grades observed along the anticlinal axial planes. Mineralization at the Enterprise deposit, the largest of the gold deposits, is hosted in greywacke along the axial plane of the Enterprise anticline.

Along many of the anticline structures are quartz saddle reefs occurring in the hinge zones. These structures are hosting the highest grades and are generally surrounded by a zone of lower grade mineralization, occurring as scattered centimeter scale quartz sulphide veining. Veining is often concordant, though it is also sometimes discordant and some deposits display several different styles of veining in the axial zone. This style of mineralization has been described in “Gold in Greywacke in Anticlinal Crest (GIGIAC) Mineralization” (Shields 1994).

Pervasive, intense to moderate chlorite wall rock alteration is present around the mineralized zones, especially in hinge zones. Closer to the mineralization bodies, silica alteration is usually observed, along with disseminated sulphides and sometimes k-feldspar mineralization.

Quartz veining hosting gold is this style of mineralization. The shears are sub-vertical to steeply dipping reverse fault structures, where movement can be inferred. The movement planes of the shear structures are orientated sub parallel to oblique to the axial planes of the folds. Where observable the orientations are between 315° and 340°.

Shear hosted deposits within the project area are generally smaller, lower grade and patchier than the anticline hosted deposits. Modern mining has not specifically targeted any shear structures. However, smaller shears have been shown to host gold within larger deposits. One example of this is the Eastern Fault Zone within the Enterprise deposit.

Wall rock alteration around mineralization in shears is silica alteration within a larger zone of moderate to intense chlorite alteration. Sulphide dissemination around veining is also a feature within the shear zones. This is similar to the wall rock alteration observed in the anticline hosted mineralization.

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7.4.1      ENTERPRISE DEPOSIT

Enterprise is the largest mine in the Pine Creek area. Pine Creek Goldfields Limited last mined it from 1985 to 1993. The Enterprise pit sits at the center of the Pine Creek deposit area and is over 880m long by about 300m wide in the center. Its deepest point is over 160 meter below the original land surface. estimates taken from government and internal mine reports suggest around 80% of total local production has been sources from the Enterprise/Czarina deposits located at Pine Creek.

The anticline that hosts the Enterprise deposit has been termed the Enterprise Anticline. It is a relatively symmetrical structure with limbs dipping between 60° and 70°. The axial plane of the Enterprise Anticline has a strike of about 315°. The fold axis plunges to the southeast but this plunge is not constant and becomes steeper as the fold moves further southeast. In the northern parts of the deposit this plunge is between 0° and 10°, this increases to about 30° in the southern parts.

Several faults were encountered in the Enterprise pit; a fault zone that has been termed the Eastern Fault Zone is the largest of these. The Eastern Fault Zone strikes sub-parallel to the Enterprise Anticline along the eastern margin of the deposit (See Figure 7-26). This fault has been traced for over 600m with a dip of between 60° and 70° to the west.

Two types of wall rock alteration in the mineralization zone have been identified.

  1.

Silicification with the development of biotite and chlorite.

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  2.

Alteration resulting in a mineral assemblage of K-feldspar, arsenopyrite and minor pyrite and pyrrhotite.

The southern extension of the axial plane between the pit crest and the boundary of MLN13 was previously subjected to small programs of RC drilling. Greywacke, siltstone and grit of the Kohinoor Grit Member was intersected. Mineralization was extremely weak and spotty with few samples in excess of 0.5g/t Au and was fault-shear related. It is likely that the potential of this area is deeper than so far tested.

Over 9Mt of ore was extracted from the Enterprise Pit when it was operated. An historic remaining mineral resource (1) was estimated in 2004 by Burnside Operations Pty Ltd of 1,394,982t of mineralization at an average grade of 2.65g/t Au containing 119,040oz of gold. This mineral resource lies below the pit floor, which in now underwater and was rated as a ‘high risk’ mineralization target because the pit would need to be widened to be mined safely. The mineral resource review conducted by Newmarket Gold in 2012 down sized the inferred mineral resource to 1,061,000t of mineralization at 2.6g/t Au and 87,000oz of gold, relating to what could be economically extracted.

Even though there is an inferred mineral resource of over 1Mt remaining at Enterprise it is suspected that the average grade is too low, at 2.6g/t Au, to support an underground operation. If the mineral resource could be extended at depth or linked to the South Enterprise Deposit then further work is warranted. The dimensions of the Enterprise model are approximately 800m along strike, 200 down dip and 30m wide.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves.

7.4.2      SOUTH ENTERPRISE DEPOSIT

The area just to the south of the Enterprise pit has been an area of exploration interest since the 1980’s. Significant amounts of drilling have been conducted and a mineral resource has been outlined at South Enterprise deposit. Grades begin at surface and the area represents a potential open cut operation.

The South Enterprise deposit sits on the southern margin of the Enterprise deposit within the Enterprise Anticline.

Gold mineralization is relatively patchy and weak compared to other mineralization bodies in the Goldfields. This indicates that the mineralization is more likely related to a shear than to the anticline structures that other deposits are related to.

Nine different loads were identified by Makar (Makar and Muller 2006)in the South Enterprise deposit ranging in size from 17,000t to over 100,000t.

The dimensions of the South Enterprise deposit are 400m along strike, 230m down drip and 20m in width.

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7.4.3      CZARINA AND SOUTH CZARINA DEPOSITS

Czarina deposit is a zone of mineralization that was mined along with the Enterprise deposit in the early 1990’s by Pine Creek Gold Fields Ltd. The pit was back filled and rehabilitated when mining was finished. A mineral resource remains under the old pit floor at Czarina.

The Czarina deposit is to the east of the Enterprise deposit, situated in an anticline roughly 200m to the east and sub-parallel to the Enterprise Anticline. This structure has been termed the Czarina Anticline. Mineralization is hosted in the western limb of the gently southwards plunging Czarina Anticline, which has an inter limb angle of 30°-40° degrees and a strike of 315°. The axial plane of the anticline dips steeply to the west.

The deposit at Czarina is hosted in contact metamorphosed greywacke, mudstone and siltstone as are most other deposits in the Pine Creek area. Gold is hosted in quartz sulphide veins that make up between 5% and 20% of the rock in the zone. As with the mineralization in the Enterprise deposit, up to 50% of the gold is occurring as inclusions within sulphides, the rest is free. The main sulphides present in the deposit are pyrite, arsenopyrite, pyrrhotite, galena and chalcopyrite. The best zones of mineralization are associated with increased silicification, quartz veining and pyrite, mostly within siltstone units.

Mineralization is not well understood at South Czarina deposit and preliminary work inferrs a much smaller mineral resource than at the Czarina deposit.

The dimension of the Czarina deposits is 600m in strike, 110m down dip and 40m wide.

7.4.4      COX DEPOSIT

The Cox deposit is located to the north of the Bashi Bazook deposit. It is located in rugged terrain covered in numerous old workings. As with Bashi Bazook the orientation of the mineralized zone is oblique to the anticline structures that dominate the Pine Creek area.

Diamond drilling programs were completed that had the objective of targeting the 40m to 60m wide mineralized zone that dipped west at 75 to 80 degrees within the southern limits of the pit.

High gold grades were known to be associated with linear/planar structures 100-150m in strike extent and relatively continuous down dip and the intersection of planar structures that form pods of variable orientations and average dimensions of 15m x 15m.

Six deep holes were targeted beneath the pit, however, these only met with narrow modest gold grades. There is up to 100m of untested zone beneath the pit floor to the RL of the deep holes. Testing of this zone would require large step backs and shallow dips. Mineralization is hosted in the Cox’s Shear, which strikes at roughly 335°, sub-parallel to Bashi Bazook. The shear dips between 65° and 80° to the west. Marjoribanks (Marjoribanks 1993) concluded that Chinaman’s deposit is hosted in the same shear structure at its southern extent. The Cox’s Shear cuts across the Czarina and Kohinoor anticlines and this is where the highest gold grades have been observed. Cox’s Shear is characterized by black gossanous material at the surface, breccia development, quartz veining, intense chlorite wall rock alteration and disseminated sulphides.

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The dimensions of the Cox deposit are 300m in strike, 130m down dip and 40m wide.

7.4.5      BASHI BAZOOK DEPOSIT

Battery Shear/ Bashi Bazook is a zone of mineralization located between Cox’s Shear and Chinaman’s Shear and is probably on a common structural set. The deposit has been referred to as Battery Shear, Bashi Bazook and sometimes as Battery Shear in the south and Bashi Bazook in the north, for the purposes of this report the deposit as a whole has been called Bashi Bazook. The area is covered with numerous shafts and historic workings.

Mineralization has been interpreted as occurring along a shear on the western dipping limb of the Enterprise Anticline. The strike of the shear has been interpreted as striking at about 335°, or roughly 20° oblique to the anticline structures that dominate the project area. It is difficult to determine which anticline it is sitting in as the structures have been offset by faulting. Different interpretations of the orientation of mineralization have been made; Marjoribanks (1993) (Marjoribanks 1993) interpreted it as a vertical to steeply, westerly dipping shear structure hosting mineralization. The mineralization at Bashi Bazook occurs in the sediments of the Kohinoor Grit and the underlying Upper Mine Greywacke.

7.4.6      KOHINOOR DEPOSIT

The Kohinoor deposit sits directly south of the Cox deposit in the mining lease MCN523. It is a long zone of mineralization sub-parallel to the anticline structures that dominate the Pine Creek Goldfields. The Kohinoor deposit has been divided up into several deposits in the past, the northern portion being known as Henry George, the central portion known as Kohinoor and Jensen’s in the south.

Mineralization is concentrated in the axial zone of the Kohinoor Anticline, which is an upright structure with limbs dipping between 60° and 80°. The sequence is made up of inter-bedded greywackes, siltstones and conglomerates of the Kohinoor Grit and the Upper Mine Greywacke. Mineralization occurs in three domains;

  1.

Saddle reef structures in hinge zone of the Kohinoor anticline, this sits between 11450N and 11525N and host most of the gold.

  2.

A domain of veining to the west of 11450E, numerous pits and shafts fall in a line along this domain.

  3.

A third domain is characterized by a zone of strong quartz veining; this domain is thought to be lower grade.

The dimensions of the Kohinoor deposit are 1,000m in strike, 120m down dip and 60m in width.

7.4.7      ELEANOR DEPOSIT

The Eleanor deposit is situated along a collection of shallow workings and underground workings in the southeastern leg of MLN13. Roughly 200m true north is the Jensen’s adit, at the Kohinoor deposit and workings in the Eleanor deposit are orientated sub parallel to workings at Kohinoor. It has been recorded that the Eleanor Mine was a relatively high-grade producer in the Pine Creek Goldfields, though tonnage was relatively low.

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The mineralization at the Eleanor deposit is hosted in a shear structure that strikes at roughly 320°. The Eleanor shear has been interpreted as a steep SW dipping reverse fault. The fault has been traced for over 300m along strike from grid 9,800N to 10,140N. The actual extent of the mineralization may be over 400m along strike, however, there is insufficient drilling to either confirm or disprove this. A trench was cut across the shear to gain a better understanding of its geometry. Twelve meters at 0.97g/t Au were encountered in this trench. Mineralization appears to be occurring as shoots of high-grade that dip steeply to the south.

7.4.8      ELSINORE PROSPECT

The Elsinore prospect sits in MCN523 between 9,100N and 9,250N local grid. This is generally a lower grade zone of mineralization. However, this mine was one of the first hard rock workings developed in the Pine Creek Goldfields. The prospect has had extensive historic work done, in the form of both surficial workings and underground workings. Heritage listed sites relating to Chinese mining exist around the prospect and require minimal disturbance.

Mineralization at the Elsinore prospect is hosted in the eastern limb and axial zone of the Kohinoor Anticline. The host rock is inter-bedded greywacke and mudstone in a Bouma turbidite sequence. Interpretations made from old workings indicate that the mineralization was occurring mostly in two quartz reef structures. One was an easterly dipping, bedding conformable quartz reef on the eastern limb of the anticline. The other was a sub vertically dipping quartz reef on the in the central/ western zone of the Kohinoor Anticline. The Dashwood shaft passes below the deposit, this indicates that the reefs do not continue at depth.

Weak to moderate chlorite alteration occurs through the hinge zone of the anticline and disseminated sulphides (pyrite and arsenopyrite) exist in the wall rock of the mineralization.

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7.4.9      INTERNATIONAL DEPOSIT

The International deposit is located in the Czarina Anticline, which is to the east of the Enterprise Anticline, which hosts the Gandy’s deposits. The deposit is located in the Lower Proterozoic sediments of the Burrell Creek Formation.

The local geology related to the International deposit can be determined from past mining activities and is generally summarized as;

Spotted Siltstone Horizon

This unit is a 70m thick predominantly siltstone sequence overlying the Lower Mine Greywacke. The siltstones are fine grained, well bedded and contain bedding concordant chert bands and nodules. Cordierite spotting is common throughout. At the base is an 18m thick unit termed the Nodular Chert unit. This is treated as part of the Spotted Silt horizon; however, it contains abundant chert bands and is host to the Enterprise saddle reef. The base of this unit is marked by a distinct 1.5 -2.0m thick, coarse cordierite spotted and bleached silt bed.

Lower Mine Greywacke

The Lower Mine Greywacke overlies the Gandy’s Silt Horizon and is the basal unit observed within the Enterprise Mine. This unit consists of fine to medium grained greywacke, which varies greatly in exposed thickness from 55-110m. It is similar to the Gandy’s Hill Greywacke as it contains abundant mica flakes, however, it contains more common siltstone inter-beds than the former.

Metamorphic Interpretation of Lower Mine Greywacke

Summary of (P. Ashley 2013)

Prograde metamorphism displays random orientation, indicative of a contact, thermal intrusive influence. Biotite to bitotie hornfels metamorphism is present but the presence of garnet infers a higher maximum metamorphic grade of hornblende-hornfels.

Retrograde metamorphism is thought to be hydrothermal related as indicated by quartz- sulphide veining. Typical retrograde alterations of host rocks are sericite-muscovite with small amounts of sulphides, rutile and leucoxene.

Vein composition is thought to be non-uniform with early veining dominated by medium to coarse quartz grains with trace to abundant sulphides and trace chlorite. Commencement of veining is thought to be at a higher metamorphic grade indicated by biotite on vein margins, which is absent within the vein mass.

Pyrite is the primary sulphide in the vein, although locally arsenopyrite can predominate. Both sulphides appear to be related to hydrothermal alteration assemblages.

Gandy’s Silt Horizon

The Gandy’s Silt horizon is a similar unit to the Lower Gandy’s Silt. It consists of fine-grained, well bedded siltstones with minor greywacke beds. Chert bands and nodules are present, however less numerous than in the Lower Gandy’s Silt. This horizon hosts the South Gandy’s saddle reef at Gandy’s Hill, and is the basal unit intersected at the International deposit.

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The sulphide minerals, which have been recognized at Pine Creek, include pyrite, arsenopyrite, galena, sphalerite, chalcopyrite, bismuthinite, tetrahedrite, covellite and marcasite. Gold occurs in places as discrete accumulations, but also as inclusions in arsenopyrite and pyrrhotite, and as intergrowths with bismuthinite in massive pyrrhotite. Some pyrrhotite is recrystallized to pyrite and can contain gold.

Generally, gold is associated with quartz and/or sulphide mineralization. Some of the massive reefs can contain up to 80% sulphides, although the sulphide content is more generally of the order of 10% to 30%. Much of the sediment contains disseminated sulphide as fine crystals.

However, gold and sulphide mineralization is not confined to quartz, and high gold values can occur in samples completely devoid of quartz. Also, some quartz is totally barren of gold and sulphides.

Three main types of mineralization are evident at International deposit, these are:

In gross terms the majority of the mineralization is restricted to the greywacke (Lower Mine Greywacke) on the west limb of the anticline axis and east of the synclinal fault.

The vein systems comprise quartz, and quartz-sulphide veins and micro-veins separated by wall rock. Veins vary in thickness from micro-veins of <1mm to veins up to 50cm thick, with an occasional thick macro-vein of up to 1-2m. Pyrite is the dominant sulphide present with lesser amounts of arsenopyrite and pyrrhotite. Sulphides occur within quartz veins and as mono-mineralic sulphide veins, and also disseminated through the wall rock adjacent to veins. Chlorite alteration is pervasive particularly within the siltstones adjacent to the fault. Gold appears to be mainly associated with veining and grade of up to 50.0g/t are present but typically grade 1.0 -1.5g/t in oxide material and 2.50g/t in the primary zone.

7.4.10      GANDY'S DEPOSIT

The stratigraphy at North & South Gandy's consists of turbiditic sequences of siltstones and greywackes, with minor mudstones and tuffs. At South Gandy's three units have been identified (Fawcett 1993):

These units are similar in composition to those seen at North Gandy's.

At North Gandy's deposit, two major units have been identified from diamond drill core:

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The greywacke is medium grained, contains minor siltstone and is generally weakly mineralized. The lower siltstone unit hosts the bulk of the mineralization. It is a fine-grained, thinly bedded siltstone with minor mudstones and greywackes, which form cyclical upward fining sequences within the main siltstone unit.

The turbiditic sequences have an overall fining upward cycle from greywacke to siltstone with shale tops and contain 'sub-cycles’ of a similar nature of approximately one meter in thickness. Bedding features such as load structures are present in the siltstones

Several styles of mineralization in both the oxidized and primary rock are demonstrated at Gandy's Hill deposit with the bulk of the mineralization contained within siltstones. These styles include:

Two saddle reefs have been observed in outcrop at Gandy's Hill. One (the Gandy's Hill reef) is situated near the Gandy's Hill trig, and the other (the North Gandy's Hill reef) is located 500m to the north. The North Gandy's Hill Reef is well developed near surface where it is up to 5m in width, and becomes poorly developed at depth. The reef is banded, pale greyish white quartz containing 2-5% veinlets and disseminated pyrite and arsenopyrite with minor chlorite and gold grades in the range 1.0 -12.0g/t Au. The saddle reefs are hosted by cordierite spotted siltstones containing 1-2% disseminated sulphides. The Gandy's Hill reef is very poorly developed in the South Gandy's area. At South Gandy's and at depth on North Gandy's, mineralization is expressed as a zone of intense stockwork veining near to or across the anticlinal axis. Quartz veins in this zone have numerous orientations and are up to lm thick with an average of 5-10cm. They contain 5% veinlet and disseminated pyrite, arsenopyrite, chalcopyrite and minor pyrrhotite, galena and sphalerite. Gold grades are typically in the range 1.0 to 15g/t Au.

Narrow quartz veins peripheral to the fold axis are contained within siltstones and minor greywackes and form a minor proportion of the total mineralization volume. The veining is either sub-parallel to or crosscuts bedding and has been observed mainly in the oxide zone. Veins contain up to 1% disseminated sulphides and have gold grades in the range 0.9 - 3.0g/t Au.

Fault related mineralization also makes up a small proportion of the total mineralization volume. It is located within or near major fault zones and occurs mainly in siltstones. The siltstones are highly foliated and silicified, and contain numerous narrow fracture-fill quartz veinlets and veins generally less than 1cm thick. The veins carry up to 1% disseminated pyrite, arsenopyrite and chalcopyrite and typically grade 1.0 to 8.0g/t gold.

7.5 BURNSIDE GEOLOGY

Gold mines within the Burnside area have been responsible for a large portion of the historical gold production in the Pine Creek Orogen (Figure 7-37). Early prospectors first located most of the recent and current modern gold mines in the late 1800’s when alluvial production was significant. Today these occurrences and mines contain the bulk of Newmarket Gold’s mineral resources.

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Although there are many similarities amongst the deposits described below, with most having some structural control, each is uniquely different in its structural setting and style of mineralization.

7.5.1      HOWLEY LINE DEPOSITS

A description of the Howley Anticline is a macroscopic NW swinging to north trending, asymmetric, tight, non-cylindrical fold with a strike length of 30km, located 5-10km from the southwest and western surface boundary of the Burnside Granite.

The Howley anticline hosts the largest mineralized deposits in the area. Due to the doubly plunging anticlinal structure, the sediments exposed at surface change along the axial trace younging from Cosmo-Howley to Big Howley deposits, reversing south of Bridge Creek deposit and again around Mt Paqualin deposit (Figure 7-29). The stratigraphic position of these deposits is important as this controls the style of mineralization.

Parallel fold axes lie east and west of the Howley anticline, the hinge zone of these anticlinal structures can also host gold mineralization.

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7.5.2      HOWLEY BIG PIT

Located about 2.0km NW of the Cosmo Mine, stretching for approximately 5km along a NW trending part of the Howley anticline is a historically rich Chinese-Howley alluvial gold mining site. Modern hard rock mining occurred in the late 1980’s/early 1990’s from several small pits with further mining by Crocodile Gold from late 2009–2011. During 2010 a total of 1.26Mt of ore was mined from the Howley area and was the main source of gold for operations during this time.

The Howley Big Pit is a combination of the pits in this area and includes (from south to north) Howley South Extended, West Howley (mined by Crocodile Gold), Chinese-Howley pits 1, 2 and 3, Mottrams (mined by Crocodile Gold) and Big Howley. Historically, these were reported separately but are now combined due to the geological and spatial relationship of these deposits. These deposits occupy a shallow 10° NW plunging portion of the Howley anticline resulting in surface stratigraphy to young in this direction. All mineralization in this area is structurally controlled (due to absence of the highly carbonaceous sediments of the Koolpin Formation at Howley), but the change in stratigraphic position has resulted in slightly differing geology and different structural hosts for mineralization within Howley Big Pit. The whole area is mostly depleted at surface, but drilling indicates that mineralization continues at depth.

Mineralization is hosted in thinly laminated, moderately carbonaceous pelites and siltstones and more massive thin beds of tuff and chert of the middle Gerowie Tuff Formation. The Gerowie Tuff was conformably intruded by the Zamu Dolerite and other smaller dolerite sills, which were then folded during D2 to form the Howley Anticline. Three structural hosts for quartz-sulphide vein mineralization are identified: folded Zamu Dolerite, axial planar shears and “saddle reef” style.

The Zamu Dolerite associated mineralization is close to surface in the south, i.e. within the Howley South extended pit. Here, the brittle dolerite was fractured during D2 folding and later in-filled with auriferous fluids, forming the “saddle reef” geometry of these lodes. However, the highest grades (~2g/t Au; (F. V. Muller 2008) occur in quartz stockworks along the lower contact of the dolerite, particularly on the western limb of the anticline. Due to the north plunging anticline the Zamu Dolerite and associated mineralization plunges deeper below surface to the north, below Chinese-Howley 1, 2, 3 and Mottrams deposits where mineralization is then hosted at surface by axial planar shears.

Axial planar shear mineralization is controlled by D2 thrust faulting on the limbs of the Howley anticline and preferentially occurs in moderately carbonaceous pelites. Lodes are sub-vertical, sub-parallel to bedding and occur in a series of stacked planar shoots. In Chinese-Howley pits 1 and 2, shears are proximal to the fold hinge/close of the Howley anticline. West Howley, Chinese-Howley No.3 and Mottrams deposits follow shear zones along the western limb of the Howley anticline.

The geometry of mineralization lodes is complicated by northerly striking duplex thrust zones, which occurred during the main folding event (D2) as well as deformation from subsequent events (D3-D4). This has resulted in structurally complex lodes which pinch and swell creating quartz ‘pods’ of mineralization and in some cases can terminate the mineralized lode such as in Chinese-Howley pit No. 1.

Big Howley occurs 3km NW along strike from the Chinese-Howley pits; the structural setting for mineralization is different again. Mineralization occurs in the anticline hinge as 1-2m wide quartz-sulphide stockwork zones in proximity to saddle reefs (A. K. Sener 2004). There are 3 types of lodes: vertical eastern lodes, central lodes (dipping 40° to 70° to the west) and western lodes (dipping 60° to the west). The westerly dipping lodes have the highest average grade of ~3.5g/t Au (F. V. Muller 2008). All are hosted in the Gerowie Tuff Formation.

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Conceptual geological modeling by past operators, taking into account the stratigraphic relationships of the Koolpin Formation, Gerowie Tuff and Mt Bonnie Formations together with local structural evidence, has highlighted the possibility of continuations of the Cosmo gold system beneath the conceptual Howley Big Pit Area (Figure 7-30) illustrates the integrated Cosmo-Howley model and show target zones that should be tested in future exploration drilling. Significantly, there is virtually no effective historical drilling in the gap between Cosmo and the Howley area.

 

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7.5.3      PAQUALIN GROUP OF TENEMENTS

The Paqualin Group comprises deposits along the 12km north-south trending portion of the Howley anticline, 5km to the west of the Burnside Granite. The Bon’s Rush deposit is the northern most prospect with mineralization sitting in the Howley anticline. Kazi and Western Arm deposits occur on folds parasitic to the Howley anticline.

Along the 12km north trending strike length, the Koolpin Formation, Zamu Dolerite, Gerowie Tuff and Mt Bonnie Formations of the South Alligator Group are exposed as a gross domal structure on the Howley anticline, which plunges north at Bon’s Rush and south at Bridge Creek.

Bon’s Rush deposit was not discovered until the late 1990s as it is concealed beneath 1-5m of black soil. In 2001, Northern Gold N.L. delineated an inferred mineral resource of 540,000t at 2.51g/t Au for a total of 43,400oz (1) ( (Hardy and Hague 2001a)). Gold is interpreted to occur in the hangingwall of the Zamu Dolerite. The Zamu Dolerite is positioned stratigraphically between the carbonaceous shale from the Upper Koolpin Fm, and tuff, shale and chert beds from the Gerowie Tuff Formation. Mineralization occurs on the eastern limb in the form of shear controlled fractures and associated crackle breccia type mineralization. Due to the similar position in the stratigraphy to Cosmo and Bridge Creek deposits, stratabound style mineralization should be considered as an exploration target at Bon’s Rush.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves.

The Kazi deposit is located on the eastern limb of a north trending syncline immediately east of the Howley anticline structure. The deposit is almost entirely concealed beneath 1-5m of black alluvium; it was discovered as a gold soil anomaly during exploration conducted in the 1980’s. The deposit is located in a thinly bedded sequence of inter-bedded laminated rhyolitic tuff, chert, tuffaceous siltstone and minor greywacke of the Gerowie Tuff, which has been conformably intruded by the Zamu Dolerite. Gold occurs in shear parallel quartz veins and en-echelon veins in the hangingwall of moderately west dipping thrust faults ( (Parrington and McNaughton 1997)). Some minor mineralization extends into the Zamu Dolerite. The main zone is interpreted to be a moderately west-dipping, north-striking, tabular high-grade lode approximately 200m in length, sub-parallel to bedding, within the hinge and east limb of an easterly overturned anticline. Gold mineralization remains open at depth.

The Western Arm deposit is located on a parasitic domal anticline structure running parallel to the Howley anticline approximately 4km to the west of Bridge Creek deposit. The main mineralized zone extends approximately 1,200m along strike and up to 50m in width. Western Arm lies stratigraphically higher than Bridge Creek deposit occurring on the contact between a major sequence of greywacke, siltstone and mudstone in the hangingwall (Mt. Bonnie Formation) and carbonaceous shale, sulphidic shale, tuffaceous mudstone, nodular mudstone (Gerowie Tuff) in the footwall ( (Hardy and Hague 2001b)). Gold mineralization at Western Arm occurs as ‘saddle reef’ style in a series of quartz-sulfide stockwork lodes semi-conformable to bedding, and is best developed in mudstone and siltstone units in the hinge and eastern limb positions of the fold.

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7.5.4      FOUNTAIN HEAD - TALLY HO GOLD DEPOSITS

The Fountain Head deposit was discovered in 1883 and was subject to intensive elluvial mining activity until 1886. From 1985 to 1989, Zapopan Mining carried out an alluvial/elluvial mining operation reporting a production of 10,104 oz of gold (J. Shaw 2006). Dominion Mining Limited developed a small trial mining open pit in 1995. The Tally Ho lodes were discovered in late 2006 through a reverse circulation drilling program and follow-up diamond core program. The deposit was quickly expanded and brought into production, with mining occurring between 2007 and 2008 by GBS Gold.

The Tally Ho deposit is located just to the west of Fountain Head deposit and sits on the western limb of the Fountain Head anticline. The Tally Ho deposit strikes sub-parallel to the Fountain Head deposit and consists of two parallel mineralization zones striking SE-NW and plunging SE (local grid). The quartz veins are 1-20cm thick and host gold with a minor pyrite-arsenopyrite association. By 2007 a strike length of 160m had been defined with an average lode width of 18m and an average depth of 60m from surface (Z. Bajwah 2007a)

Diamond drilling carried out during 2008 at Tally Ho deposit failed to intersect significant gold grades despite recovering numerous pieces of visible gold in the core. Further exploration may be warranted to test the lode extension at deeper levels. Figure 7-32 graphically depicts the mineralized zones of the deposits and the potential of the Fountain Head area. Indications are that there are high-grade north plunging shoots beneath the floor of the Tally Ho open pit. There is very little deeper drilling and minimal drilling along strike.

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7.5.5      NORTH POINT - PRINCESS LOUISE GOLD DEPOSITS

The Yam Creek area was historically one of the better-known bedrock and alluvial gold mining areas in the Northern Territory following the discovery of gold in the area in 1872. The district was famous for its gold nuggets, the largest being 700oz (22.5 kilograms). By 1901, a three-compartment shaft had been sunk at North Point deposit with two crosscuts driven west at 42m and 62m as a prospecting exercise. The lodes in the 62m cross-cut were reported to average 5.0g/t Au over a width of 20m.

Further south at Princess Louise, 2,422t grading 51g/t Au was reported as having been produced in 1891. Gold was recovered from east-dipping (50°) quartz-sulphide veins within a west-dipping greywacke unit, 4m in thickness. The shoots were reported to plunge northerly at 30°.

The host rocks comprise cyclic silt-greywacke-mudstone sediments of the South Alligator Group (Lower Mount Bonnie Formation). These rocks are overlain by Finniss River Group, comprised of greywacke (flysch) sediments of the Burrell Creek Formation. The underlying Gerowie Tuff and local sills of Zamu Dolerite are exposed in the south of the area in the core of the fold. The dominant mineralized structural feature in the area comprises the west limb of the Yam Creek anticline that dips west at 50-60°. The east limb is steep to overturned and the axis plunges north at 10-30°.

7.5.6      BROCKS CREEK GOLD MINE

The Brocks Creek underground mine (historically known as the Zapopan Mine and incorporating the historical Faded Lily & Alligator Zones) is a stratiform, meta-sedimentary hosted quartz-vein type gold deposit, formed in the steeply south-easterly plunging hinge of a tight anticline known as the Brocks Creek-Zapopan (BCZ) Anticline. The mineralized sequence consists of argillite, often highly carbonaceous near its base, with variable proportions of inter-bedded greywacke, chert and tuff. There are thin BIF beds near the top of the sequence (G. K. Miller 1998).

The Brocks Creek area was intermittently mined from the 1870s to 1935 with modern exploration occurring since the 1970’s. Alluvial mining operations occurred in the Faded Lily area during the early 1990’s followed by mining of open pits being along the BCZ anticlinal structure. In 2003 underground development commenced at Zapopan, since this time underground mining has been intermittent but was completed by Crocodile Gold in 2012 when the mine was placed on care and maintenance and the pit was allowed to flood.

The BCZ Anticline is a tight asymmetric anticline, which plunges southeast at roughly 35°. The southern limb dips south at about 55° and the north limb is sub vertical. The BCZ Anticline has been subject to axial planar failure and thrust movements during the many deformation phases of the PCO. The deposits are proximal to the Brocks Creek Shear Zone (BKSZ), a locally significant approximately E-W trending shear zone located 3km south of the Burnside Granite. The BKSZ hosts several deposits in the Brocks Creek area and is thought to be a major mineralizing control (G. K. Miller 1998). The Burnside Granite appears to have had a significant influence on the local structural regime; whereas the regional structural trend is north-south, local structures are WNW, tangential to the intrusive body (Dunn 1998).

The axial plane of the BCZ Anticline is the focus for gold deposition. Gold mineralization at Zapopan can be either bedding concordant quartz veins 1.5 -2.5m 126 thick or stratiform sulphidic chert bands up to 0.4m thick (G. K. Miller 1998). At the Faded Lily deposit, gold mineralization occurs within a number of bands of quartz as well as some bedding concordant quartz veins, along vein margins and within graphitic shears and has a close affinity with pyrite and arsenopyrite. Mineralization zones may have up to 10% pyrite and 5% arsenopyrite and small grains of visible gold are a relatively common feature of higher grade zones. In the axial zone, where the concordant veins flatten, higher grades and thicker lodes occur ( (J. Shaw 2005)).

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Four lode zones have been identified:

Lode Type Thickness Grade
5 Lode Iron rich carbonate 10 - 20cm 15g/t Au
10 Lode Quartz Reef 0.2 - 2.0m ~50g/t Au
20 Lode Iron rich carbonate 1.2 – 3.0m ~60g/t Au
30 Lode Discontinuous veins 0.5m ~30g/t Au

TABLE 7-1 BROCKS CREEK LODE TYPES AND GRADES

The mineralization body is complicated by two faults named the “North Slide Fault” and “South Slide Fault”. The two “slide faults” are located about 25-40m apart and are post mineralization shear zones comprising incompetent foliated and graphitic rocks slightly oblique to the axial plane. The intersection of the slides with the axial fold plane splits the mineralization into three identifiable units, these being (from south to north) the Fissure Lode, Main Lode and Central Lode. The Main and Fissure Lodes strike east west and dip southwards at approximately -55° to -60°, in parallel with the bedding direction. The Central Lode is located along the hinge zone of the anticline. The slides are structural margins to the lodes, defining the up dip limit of mineralization in most cases. Lode thickness and gold grade decrease down dip from the lode-slide contact (Dunn 1998)). These slide faults trend 115° and dip steeply at about 80°, movement on these structures is apparently sinistral, but principally reverse dip slip.

The Rising Tide deposit is located over 2.5km north of the Faded Lily deposit and is outside the BCZ Anticlinal Zone. Unlike the Zapopan and the Faded Lily deposits, the mineralization is thrust fault controlled. At Rising Tide, at least two mineralized structures occur, these structures comprise shallow, 25° southeast dipping reverse fault planes within carbonaceous sediments of the Koolpin Formation that parallel the underlying contact with a Zamu Dolerite sill (J. Shaw 2005). The Koolpin sediment/Zamu Dolerite contact is sheared with some mineralization occurring within this shear. Koolpin Fm host rocks comprise argillite, carbonaceous and pyritic/pyrrhotitic shale, chert bands, calc-silicates and possible iron formation. A prominent late stage, crosscutting quartz vein on 330° cuts the deposit and passes into the Burnside Granite to the north.

7.5.7      GOODALL MINE AREA

The Goodall Mine area is hosted by a sequence of turbidites that published reports indicate belong to the Burrell Creek Formation. These sedimentary rocks are folded about a north trending F1 fold axes. It is part of the extensive Howley anticline.

Newmarket Gold interprets the mine area to be hosted by Mt Bonnie Formation, as the radiometric signature is distinctly anomalous in potassium, as it is in the Margret syncline located to the southeast.

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The turbidite sequence consists of a typical greywackes and siltstones that have undergone greenschist facies metamorphism that has produced a quarts-chlorite assemblage. The mineralized zone is associated with a quartz-sericite-carbonate-K feldspar assemblage with minor tourmaline and apatite. Minor mafic dykes cut the turbidite sequence.

Within the mine area there are several folds with wavelengths of 50-100m that are related to the north trending Howley anticline. They plunge to the north at between 15° and 60°.

Mineralization of the A Pod occurs on the eastern limb of an anticline with dimensions of up to 50m wide, 800m along strike and 140m in depth. Gold is associated with thin vein arrays (5-50 mm) of quartz and sulphides, which make up about 5-20% of the rock. Mineralization appears to diminish with depth.

A more complete description can be found in Quick, D.R., 1994 (Quick 1994).

7.5.8      WOOLWONGA MINE

At the past producing Woolwonga Mine the mineralization is structurally controlled and occurs in quartz veins associated with faults, shears and zones of brecciation within a moderately tight anticline striking 310° and plunging 35–40° southeast. A shear zone (the regional Pine Creek Shear) striking 330° cuts the anticline. Quartz veining comprises saddle reefs, sub-vertical veins, stockworks, associated with shear zones, and sub-vertical veins parallel to the axial plane of the anticline and to the dominant cleavage.

The host rocks consist of tuffaceous greywacke, mudstone and carbonaceous mudstone of the Mount Bonnie Formation, frequently in upward-fining turbiditic successions. The rocks are black in color due to a high carbon content.

The richest gold mineralization occurs at the intersection of the 330°-striking shear zone with the anticlinal axis, and at the brecciated margins of quartz saddle reefs. The predominant mineralization mineral association is arsenopyrite and pyrite, with subordinate amounts of marcasite, galena, native bismuth, pyrrhotite, chalcopyrite, sphalerite, covellite and chalcocite. Gold occurs as small particles of free metal in quartz or as minute blebs in arsenopyrite. Gangue minerals are mostly quartz with minor siderite, K-feldspar and Mg-rich tourmaline (dravite). (Ahmad, Wygralak and Ferenczi 2009).

7.6 MINERALIZATION

7.6.1      COSMO MINE

The Cosmo Mine mineralization lies within a marine siltstone package located between the Inner Zamu Dolerite sill and a +30m thick pyritic carbonaceous mudstone unit identified as the “Pmc” unit. Siltstones, near the Pmc contact often contain boudinaged chert lenses. These cherts are recrystallized to resemble the sucrosic texture of quartzite. The unit intercalates with massive and banded siltstones. The width of the gold hosting siltstones is 30 to 50m in the footwall of the F1 Fault and from several meters to 50+ meters in the hangingwall due to variably developed folding.

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Four main lodes have been delineated in the Footwall Lodes and three in the Hangingwall Lodes in relation to the F1 Fault. These are the 100 Lode, 200 Lode, 300 Lode and the 400 Lode on the footwall of the Eastern Limb, with the 500 Lode, 600 Lode, and 101 Lode in the hangingwall.

These gold bearing lodes are remarkably planar in the footwall of the F1 Fault along the long Eastern Limb of the fold. The hangingwall lodes are more complex due to parasitic folding of which many are isoclinal, causing localized thickening and shortening. Each lode is correlated by grade within its stratigraphic position in the mineralization bearing siltstones.

Footwall Lodes (Eastern Lodes)

100 Lode – This lode is constrained between the contact of the Graphitic Mudstone (Pmc) and the F10 unit, which is highly sheared on the footwall side of the F1 Fault. The thickness of the 100 Lode ranges between 5m up to 8m true width. The 100 Lode contains, near its center, a thin internal Graphitic Mudstone unit approximately 10-30cm thick, which is often un-mineralized. Grades are easily correlateable in plan and section. The 100 Lode mineralization can be traced with confidence with the current drilling down to the 480mRL, giving the mineralization a vertical extent of 670m. In the footwall the 100, 200, 300 and 400 Lodes are split and offset by a northwest/southeast dextral fault called the F9 Fault. The displacement along this fault is strike slip movement with approximately a 20-30 meter offset.

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200 Lode – This is the first mineralization that occurs west of the F10 faulted unit. Grades are usually more erratic and lower grade than in the 100 Lode, but it is still clearly correlateable through changes in lithology. The thickness of the 200 Lode ranges between 5m up to 10m true width.

300 Lode – This is the next lode to the west of the 200 Lode. This lode is usually low grade with variable and indistinct grade contacts although at depth and in the southern extent of this lode grades have been demonstrated to improve. The mineralization appears to correlate with coarser grained recrystallized siltstone units and is parallel to the bedding. The thickness of the 300 Lode ranges between 1m up to 3m true width.

400 Lode – This is the inner most lode. It is located close to the dolerite and is consistently low grade and like the 300 Lode it has indistinct grade boundaries. The thickness of the 400 Lode is very similar to the 300 Lode with ranges between 1m up to 3m true width.

Collectively the 100 to 400 Lodes are referred to as the “Eastern Lodes” (Figure 7-34).

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Western Lodes

Following a review of past drilling in context of new geological knowledge gained from underground mining and drilling in 2014 it was proposed to test the western Cosmo Anticline limb near significant mineralization in surface hole CP003W1. Gold mineralization in that hole was close to the Pmc – greywacke contact analogous to areas being mined on the Eastern Lodes.

Drilling in 2015 of the Western Lodes focused on the prospective NW shoulder to the Cosmo Anticline and generated further significant gold intersections (see (J. Miller 2014) and Section 9). Although this mineralization is hosted in very similar host rocks and stratigraphic location the results did not support a structural style such as the very linear zones within the Eastern Lodes.

Late in the year structural studies were conducted, which suggested the Western Lode mineralization occurs as a series of high grade shoots, which plunge moderately north approximately parallel to the Zamu Dolerite contact Millar (J. Miller 2015), however, appreciated that the Western Lodes may be a westerly continuation of the Hinge Zone. This area remains to be drill tested in 2016.

Hangingwall Lodes:

500 Lode – This is actually a continuation of the 100 Lode as it wraps around the fold hinge becoming part of the Western Lodes. Like the 100 Lode it occurs nearest to the Graphitic Mudstone and is bounded by the F10 mudstone unit.

600 Lode – This is actually the 200 Lode and it is interpreted to wrap around the fold hinge and become part of the Western Lodes mineralization.

On the western limb of the Hinge Zone in the hangingwall, the 500 Lode and 600 Lode are offset by the F9 Fault with displacement of approximately 15-20m. Collectively the 500 and 600 Lodes are referred as part of the “Hinge Zone”.

101 Lode – Termed the “Sliver”, this lode is a subsidiary fold on the Eastern Hangingwall limb of the fold. The lower extents of this mineralization were drilled to a scoping level in 2015 and have become a high priority for mineral resource definition and exploration in 2016.

Inter Lode Zones

Within the 100 Lode there is often an intermittent graphitic phyllite (the 11-Unit) ranging from 0.1m up to 3m thick in some areas. This unit is usually un-mineralized and is highly friable in some areas, particularly on the eastern footwall where high strain along the long limb is occurring. This unit is referred to as the 100 Lode Internal Mudstone. Separating the 100 Lode from the 200 Lode is another distinct 1-1.5m wide carbonaceous mudstone unit (the 10-Unit). This unit is a stratigraphic marker that can usually be traced from the footwall of the Eastern Limb, around the fold hinge on the hangingwall and around to the Western Limb. In the Eastern Lodes, this unit is faulted and is termed the 10 Fault. While the 10 Fault is predominately a bedding-plane fault in the footwall of the Eastern Limb, in the hangingwall the 10-Unit is not faulted.

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Gold mineralization is closely associated with arsenopyrite, often seen within the Pgtb unit, especially in the 100 Lode. The mineralization styles, both on the hangingwall and footwall of the F1 Fault are very similar, with the main mineralization associated with, but not necessarily totally constrained within the “Pgtb” unit. The main sulfide minerals in the fresh rock are pyrite and arsenopyrite, with traces of sphalerite and chalcopyrite. Pyrrhotite occurs below depths of 300m and is predominantly seen in the Pmc unit.

Higher grade gold mineralization is commonly hosted within Pgtb in the 100 Lode in proximity to the Pmc contact. Mineralization hosting lower gold grades can occur in the lodes within the non-graphitic mudstones. Minor mineralization is associated with quartz veins, breccias and shears. The footwall lodes are limb parallel, stratigraphically constrained planar mineralization bodies with relatively continuous mineralization. While the hangingwall lodes are stratigraphically associated with the “Pgtb” units, they are parasitically folded along the fold hinge. As a result the mineralization lodes are less continuous than in the footwall. Many of these parasitic folds are isoclinal. Geological mapping and assaying has shown shortening along the fold limbs and thickening in the folds.

Gold mineralization occurs along areas of high strain, such as the Pgtb along with constrictions in the mineralization hosting siltstones, and faults. Essentially, gold occurs in various styles of structural traps. The graphitic mudstones appear to be impermeable units that forces mesothermal fluids through the mineralization hosting siltstone package. It is believed that gold deposition may be caused by the redox front between the Carbonaceous Mudstone (Pmc) and the adjacent coarser grained siltstone units (Pgtb). The highest endowment of gold mined to date occurs at the interface between graphitic mudstone (reduced conditions) and hematitic alteration (oxidized conditions of the siltstones forming a redox boundary).

 

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7.6.2      UNION REEFS

From (Karpeta 2011)

Gold mineralization at Union Reefs and other gold mines in the area occurs in four end member styles, summarized as follows;

1.

Lode style veins: up to 4m thick, discontinuous, with variable but locally high gold grades. Lode style veins are pod-shaped and occur in highly sheared, dominantly shale host rocks.

   
2.

Stockwork vein systems: complex en echelon vein swarms largely restricted to greywacke-dominated horizons. Tend to form around fault & shear zone intersections, bedding/joint intersections, and anticlinal hinges. Typically of moderate gold grade.

   
3.

Sheeted vein systems: sub-parallel, laminated crack-fill vein sets that tend to occur in thinly inter- bedded greywacke-shale sequences. Typically of low gold grade.

   
4.

Saddle reef association (not seen at Union Reefs): in some deposits, gold is associated with the margins of folded quartz veins in plunging anticlinal hinges (“saddle reefs”). Studies have shown that the quartz reefs themselves are barren and pre-date gold mineralization, but their fractured and sheared margins have served as traps for gold mineralization during subsequent deformation and hydrothermal activity. Typically moderate to high grade with short strike length, and may have significant plunge extent.

Mineralization at Union Reefs is predominantly of the stockwork and sheeted-vein type, with lode-style veins comprising a lesser proportion of the deposits. Photos and descriptions of the Crosscourse E-Lens suggest that it is composed mainly of stockwork-type veining in a greywacke host and that elevated gold grades within the mineralization shoot occur due to the overlap of multiple generations of gold-bearing veins. The steeply plunging aspect of the E-Lens suggests that mineralization shoot location and morphology is strongly controlled by structural intersections.

7.6.3      PINE CREEK

At Pine Creek mineralization occurs in a zone along the anticlinal axis of the Enterprise Anticline. It occurs in saddle reefs, veins and in the wall rock of the Enterprise Anticline. The highest grades were observed in the saddle structures of the deposit and range from about 5-15g/t. Au Veins in the mineralized zone returned grades between 2 and 3g/t Au and in the wall rock of the mineralized zone grades range from 1-2g/t Au.

Up to 50% of the gold occurred as free gold, while the rest was bound up in sulphides. Sulphides within the deposit include pyrite, pyrrhotite, marcasite, arsenopyrite chalcopyrite, galena, sphalerite, bismuthinite, tetrahedrite and covellite. Gold occurs as inclusions 2-30µm across primarily within arsenopyrite and also in pyrrhotite and bismuthinite.

Three categories of veining have been recognized. The following is a simplification recognized by Dann and Delaney (1984) (Dann 1984). These are:

  1.

Saddle, spur and hinge zone veins. Forming the saddle structures present in the deposit.

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  2.

Ladder and sheeted veins are regularly sized and spaced veins occurring only on the western limb of the anticline.

     
  3.

Late stage veins and breccias, which are rich in sphalerite and galena and occur mostly in the Eastern Fault Zone.

Gold mineralization is strongest in the saddle reef structures.

7.6.4      BURNSIDE

7.6.4.1  Howley Group Deposits

The Howley Group includes the largest deposits in the Burnside deposit area. All deposits are proximal to the Howley anticline or parasitic anticlinal structures. Several plunge reversals occur along the fold axis, resulting in a range of host stratigraphy with the Middle Koolpin Fm, Gerowie Tuff Fm and Lower Mt Bonnie Fm along with bedding concordant sills of Zamu Dolerite, subsequently, several different mineralization styles occur. These can be broadly classified by two end members (A. K. Sener 2004):

The carbonaceous and iron rich units in the mid to lower Koolpin Formation host the vein stratabound type deposits of Cosmo with exploration potential at the Bon’s Rush deposit.

In the lower stratigraphy, where the Zamu Dolerite is present, gold is hosted in fractures, which formed in the brittle dolerite during the D2 folding event and were later filled with auriferous fluids. The fracturing is most intense in the Hinge Zone resulting in the saddle reef geometry of the mineralization. Mineralization also occurs in quartz shears along the dolerite/sediment contact. Dolerite hosted mineralization occurs at South Howley Extension and Bon’s Rush.

Dolerite hosted gold lode thickness can range from 2-10m (thickening of the dolerite unit due to duplex thrusts is evident at Howley South extension). Gold occurs in extensional quartz-carbonate veins with disseminated pyrrhotite, pyrite and chalcopyrite, widespread chlorite alteration and localized sericite alteration occurs around mineralized veins. Visible gold is often present.

Axial parallel shears occurred during the intense D2 folding which created duplex thrusts along the limbs of the Howley anticline. Lodes are sub-parallel and roughly concordant with the bedding along a 900m strike length of the tightly folded Howley anticline at the Chinese-Howley pits and at Mottrams. At Kazi, which is located on a parasitic fold to the east of the Howley anticline, a mineralized strike length of 200m has been outlined.

The majority of gold mineralization is strongly associated with an en-echelon quartz-sulphide vein system, gold mineralization also occurs in conjugate vein arrays and stockworks. Gold occurs as <40 micron sized grains in quartz-sulphide veins containing sulphides pyrite, arsenopyrite, chalcopyrite and sphalerite. Gold mineralization is strongly associated with arsenopyrite and pyrite (A. K. Sener 2004). Chlorite-sericite alteration associated with mineralized zones is common.

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At Western Arm and Big Howley mineralization is hosted in quartz-sulphide stockworks conformal to bedding along the anticline limbs and as “saddle reefs” in the anticlinal hinge zone. These saddle reefs occur between units of contrasting competencies; the upper mudstone of the Gerowie Tuff Formation and lower greywacke unit of the Mt Bonnie Formation at Western Arm deposit and inter-bedded shale and siltstones of the upper Gerowie Tuff Formation at Big Howley deposit.

At Western Arm deposit, gold is mostly fine grained (<40 microns) but some grains observed are as large as 200 microns (Zerovitch 1994). Gold mineralization is associated with a quartz vein stockwork zone containing silicified wall rock. Most of the gold occurs in late cross cutting fractures, spider veinlets, veins of K-feldspar, veins of massive pyrite and at the contact between quartz and carbonaceous mudstone (Hardy and Hague 2001b). Sulphides present include (in decreasing order of abundance) pyrite, chalcopyrite, arsenopyrite, pyrrhotite, sphalerite and galena. High gold values are strongly associated with arsenopyrite. Visible free gold occurs in late fractures, spider veinlets, and near contact between carbonaceous mudstone and quartz veins.

7.6.4.2  Hayes Creek Deposits

The Hayes Creek Group deposits are located in the southeastern portion of the Burnside area. Currently the most significant deposits of the Hayes Creek Group comprise the Fountain Head, Tally Ho, Yam Creek, North Point and Princess Louise deposits.

All these deposits contain gold mineralization within sulphide disseminations and quartz-sulphide veins of similar mineralogy. Sulphide is predominantly pyrite with minor arsenopyrite and trace amounts of chalcopyrite, pyrrhotite and galena. Some siliceous alteration and disseminated euhedral arsenopyrite occurs in country rock proximal to the mineralized quartz veining. Gold can be quite coarse and is often observed in drill core within quartz veins and along quartz vein selvages (Stephens 2000).

The orientation and style of quartz veins is mostly controlled by the competency of the host rock. In the finer grained lithology, veining is sub-vertical and appears to be axial planar, in the more massive brittle greywackes the veins are stockworked. Lithological contrasts between siltstone-mudstone packages and massive greywackes have been a further focusing for auriferous quartz veining (Z. Bajwah 2007b). These host rock properties have resulted in subtle differences between deposits.

Brocks Creek – Zapopan Group Deposits

The Brocks Creek – Zapopan Group deposits occur along the northwest trending Brocks Creek – Zapopan anticline, which is centrally located within the Burnside area. The Brocks Creek – Zapopan anticline is cut at a low angle by the Brocks Creek shear zone and lies within the thermal aureole of the Burnside Granite. Currently the most significant deposits of the Brocks Creek – Zapopan Group comprise the Brocks Creek/Zapopan underground and Rising Tide deposits.

Along the BCZ anticline, gold mineralization occurs mostly within bedding concordant quartz-sulphide veins. Gold is most closely associated with pyrite and arsenopyrite. Visible gold is common, mostly in the 40-60µm range (G. K. Miller 1998). Trace amounts of chalcopyrite, sphalerite and galena occur in later stage carbonate veins.

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The best mineralization at Brocks Creek occurs where quartz veins have formed during a number of brecciating events (Dunn 1998). This is evident with higher grade occurring in main lode, which has undergone at least two phases of brecciation (F. H. Muller 2007). Brecciation events appear to preferentially occur in a horizon of interbedded fine sandy layers about 1-5cm thick. Individual sulphide grains commonly show signs of multiple stages of brecciation. Arsenopyrite bands are common, ranging up to 10cm thick, comprising aggregates of coarse subhedral to euhedral arsenopyrite. In the absence of quartz veining these bands occur in the fine sandy beds proximal to main lode and often include fine pyrite laminae.

Mineralization at Rising Tide is predominantly stratabound within pyrrhotitic, carbonaceous sedimentary units of the lower Koolpin Fm, adjacent to Zamu Dolerite intrusives and in close proximity to the southern margin of the Burnside Granite (F. H. Muller 2006). The fault hosted gold mineralization is associated with quartz-pyrite-pyrrhotite-arsenopyrite vein sets (G. K. Miller 1998). Gold is thought to be supergene enriched and associated with structures leading to the Burnside Granite.

Fountain Head/Tally Ho

The Fountain Head deposit was discovered in 1883 and was subject to intensive elluvial mining activity until 1886. From 1985 to 1989, Zapopan carried out an alluvial/elluvial mining operation reporting a production of 10,104oz of gold (J. Shaw 2006). A small trial mining open pit was developed in 1995 by Dominion Mining Limited. The Tally Ho lodes were discovered in late 2006 through an initial reverse circulation drilling program and follow-up diamond core program. The deposit was quickly expanded and brought into production.

Gold mineralization in the Fountain Head deposit is hosted by siltstone, mudstone and greywacke packages of the Burrell Creek Formation of the Finniss River Group and is associated with quartz-pyrite-arsenopyrite veins (Figure 7-4). Mineralization is hosted in the Fountain Head anticline, which is interpreted to be a parasitic fold of the Margaret syncline. The Fountain Head anticline plunges gently to the SE, is asymmetric and has a tightly closed fold. The limbs dip at steep angles; the NE limb dips at 70° and the SW limb dips 50°-65°. The mineralization sits within the hinge zone and is structurally related to a NW striking fault system. Fountain Head mineralization is focused at the culmination of a doubly plunging domal structure along the axial trend (Z. Bajwah 2007a).

Mineralization is hosted by sub-vertical shear related stockworks, fracture zones in greywackes and saddle reefs at lithological contacts over a strike length of 420m. Most of the mineral resource is in the hinge zone of the anticline with gold grade rapidly tapering off down dip on the limbs. Fracture zones within the hinge zone lie parallel to the axis of the fold and have acted as a focus for fluid channelling.

The Tally Ho deposit is located just to the west of Fountain Head (Figure 7-31) and sits on the western limb of the Fountain Head Anticline. The Tally Ho deposit strikes sub-parallel to the Fountain Head deposit and consists of two parallel mineralization zones striking SE-NW and plunging SE (local grid). The quartz veins are 1-20cm thick and host gold with a minor pyrite-arsenopyrite association. By 2007 a strike length of 160m had been defined with an average lode width of 18m and an average depth of 60m from surface (Z. Bajwah 2007a).

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Diamond drilling carried out during 2008 at Tally Ho failed to intersect significant gold grades despite recovering numerous pieces of visible gold in the core. Further exploration is warranted to test the lode extension at deeper levels. Figure 7-32 graphically depicts the mineralized zones of the deposits and the potential of the Fountain Head area. Indications are that there are high-grade north plunging shoots beneath the floor of the Tally Ho open pit. There is very little deeper drilling and minimal drilling along strike.

North Point/Princess Louise

The Yam Creek area was historically one of the better-known bedrock and alluvial gold mining areas in the Northern Territory following the discovery of gold in the area in 1872. The district was famous for its gold nuggets, the largest being 700oz (22.5 kilograms). By 1901, a three compartment shaft had been sunk at North Point with two cross-cuts driven west at 42m and 62m as a prospecting exercise. The lodes in the 62m cross-cut were reported to average 5g/t Au over a width of 20m.

Further south at Princess Louise, 2,422t grading 51g/t Au was reported as having been produced in 1891. Gold was recovered from east-dipping (50°) quartz-sulphide veins within a west-dipping greywacke unit, 4m in thickness (Figure 7-43). The shoots were reported to plunge northerly at 30°.

The host rocks comprise cyclic silt-greywacke-mudstone sediments of the South Alligator Group (Lower Mount Bonnie Formation). These are overlain by Finniss River Group, comprised of greywacke (flysch) sediments of the Burrell Creek Formation. The underlying Gerowie Tuff and local sills of Zamu Dolerite are exposed in the south of the area in the core of the fold. The dominant mineralized structural feature in the area comprises the west limb of the Yam Creek anticline that dips west at 50-60°. The east limb is steep to overturned and the axis plunges north at 10-30°.

Auriferous quartz-sulphide veining occupies structurally prepared fault-fold sites on splays from the regionally important Hayes Creek Fault that trends NE through the area. Mineralized quartz-sulphide veins are associated with greywacke-dominated packages within the west limb and axial zones of the Yam Creek fold, particularly where bedding slip, reverse faults and splays cut the limb at shallow angles (Z. Bajwah 2007b). A number of NE trending faults displace bedding trends in the Yam Creek area. The faults are thought to be sub vertical and appear to post-date mineralization.

At the North Point deposit mineralization occurs as linear, northerly oriented, multiple lode system that dip conformably with bedding at ~45° west (Figure 7-42). The system is reasonably continuous along strike for about 300m and down dip for 60m (Ahmad, Wygralak and Ferenczi 2009).

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Paqualin Group of Tenements

In the Paqualin area, the sequence has been folded into south-plunging anticlinal structure where western limb has been affected by a NW-trending fault. The Zamu Dolerite occupies the hinge of the fold and appears to have been interlayered within the Koolpin Formation and Gerowie Tuff – an artifact of deformation.

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Bon’s Rush is associated with quartz-carbonate veins that dip shallowly (~25-30°) to the northeast. The mineralized zone is hosted by a carbonated, sulphidised, sericitised and occasionally silicified granophyric phase of the Zamu Dolerite. The higher grades are associated with a zone of quartz veining, chloritization, pyrite, arsenopyrite and minor pyrrhotite, hosted within a shear zone in the hangingwall of the Upper Zamu Dolerite Sill.

Outcrop within the Rhodes Group of tenements (including the Kazi deposit) is very poor due to extensive black soil and creek alluvium deposited by Howley Creek immediately to the south. Interpretation from regional mapping, and supported by drilling, shows that the Gerowie Tuff, of the middle South Alligator Group, underlies much of the area. This unit has been intruded by mafic Zamu sills and folded into north trending structures.

Gold mineralization at the Kazi deposit has been interpreted to occur in four parallel lodes within the interbedded metasedimentary units of the Koolpin Formation. Each lode is divided by the reverse faults to create enechelon zones within each lode. Lodes 100, 200, 300 & 400 (from surface down) are subparallel to the Zamu Dolerite intrusion and strike ENE with a southerly dip ranging 200 to 350. The lodes range from 2m to 6m vertical thickness. Minor mineralization extends into the Zamu Dolerite, mainly represented by lode 400. A high-grade zone exists between 10,000N and 10,130N and is coincident with a quartz-pyrite rich, sheared fault zone interpreted to thrust the Koolpin to the north, over the Zamu Dolerite (Harris 2005).

The Western Arm mineralization is hosted by a folded and sheared sedimentary sequence of Lower Proterozoic age. Rock types included in the deposit are siltstones, greywacke and mudstone with some tuffaceouse members and variable diagenetic pyrite. The rocks have been folded along the north-south axes and high strain zones occur on the limbs of the folds and upon parasitic fold structures sub-parallel to the fold axes. The high strain zones have been channelways for gold bearing fluids, which have introduced quartz, pyrite and arsenopyrite into fractures. Potassic alteration reflected in pine feldspathic veinlets and patches attended the mineralizing event.

The gold distribution is controlled by steep east and west dipping fracture sets with east dipping sets dominating. Bedding and lithology exerts some control on gold distribution. Mudstone-siltstone packages are favored hosts while massive greywackes and tuffaceous rocks are less so (Shaw, 1993).

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7.7 DEPOSIT DIMENSIONS

Deposit Lode Horizontal Length
(m)
Vertical Length
(m)
Horizontal Widths
(m)
Cosmo 100 750 700 7-20m
101 375 100 2-80m
200 300 650 6-14m
300 675 350 5-14m
400 300 625 5-15m
500 425 500 6-60m
600 250 375 3-17m
Howley Combined 3200 250 700
Mottrams Combined 670 120 60
North Point Combined 900 80 25
Princess Louise Combined 580 156 60
Rising Tide Combined 300 860 40
Fountain Head Combined 420 160 90
Tally Ho Combined 160 60 18
Kazi Combined 200   2-6
Western Arm Combined 1100 130 100
Bon's Rush Combined 650 100 50
Prospect Combined 650 450 70
Crosscourse E-Lens Combined 430 450 130
Crosscourse Western Lodes Combined 270 490 30
Esmeralda Zone A 1150 150 200
Zone B 570 150 115
Lady Alice Combined 270 135 22
Millars/Big Tree/Ping Que Combined 1100 130 60
Orinoco Combined 460 160 40
Union North Combined 1300 300 60
Union South/Temple Combined 800 220 40
Cox Combined 300 130 40
Czarina Combined 600 110 40
South Czarina Combined 660 135 60
Enterprise Combined 800 200 30
Gandy’s Combined 1280 115 60
Kohinoor Combined 1,000 120 60
International Combined 1300 120 90
South Enterprise Combined 400 230 20

TABLE 7-2 DEPOSIT DIMENSIONS

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8 DEPOSIT TYPES

The contribution from gold deposits in Proterozoic sedimentary basins to total gold production has increased markedly over the past two decades, both globally and within Proterozoic basins within Australia. Consequently, many Proterozoic basins are now considered high priority exploration targets.

8.1 MINERALIZATION DEPOSIT MODELS

A variety of genetic models, ranging from magmatic through hydrothermal to syngenetic, have been postulated in the past for the formation of gold deposits in the Pine Creek Geosyncline. Gold and base metal mineralization in the Pine Creek Geosyncline is commonly associated with granite intrusions and have often been classified as high temperature contact aureole deposits. A secondary host rock control has also been suggested due to the association of gold mineralization with carbonaceous metasedimentary rocks, such as at Cosmo Mine.

However, much of the gold mineralization occurred after the main intrusive event, the intrusion of the Cullen Batholith, and the relationship of gold mineralization and carbonaceous rocks is not the most important control on mineralization. More recently, authors have argued that gold mineralization is structurally controlled; occurring in brittle-ductile structures at the greenschist-amphibole facies boundary and hence has an epigenetic origin (Parrington and McNaughton 1997).

In places, such as the Cosmo-Howley area, duplex thrust folds with buckle folding or basin and dome structures appear to be more significantly mineralized. The presence of shear systems linking anticlines higher in the sequence also appears to have provided the ideal fluid focusing mechanisms to localize gold-bearing fluids.

Accepting that gold deposits of the Northern Territory have a structurally controlled mesothermal setting, then on the basis of host rock and mineral association they can be divided into seven types:

Over half of the gold occurrences are gold-quartz vein deposits.

Native gold is the main mineralization mineral and is commonly present as micron sized grains; coarse nuggets are rare. Gold is commonly associated with pyrite, arsenopyrite and pyrrhotite and in places with minor base metal sulphides. Quartz, chlorite, sericite and carbonates are the common gangue minerals in the gold-quartz deposits.

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All gold deposits in the Northern Territory show some structural control at the regional and deposit scales, with most deposits within the Pine Creek Orogen trending northwest-southeast.

Base metal mineralization in the Pine Creek Orogen strike significantly differently than the gold veins, suggesting different discrete mineralizing events. They are interpreted to be syngenetic.

Most deposits show a preference for competency contrast situations in dilatant or low pressure zones, such as anticlinal crests, recurrent shear zones and necking zones. Gold mineralization is invariably late, occurring after orogenic events.

Common factors for most gold deposits include:

Five main types of mineralization have previously been recognized within the Pine Creek Orogen. These include:

Most gold mineralization in the Pine Creek Orogen occurs within the South Alligator Group, especially above the Middle Koolpin Formation, and in the lower parts of the Burrell Creek Formation. At Maud Creek, gold mineralization is hosted by the Tollis Formation that represents the uppermost unit of the El Sherana Group and unconformably overlies the Burrell Creek Formation. Most of the fold-associated deposits were probably formed during intrusion of granitoids such as the synorogenic Cullen Batholith and the Burnside Granite.

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The most important regional-scale exploration vectors to the orogenic style of gold mineralization are:

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Dilatant Zone Mineralization
Compressive Zone Mineralization
Type: Fold Structure

“Telfer Type”

Fold - Brittle
Fracture
Thrust

Examples   • Cosmo • Fountain Head • Rising Tide
  • Brocks Creek • Goodall • Woolwonga • Kazi
  • Faded Lily • Howley Ridge    
         
Economics   • +5g/t Au • +2.5g/t Au • +2.5g/t Au
  • +10g/t Au • 100,000- 4,000,000oz • 20,000- • 60,000 - 1,500,000oz
  • 50-800,000oz • Open Pit, 1,000,000oz • Open Pit
  • Open Pit • u/g extensions • Open Pit • Small u/g
  • u/g extensions      
Geological • Anticline hosted • Anticline hosted • Anticline hosted • Reverse fault hosted
Features   • Stratabound • Stratabound • Discordant
  • Strataform • Dilational (Area B & • Compressive • Compressive (Area A
  • Dilational (Area C) (Area A & D) & D)
     B & C) • Fe-Carb stratigraphic • Stratigraphic • Stratigraphic
  • Fe-Carb association association with association with
     Stratigraphic • Greywackes, Fe-Carb Greywackes &
     association Fe/carbonate altered • Greywackes & siltstones
  • Greywackes & silts & graphitic silts siltstones • Often amphibolite
     Graphitic     facies alteration
     Siltstones      
Target        
Ranking #1 #2 #3 #3
  HIGH Priority High Priority O/P Moderate Priority Moderate Priority O/P
      O/P  
  • Small • Large tonnage   • Moderate tonnages
  • high grade • Elevated grade • Moderate • Moderate grades
      tonnages  
      • Elevated grades  

TABLE 8-1 PINE CREEK OROGEN MINERALIZATION MODELS

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8.2 STRUCTURAL MODELS

8.2.1      NORTHERN TERRITORY PROPERTY

Assuming that the majority of gold deposits within the Pine Creek Orogen are structurally controlled and mesothermal/orogenic (Groves 1998) in origin, it is likely that the known gold deposits are associated with regional shear zones and fault systems that were formed during orogenesis. By analyzing maps displaying total magnetic intensity (TMI) data, a number of continuous, NNW-tending first-order faults can be defined within the sedimentary-dominated rock sequences of the tenement area (Figure 8-2).

The majority of known gold deposits within the tenement area are spatially associated with the first-order, NNW-trending shear zones. It is therefore likely that these first-order shear zones acted as conduits for epigenetic gold-bearing fluids during/after orogenesis and they control the distribution of gold mineralization known in the tenement area. Additional factors such as the presence of the South Alligator Group, proximal antiformal hinges (e.g., Cosmo-Howley) or converging secondary shear zones (e.g. Crosscourse deposit) would also play an important role in localizing gold mineralization.

The major shear zones are separated by rock sequences that regularly preserve NNW-trending, doubly-plunging antiformal hinges with no clear evidence for strike-slip deformation along these NNW-trending structures. South of the Burnside Granite area, a series of NE-trending shear zones and faults have also been defined (Figure 8-2). Based on preserved asymmetries of rock sequences either side of these NE-trending faults, dextral-dominated strike-slip deformation possibly occurred along these relatively later structures.

Consulting geologist Paul Karpeta’s comments, which the Authors have reviewed and agree with, include the following:

The origin of the gold mineralization in the Pine Creek Belt is controversial. Matthai et al (S. H. Matthaei 1995a) (S. H. Matthaei 1995b)argue for an intrusive-related thermal aureole model associated with the Cullen Granite Batholith. However, Partington and McNaughton (1997) (Parrington and McNaughton 1997) and Sener et al (2003, 2005) (A. K. Sener 2003) (A. K. Sener 2005) prefer a structurally related model with gold mineralization being related to duplex-fold-thrust systems. This study indicates that gold mineralization in Howley, Brocks Creek and Union Reefs areas is structurally controlled and three structural styles of mineralization can be identified.

The first style of mineralization is found in the Koolpin Formation at Cosmo Howley and Rising Tide and is thrust related being found as bedding-concordant quartz veins and lodes associated with thrusts. The optimum location for such mineralization appears to be where the main thrust surface changes orientation possibly because of a change in lithology/competency or the buttressing effect of normal fault planes. Ore bodies will be oriented orthogonal to the (D1a) thrust direction and plunge in the direction of the local D3 cross fold limb.

The second style of mineralization is found in the Gerowie Tuff and Burrell Creek Formations at Howley, Faded Lily, Zapopan and Union Reefs and comprises gold mineralized quartz veins associated with large amplitude anticlinal folds. These large amplitude folds would be formed by buttressing on an inverted normal fault plane (Howley Structure or Pine Creek Fault) and be associated with thick competent sills of Zamu Dolerite or beds of greywacke inter-bedded with less-competent tuffs and meta-pelites. Such folds would propagate backwards away from the buttressing fault plane, the fold nearest the buttress being the oldest and largest, which would act as the optimum location for mineralization by fluids migrating up the fault plane. Ore bodies would form parallel to the D1b fold axis and plunge in the direction of the local D3 cross fold limb.

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The third style of mineralization is found on the Gerowie/Mount Bonnie contact at Mottrams and in the Burrell Creek Formation at Union Reefs. The third style is shear zone related where an inverted normal fault is reactivated as a strike slip fault and mineralizing fluids migrate up the fault zone. Optimum mineralization will be in areas of transtension (sinistral stepovers) with good competency contrasts and a good reductant (e.g. graphite – though the graphite may itself be a product of methane migration up the shear zone) that is where the shear has a releasing bend (e.g. Crosscourse). Ore bodies would form parallel to the local D2 fold axes, orthogonal to the shear direction with a plunge related to the local D3 cross fold orientation.

In all these styles, previously existing structures (rift-related normal and transfer faults) play an important role in localizing the mineralized structures.

Sedimentology and Volcanology

This evolution from subaerial through shallow marine to deep marine sediments accompanied by bimodal volcanism suggests of the opening of an intra-cratonic rift system accompanied by a marine transgression. The Koolpin meta-pelites and BIFs are thought to be shallow lagoonal sediments. The Gerowie Tuffs are thought to be the products of repeated subaqueous eruption of felsic magmas into shallow water similar to the Bergslagen area in Sweden. A gradual change to more mafic volcanism occurred towards the top contact with the overlying shallow marine Mount Bonnie Formation clastics. At this contact syn-volcanic VMS-style mineralization can be expected at favorable locations (Iron Blow). The Burrell Creek Formation comprises deep water turbidites with thick greywackes and slates.

Structural Geology

The structural evolution of the Pine Creek Belt in the Howley-Pine Creek area is thought to be as follows:

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This episode of compressive structural deformation is interpreted as a phase of basin inversion during which previously extensional structures were reactivated as compressional structures.

Mineralization

The Early Proterozoic tuffs in the Bergslagen area in Sweden provide a close analogue for the Gerowie Tuffs. In the Bergslagen area subaqueous acid volcanic tuffs are associated with stratiform polymetallic tuff associated deposits (Broken Hill Type), stratabound polymetallic limestone skarn deposits and distally, BIF-associated gold deposits. The REE-bearing skarns at Basnas are nearby.

Syn-Volcanic Mineralization: Iron Blow in the Pine Creek Belt probably represents polymetallic syn-volcanic mineralization i.e. a VMS-type deposit, which would explain the high silver-gold ratio. The deposit needs to examined in more detail and alteration and volcanic facies mapped.

Skarn Mineralization: The mineral assemblage (diopside-garnet-calcite) at Rising Tide suggests that it could be a skarn though the gold mineralization there is not thought to be skarn related but is probably much later. The presence of this skarn coupled with the anomalous REE content of the Burnside Granite suggests that Rising Tide may have enhanced REE mineralization.

Gold Mineralization: Three structural styles of gold mineralization have so far been identified:

All of these structures appear to be directly related to pre-existing structures and result from the buttressing effect of earlier rift-related normal faults.

The plunge of the ore bodies can be directly related to the plunge of lineations, fold axes and boudins.

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8.3 COSMO MINE MODELS

There have been significant advances to the structural understanding of the Cosmo-Howley area over the past five years since underground mining commenced at the Cosmo Mine. (Note; local Cosmo Mine Grid is used in this discussion).

The Cosmo deposit is hosted by a folded anticline with numerous limb parallel faults and major east west crosscutting faults, which have recently been concluded to roll across the Cosmo Anticline nose and splay into major limb parallel north-south faults. In addition, a series of northwest orientated faults and fault-splays are appreciated as locally important to gold mineralization along with internal accommodation structures, which are less well understood.

There are five known distinctive styles of faulting identified within the Cosmo Mine. The majority of these have been demonstrated to have an impact to the Cosmo East Limb and associated gold mineralization;

8.3.1      COSMO MINE EAST-WEST FAULTS

The most important east-west fault currently identified in the Cosmo System is the F1 Fault. The F1 Fault is a thrust fault with a southeast strike slip movement of the hangingwall. The total up thrust displacement is currently interpreted to be approximately one hundred meters, but the east-west displacement is tens of meters. The movement was estimated by correlating a distinct strike change in the lodes in the footwall and the hangingwall. The F1 Fault divides the Cosmo system into two domains, the Footwall Domain and the Hangingwall Domain. The highest grade mineralization on the 100 Lode plunges parallel to and within 120m of the footwall of the F1 Fault (see Figure 8-3). This fault postdates mineralization.

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Previous reports have described a second major east-west fault at Cosmo Mine, identified as the F3 Fault. This fault was interpreted from a single fault exposure in the southeast corner of the Phantom open pit and four diamond drill holes. Underground diamond drilling of the Inner Metasediments, targeting the Lantern gold mineralization do not support the interpretation of the F3 Fault.

8.3.2      COSMO MINE NORTHWEST STRIKING FAULTS

The F9 and F8 Faults are significant northwest structures, which crosscut the Eastern Limb of the Cosmo deposit (J. Miller 2014) (See Figure 8-4). The F9 Fault can be traced from the footwall of the Eastern Limb through and into the hangingwall, where it is displaced by approximately 100m by the F1 Fault. This fault can be observed on the western high wall of the pit as a sub-vertical fault with up to 30m of dextral strike slip movement. Both faults predate the F1 thrust fault and although being un-mineralized, empirical evidence shows a spatial relationship with higher gold grades in the adjacent Eastern Lodes (e.g. F8 Fault & 300 Lode). (J. Miller 2014).

A further northwest cross fault is suggested from drilling data to the north of F9 Fault, beyond the present Sliver mineral resource model. Further drilling in 2016 will aid in further identification of this newly proposed fault and the potential for similar elevations in gold grade and mineral resource thickness as within the Eastern Lodes mined to date.

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8.3.3      COSMO MINE NORTH-SOUTH STRIKING FAULTS

A large north-south striking fault has been identified on the Western Limb of the fold hinge. This fault is known as the F2 Fault. The F2 Fault occurs immediately to the west of the 500 Lode and can be traced south to the western side of the Cosmo and Phantom open pits, and also north to the 2200N section. This large fault has a relatively uniform strike and consists of 5-20m of highly graphitic gouge and finely crushed rock. It is modelled that the western extents of the F1, F8 and F9 Faults terminate into this fault. Gold mineralization appears to be almost completely confined to the east of this normal fault, which strongly suggests it to be a major control on gold deposition. The F2 Fault is found to flatten significantly to shallow depths such as modelled in the Phantom pit, but otherwise is well defined, by diamond drilling data, to dip about 60° to the west.

8.3.4      COSMO MINE LIMB-PARALLEL STRUCTURES

In the footwall of the Eastern Limb, there are a series of limb-parallel faults. The most continuous of these faults is the F10 Fault. This is a bedding-plane fault in the Eastern Limb of the deposit, which occupies the 10 Unit, a graphitic mudstone and a zone of weakness. The F10 Fault is associated with limb-parallel deformation associated with the high strain zone on the long limb of the fold and is usually about 0.5 -2.0m in width. This shale unit is graphitic and metallurgical studies have shown it to have preg-robbing characteristics (can lower metallurgical recovery).

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The F10 Fault separates the 100 Lode from the 200 Lode. It has minor and major discontinuous northwest splays associated with changes in orientation of the dolerite and sedimentary stratigraphy. Strong shearing, significant brecciation and milled gouge zones characterize the fault.

There is evidence that the F10 Fault intersects the base of the Cosmo pit. There are also a series of other faults that can be identified on surface at the 1100mRL. These faults at surface are reflected by deeply incised and eroded slots in the Cosmo pit wall (Figure 8-6); these faults are not seen as major structures in the underground and appear to be discontinuous.

There are small 0.2 -0.5m fault splays off the main F10 Fault in the footwall that can be identified and mapped underground. These do appear to offset mineralization on the 200 and 300 Lodes.

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8.3.5      COSMO MINE - FLAT TO SHALLOW-DIPPING FAULTS

Underground mapping has identified a series of shallowly dipping faults with the Zamu Dolerite especially exhibiting these faults. These are relatively tight faults that are accommodation deformation structure formed during folding and thrusting.

There are several other flat dipping faults that have been identified within the Cosmo deposit. The most obvious are the north shallow dipping reverse faults that can be seen in the eastern wall and the western wall of the pit. The displacement along these faults is in the order of 5-10m (See Figure 8-9).

Within the development of the Cosmo underground a major shallowly north 10° dipping, east-west striking fault has been identified on the 875mRL level. This fault appears to be a splay off the major F1 Fault that splits the Cosmo deposit into two domains the Hangingwall and Footwall. This fault has a thickness of 1.0 -1.5m. The strike extent is currently unknown but has been identified within the Zamu Dolerite in development drives and from drilling. This fault may extend into the interior siltstones.

8.4 UNION REEFS MODELS

The Union Reefs deposit model (including Esmeralda) generally conforms and supports the Pine Creek Orogen model as outlined in Section 8.1 . Gold mineralization has been focused within two zones, (Union and Lady Alice Line at Union Reefs and Zone “A“ and Zone “B“ at Esmeralda) in the sheared axial zones of two adjacent faulted antiforms that strike NNW-SSE. At Esmeralda the northeastern “Lens A” is within 300m of the contact of the Allamber Springs Granite of the Cullen Suite and lies within the outer metamorphic aureole of the granite. It dips steeply southwest and has been significantly silicified and brecciated. Chert facies rocks are reported to coincide with the mineralized zones, which locally contain visible gold.

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Gold is generally hosted within a series of NNW-SSE striking, bedding-plane parallel quartz and quartz-tourmaline veins, with associated pyrite-sericite alteration, in an alternating slate-greywacke sequence of the Mt Bonnie Fm.

Bedding in the Esmeralda area is dominated by steep dips (>70°) towards the NE and SW, related to the tight to isoclinal NW-SE striking folding.

  1.

Early bedding plane parallel SW-over-NE thrusting, which may locally have produced imbricate fans of fault bend folds where the thrusts stepped up through the stratigraphy (D1).

  2.

SW-NE oriented compression tightened the D1 folds and producing new folding such as second order parasitic folds on the SW limbs of the first order folds (D2).

  3.

Clockwise rotation of the shortening to W-E orientation resulting in sinistral strike-slip faulting often reactivating the now subvertical thrust planes (D3).

  4.

Further clockwise rotation of the shortening direction to NW-SE producing resulting in subtle cross folding and the doubly plunging axial planes of the D2 folds (D4).

  5.

Local reactivation of the strike slip faults as dextral faults (D5).

  6.

Intrusion of the Allamber Springs Lobe of the Cullen Batholith resulting in the uplift of the Mount Bonnie Formation and the formation of the late fracture cleavage (D6).

The model for the structural evolution of the Esmeralda area resembles that proposed for the Union Reefs area with the exception of the early bedding plane parallel thrusting D1, which is probably present at Union Reefs but was not observed.

8.5 PINE CREEK MODELS

Gold mineralization at Pine Creek is focused on the axial zones of parallel major upright folds. The most productive is termed the Enterprise anticline; others include the less productive International-Czarina anticline. The folds plunge shallowly towards 135 degrees at around 10 degrees and the limbs dip southwest and northeast at around 65 degrees. The fold axes are sub-vertical.

The south plunging Enterprise fold exposed a well-stratified succession of alternating mudstones, nodular cherty siltstones and greywackes that has been correlated in detail throughout the Pine Creek Gold Field. To the southeast of the Enterprise Pit, N-S faulting coincides with a kink in strike that imparts a more southerly strike and seems to offset the principal fold axes in a sinistral sense. The continued southerly fold plunge takes the Mt Bonnie sequence beneath the Burrell Creek Formation in MCN 523. These lithologies have been less gold-productive in the Pine Creek field but nevertheless host several historic gold workings at the south end of the field. (Cox’s, Battery Shear/Bashi Bazouk, Eleanor, Elsinore, Kohinoor and Jensens deposits).

8.6 GOLD - URANIUM MINERALIZATION

Significant gold mineralization is associated with world-class uranium deposits, particularly in the South Alligator River region of the eastern Pine Creek Inlier. Bonanza grade gold mineralization was reported while mining high grade pitchblende mineralization in the 1950s and 1960s but since the discovery of high-grade gold mineralization associated with the Jabiluka II mineralization body in the late 1970s there has been more attention paid to this style of mineralization.

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The Jabiluka uranium-gold deposit is located 20km north of Jabiru. In 1998, North Limited reported “mineral resources”1 (non NI43-101compliant) of 19.53Mt of mineralization grading 0.46% U3O8 and a gold mineral resource (non NI43-101) of 1.1Mt at 10.7g/t Au; while these numbers are not current, they are indicative of the gold grades present (Ahmad, Wygralak and Ferenczi 2009).

Gold has also been located at the Koongarra and Ranger uranium deposits and gold and platinum group elements have also been located at Coronation Hill (Nicholson 1990)

This mineralization is characteristically associated with the Early/Middle Proterozoic unconformity surface or with late orogenic acid volcanics and sediments deposited on that surface. There is a strong structural and stratigraphic control on the mineralization and an association with ferromagnesian metasomatic alteration (chloritization, magnetitization, hematitization).

In 2009-2010 Thundelarra Exploration reports that uranium mineralization at the Thunderball deposit in the Pine Creek area appears to be shear hosted and consists of massive veins and disseminated uraninite (pitchblende). The mineralized zone is in an anticline that plunges to the north at approximately 40° and remains open down plunge.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves.

8.7 POLYMETALLIC DEPOSITS

The know polymetallic deposits, which include the Iron Blow and Mount Bonnie deposits, contain inter-bedded pyritic shale, dolomitic siltstone and tuff of the Mount Bonnie Formation. They contain significant amounts of zinc, lead, copper, silver and gold.

The largest base metal mine within the Pine Creek Orogen was the Woodcutters Mine located approximately 47km southwest of the Tom’s Gully Mine. It produced 4.65Mt grading 12.3% Zn, 5.6% Pb, and 87g/t Ag from 1985-1999. It is hosted by a tight anticlinal fold within the Whites Formation (dolomitic shale). The anticline is faulted along its axis and intruded by lamprophyre dykes. Several thin, tabular, en echelon bodies of sulphides developed over a strike length of 1.4km. The deposit is considered epigenetic. There are possibilities it may be remobilized from an initial stratiform zone.

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9 EXPLORATION

9.1 COSMO EXPLORATION

The Cosmo Mine is currently the only operating mine for the Newmarket Gold within the Northern Territory Operations project. Ongoing geological studies and projects were undertaken in 2015 to better understand the deposit and generate predictive exploration models. Those studies, which have made substantial conclusions or prompted changed understandings, are summarized below (for additional information outside of the summary below, please review previous technical reports on the Cosmo deposit for example (Gillman, et al. 2009) and (Smith and Pridmore 2014).

Late in 2014 a visit to Cosmo Mine was made by Dr. John Miller from the Centre for Exploration Targeting (“CET”, University of Western Australia & Curtin University collaborative research unit) to define and document the controls on gold mineralization at Cosmo Deeps to generate near mine exploration targets (J. Miller 2014). A comprehensive report, presentation and new ideas were produced from Miller’s work resulting in four ‘in-mine’, and four ‘near-mine’, prioritized drill targets and recommendations to reprocess geophysical data and conduct additional targeted research projects around the mine.

Much of the exploration drilling and research conducted by Newmarket Gold in 2015 was a follow up of the initiatives suggested by Cosmo geological staff and Miller. Two return visits to the mine were made by John Miller to review the results of testing the targets he proposed. Cosmo exploration growth drill programs were conducted at (additional details can be found in Section 10.1);

Complimentary to the above exploration programs at Cosmo Mine was the mining of a drive at the 640RL level with purpose to provide optimal drill platforms to drill targets such as the Sliver, Hinge and Western Lodes to the deeper northern end of the underground mine.

Other exploration studies described in this section are;

A key aspect of Miller’s 2014 studies was the importance of fault orientation to the northwest faults, which splay out of the eastern Cosmo limb and across the fold nose where they become discordant to bedding at encounter the F1 Fault. Miller illustrated, what the mine geologists appreciated, that gold mineralization can be elevated adjacent to the NW cross-faults and where fold hinges and adjacent fold hinges in the greywacke units occur (J. Miller 2014) & (Figure 9-1)

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Figure 9-1 is showing the importance placed on NW cross faults as a control on gold mineralization. LHS shows 3 of the 4 targets proposed in late 2014 and drill tested in 2015; 1=Western Lodes, 3=Inner Metasediments inside the western Dolerite limb (later known as “Lantern”) and 2=Cosmo South.

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Figure 9-2 is a plan of the underground geology at the 920RL illustrating the bedding and fault strike of high-grade gold areas such as Hinge and fault splays (J. Miller 2014).

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Figure 9-4 illustrates the two most successful areas of new mineral resource potential;

  1.

The northern down-plunge continuation of the Sliver Lode with successful surface and underground drilling programs suggesting a shallower dip to the F1 Fault and a potential new NW- striking cross fault; and

  2.

The area of Inner Metasediments under the Zamu Dolerite and beneath the former producing Phantom pit. This area, known as the Lantern Target is the down-plunge continuation of mineralization mined in the Phantom open pit during the 1990’s. Circular gold intersections are 2m minimum composites using a 2.0g/t Au lower cut-off with minimum 2m internal dilution.


9.1.1      STRUCTURAL STUDIES AND 3D MODELING

With recognition that the host rocks and mineralization style within the first holes completed into the Lantern target were markedly different to that known in the Cosmo Mine mineralization, it was decided to have Dr. John Miller return to review the new holes, and to gain some petrographic understanding of the mineral assemblages found.

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9.1.1.1   Lantern Petrographic Study (Ashley, May 2015)

Paul Ashley from Petrographic and Geological Service of Armidale, NSW was contracted to prepare rock thin sections and petrographically describe three samples from hole CW101002;

  Sample A 15.63m – 15.79m
     
  Sample B 161.48m – 161.65m
     
  Sample C 186.36m – 186.54m

and one underground hand grab Sample D of Phyllite from between the 300 Lode and 400 Lode mineralization the Cosmo Deeps (exact location undefined).

Polished thin sections (PTS) were prepared from each sample at Petrographic International Pty Ltd in Brisbane. The purpose of the investigation was to ascertain rock types (including protolith material), metamorphism and subsequently imposed retrograde alteration with aim to characterize associations with gold mineralization. Summary descriptions and interpretive report were supplied to Newmarket Gold (P. Ashley 2015) and are summarized below.

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Some geologists are skeptical that supergene processes related to near-surface meteoric ground waters has caused the distinctive hematite and associated retrograde mineral assemblages across the inner metasediments. The diamond core from Lantern is more than 150m below surface and shows little evidence of fracture oxidation.

Further petrographic and other alteration characterization studies are warranted and planned for 2016.

9.1.1.2   Western Lodes Review (Miller, June 2015)

Dr. John Miller revisited Cosmo Mine site when the second phase of drilling into the Western Lodes target was nearing completion. Six holes with full results including assay information were available and a seventh in progress. A wireframe with the new drill results was generated and drill density gaps with gold potential identified. Recommendations from this review are;

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9.1.1.3Western Lodes Structural Study (Beeson, November 2015)

In October 2015, Dr. John Beeson from Geoscience Now Pty Ltd conducted a structural study and review of the Cosmo Western Lode mineralization. The study aimed to develop a more detailed structural interpretation of Western Lode using diamond drill data. Wessley Edgar of Mining Employment Services assisted with structural logging and integrated the results of structural logging with other geological data in order to develop a 3D geological model of the Western Lode.

Below is a summary of the work completed by Beeson and Edgar (Beeson 2015) during late 2015;

1.

Structural logging of mineralized intervals and adjacent intervals intersected in diamond core through Western Lode (16 diamond holes on five drill sections).

   
2.

Visualized existing drill data and wire frames in 2D and 3D (Micromine software).

   
3.

Interpreted structural features (faults, major contacts, folds) and gold mineralized domains.

   
4.

Constructed a preliminary 3D interpretation of the Western Lode geology and gold distribution.

   
5.

Produced synthesis figures illustrating the interpreted shoot-like nature of mineralized domains and their relationship to key geological features, as well as the relationship between the F1 and F2 Faults.

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The mineralization intersected in the Western Lode appears to be hosted in one or more tabular, plunging, shoot-like bodies oriented sub-parallel to the hinge of the Cosmo-Howley Anticline. These bodies may link between drill sections along fold limbs, following the sigmoidal geometry of the folded nodular greywacke host sequence located in the footwall of the F1 Fault. Some potential gaps are apparent in the existing drill coverage that may warrant consideration for infill drilling. The F2 Fault appears to be a key master structure developed along the western limb of the Cosmo-Howley Anticline. The F1 Fault appears to be a fault splay localized in the footwall of the F2 Fault.

This study of the Western Lode system recognizes the structural controls on gold mineralization and associated veining related to folding and faulting as outlined in the 2014 technical report (Smith and Edwards 2015). Fold plunge, together with proximity of the host nodular greywacke unit to fault zones such as the F1 and F10 Faults, appear to be a significant influence on mineralization and vein development in the host nodular greywacke sequence. Gold mineralization in the Western Lode is associated with arsenopyrite, both as fine to medium grained disseminations along arsenopyrite-bearing quartz-carbonate-sulfide veins, and crystal aggregates concentrated around the margins of silica nodules in the host greywacke. Multi-stage syn- and post-mineralization veining is evident in the Western Lode. Broadly, syn-mineralization quartz-carbonate-sulfide tensile and shear veins formed in various orientations commonly show a spatial association with gold mineralization. The selvedges of quartz-carbonate-sulfide veins typically host relatively higher-grade gold mineralization than the veins. A penetrative structural fabric is locally developed in the Western Lode host sequence, oriented approximately axial planar to the Cosmo Anticline. Locally, this fabric has been crenulated, mostly in proximity to late veins.

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Figure 9-7 displays major structural features and surfaces are annotated on the oblique view and structural data measurements are shown as apparent dips and strikes (black = bedding measurements; red = vein measurements). The northern and central mineralized domains probably join into a single continuous mineralized domain thus representing a sigmoidal mineralized volume enclosed within the folded nodular greywacke. A third domain of mineralization is evident on one of the more southerly drill sections and may also connect through the folded greywacke sequence to the mineralization intersected on the central and northern drill sections.

Lantern Exploration Targets (Inner Zamu Dolerite Meta-sediments)

The Lantern exploration target, previously named the Inner Metasediments Target, is located down plunge of the previously mined Phantom open pit. The target was proposed in 2014 to test between prospective results in two existing surface drill holes the results of which are 340m apart. The drill intersections are located close to the Inner Sediment Contact with the Zamu Dolerite on the western limb of the Cosmo Anticline (Figure 9-9).

An initial phase of drilling consisting of 2 holes were drilled in the first half 2015 to evenly bisect the existing results in hole CP009W1 (2.14m @ 8.34g/t Au) and hole PHP0001 (8.6m @ 5.14g/t Au) a small distance inside the Zamu Dolerite limb contact. Not unexpectedly both holes encountered strong gold grades (see Table 9-1) just 10m to 25m from their drill collars and with continued similar intersections in the 2nd Phase of diamond drilling this cluster of gold results became known as the 700 Lode and was the focus of both underground trial mining and subsequent drill program, which is in progress (2016). The equivalent stratigraphic Zamu Dolerite contact on the eastern flank of the Cosmo Anticline intersects the main decline development where it has been demonstrated to contain sporadic gold mineralization within adjacent metasediments (see Figure 9-12).

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Of greater possible economic significance from the 1st Phase drilling was the highly hematite-sericite-chlorite altered fine grained metasediments central to the core of the Cosmo Anticline, which contained high gold grade drill results. Best results at that time included 6.15m @ 6.79g/t Au, 22.75m @ 4.27g/t Au and 13.7m @ 4.44g/t Au in hole CW101002 (Table 9-1).

The mineralization in hole CW101002 was recognized as not typical for the Cosmo area occurring as an isoclinal folded horizon, fine grained metasediments, which appears as a strongly green color owing to the chlorite + sericite +/- carbonate alteration and regions of intense hematite alteration. Sulphides present are generally pyrite but are of modest quantities compared to gold bearing greywackes within the main Cosmo mineralization.

Lantern (“Green Rock”) Review (Millar, June 2015)

In June 2015 John Miller conducted a short study of the initial Lantern area drill results. This work characterized the gold and alteration styles found and developed an excellent cross section structural interpretation for the anomalous hole CW101002 drill core (Miller, 2015b).

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All of the holes drilled across the Lantern target encountered gold mineralization approximately 11-13m beyond the drill collar. The host horizon of altered siltstone adjacent to a vertical dolomite bed became known as the 700 Lode and was subject to a small exploratory development sill from the 1010RL cuddy. Results of the underground mining were strong gold grades to the southern mineralization drive but lower to the northern side. It was then concluded that a steep plunge to the 700 Lode was likely and development halted until a diamond drill program could test the areas below the known mineralization at 1010RL.

This area holds good potential for additional mineral resource with past shallow drill holes containing widths up to approximately 10m and results of 6.71g/t Au over 3.95m in hole CW101003 (Table 9-1) and 4.04g/t Au over 4.7m in hole CW101002 (see Figure 9-9). These intersections are close to current development and planning has been completed on the next phase of in 2016.

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Hole ID Intercept (Down-hole widths) Lode
CW101001A 7.54m @ 5.33g/t Lode 700
CW101002 4.7m @ 4.04g/t Lode 700
CW101002 3.25m @ 4.78g/t Lantern Central
CW101002 3.3m @ 2.84g/t Lantern Central
CW101002 6.15m @ 6.79g/t Lantern Central
CW101002 22.75m @ 4.27g/t Lantern Central
CW101002 13.7m @ 4.44g/t Lantern Central
CW101003 3.95m @ 6.71g/t Lode 700
CW101003 3m @ 2.24g/t Lantern Central
CW101003 5.75m @ 3.86g/t Lantern Central
CW101004 5.05m @ 3.03g/t Lode 700

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Hole ID Intercept (Down-hole widths) Lode
CW101004 3.25m @ 3.18g/t Lantern Central
CW101004 3.4m @ 2.44g/t Lantern Central
CW101004 5.3m @ 2.15g/t Lantern Central
CW101004 5m @ 2.43g/t Lantern Central
CW101005 5m @ 2.43g/t Lode 700
CW101005 4.2m @ 7.16g/t Lantern Central
CW101006 4.65m @ 3.41g/t Lode 700
CW101006 6.0m @ 9.64g/t Lantern Central
CW101006 2.05m @ 7.98g/t Lantern Central
CW101006 2.9m @ 4.37g/t Lantern Central
CW101006 4.1m @ 6.13g/t Lantern Central
CW101006 2.5m @ 5.36g/t Lantern Central

TABLE 9-1 LANTERN TARGET DRILL INTERSECTION ASSAY RESULTS

Figure 9-12 shows a cross section looking NNW across the Lantern target area of inner metasediments between the folded Zamu Dolerite limbs to the Cosmo Anticline. The Cosmo & Phantom open pits are also shown, as is the Phantom Dolerite which cores the anticline and hole traces from the CW1010 Phase 2 drilling conducted in 2015. The pink to blue thematic cirles are gold in drilling 2m > 2g/t Au composite intersections.

All six diamond drill holes reviewed for Lantern were drilled in a broadly SW-plunging fan from the 1010 drill cuddy. These holes initially encountered nodular greywacke of the 700 Lode (Eastern Lode) in the top 20-30m of the hole, then drilled through a 20-30m sequence of well-foliated phyllite, with abundant sericite/chlorite poikiloblasts (pseudomorphs after cordierite), prior to encountering rocks that host the Lantern mineralization. The Lantern host rocks comprise an intercalated and folded sequence of greywacke and siltstone. Gold mineralization is associated with a prominent banded, dark green-colored alteration assemblage comprising chlorite-sericite-carbonate-pyrite (retrograde after an amphibolite-facies biotite-quartz-Kfeldspar-hornblende-garnet-tourmaline assemblage); the retrograde alteration assemblage is locally red-brown in color reflecting hematite (dominantly as pigmentation of carbonate according to (P. Ashley 2015). Examples of these rocks are shown in Figure 9-8. Domains of the chlorite-dominant assemblage alternate with pale-grey colored greywacke and siltstone that preserve the prograde amphibolite facies assemblage; these alternations reflect folding along the western limb of the Cosmo Anticline.

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The Lantern host sequence has been moderately to strongly deformed and a variety of structures are apparent, including folds, penetrative fabrics, faults, breccia zones and multi-stage veins. These structures have been logged and used to assist in interpretation of a preliminary geological cross section through the drilled area at Lantern. Key features recognized during structural logging are summarized below:

Lantern XRF Litho-Geochemical Analysis

In late 2015 a litho-geochemical study was initiated across the available 6 Lantern diamond core holes (CW1010 prefix) using a handheld Niton XRF analyzer. Approximately 810 XRF analyses were made across the 6 holes to obtain a suite of 38 elements. This data was then merged with geological logging information to extract the fire assay gold grades, logged lithology and alteration present for each XRF data record. The analyses were deliberately made on the least altered portions of the core samples, avoiding veining and other irregular features, which may cause spurious results. The purpose was to see if the Lantern Sequence could be characterized and specific stratigraphic horizons most prospective for gold could be identified.

Graphical results of this work are presented in Figure 9-13.

Areas of high total iron are only temporally associated with gold, and not highly correlated with the gold grades and significant intersections.

Base metals Cu, Pb and Zn do show a degree of correlation with high gold grades especially zinc in the central Lantern portion of holes. Elevated Cu, Pb and Zn values were found to be associated with the 700 Lode to a modest degree. Base metal anomalism on the western limb, which is not highly related to gold is of interest.

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Lantern Conclusions & Recommendations:

Given that the Lantern target has a distinctive hematite- and ‘green’ bedded chlorite- alteration assemblage compared to the nodular greywacke + Pmc stratigraphy, which is strongly correlated with Cosmo gold mineralization, it is logical to conclude there is a strong alteration relationship to gold grades. This conclusion, however, is possibly too simplistic and not fully supported by logging and geochemical data;

These conclusions must be cautioned with the comment that narrow and imprecise handheld XRF data is not a bulk geochemical method and results are subject to high error and biased by the geologist’s selection of sites analyzed. Also the logging of hematite, chlorite, silica and carbonates is less quantitative, with inconsistency errors due to different geologists subjective logging practices.

In general terms, silicification in the presence of moderate to strong chlorite and/or hematite alteration is a good indicator of gold potential across the Inner Metasediments. Logging drill data suggests that the hematite-chlorite-silica alteration represents an alteration footprint that is broader in scale than the gold metal distribution, and so may provide a useful near-miss indicator in sparsely drilled areas close to the Cosmo Mine.

Recommendations concerning this new exploration region of the Cosmo Mine are;

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Footwall Fault Hinge Zone

The thrust movement associated with the F1 Fault has been estimated to be in the order of 100m. The estimated location of the Cosmo Anticline Hinge Zone (the 550 and 650 Lodes) transformed back along this thrust displacement (to the projected pre F1 Fault location) is estimated to be located in the footwall of the F1 Fault (as shown in Figure 9-15). This zone forms the Footwall Fault Hinge target. To date three drill holes have intersected this zone and the planned program will build upon these drilling intersections.

9.2 EXPLORATION PLANS FOR 2016

9.2.1      COSMO EXPLORATION 2016

An active near mine growth program of exploration is planned for 2016 at Cosmo Mine. This work is mainly drilling based but is designed to confirm the understanding as outlined above. A summation of this work is outlined below.

The continuing success of the Sliver Lode to add mineral resource down plunge and confirmation of significant Sliver gold mineralization at 2200mN make this 160m of strike a priority target for additional drilling in the first half 2016 (Figure 9-15). To facilitate suitable drill intersections and desired target accuracy an extension to the 640 drill drive is planned in Q1 2016 (Figure 9-16).

Complimenting this north mine Sliver exploration will include Eastern Lode extensions down plunge from the 640 (extended) and 840 drill drives underground platforms. Known as the Cosmo Deeps series of drill programs this area is expected to add substantial mineral resources and mine life to the Cosmo mining operation in 2016.

The Hinge mineralization below the F1 Fault can be drilled as a secondary target in many of the Sliver drill holes and is expected to provide a number of small to modest, yet possibly high-grade short-term mineral resource targets.

Other exploration drill programs will test the Western Lodes and emerging Lantern area targets.

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9.3 UNION REEFS AREA

Historical exploration activity is summarized in Section 7.3 of this report. The exploration work completed by Crocodile Gold/Newmarket Gold relates to improving the geological understanding. This information is outlined below.

9.3.1      ELIZABETH MINE AREA

In 2012 Crocodile Gold elected to selectively look at some of its gold assets in the Union Reefs area to ultimately determine if some of them required additional work to determine their potential economic viability. The Elizabeth Mine was chosen for further study due to the high-grade nature of past production and its proximity to the Unions Reefs mill facility.

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R.M. Biddlecombe (1985) (Biddlecombe 1985) estimated a possible mineral resource volume of approximately 300,000t of material grading 8g/t Au and 25g/t Ag based on an unsubstantiated outline of mineralization displayed in Figure 9-18. Without further drilling and confirmation at depth there is no reason to believe this mineralization may be present (NB1). An apparent examination of the old underground workings by Biddlecombe indicated everything within 30m of surface has been mined out. The Author has noted that there is evidence of open stopes at surface.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves.

The historic workings at Elizabeth are centered on narrow 1 to 2m wide steeply east dipping shear hosted quartz veins near the interpreted western margin of the Pine Creek Shear Zone. The old workings occur over two NW striking ridges that are cut by the McKinlay River with the northern ridge having been subjected to more extensive historical mining activity.

During the period 1875-1897 the Elizabeth Mine reportably produced 3,450oz of gold averaging about one ounce per tonne (Stuart-Smith P.G. 1993). The down dip and plunge potential of mineralization is largely untested.

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ID Easting Northing   Type Location
EL-01 797694 8489289 chloritic slate with qtz dump N most shaft, central workings
EL-02 797713 8489226 Qtz with chloritic slate dump central workings
EL-03 797721 8489140 chlorite, sericite altered bedrock shaft dump E of central workings
EL-04 797715 8489139 gossan like oxidized qtz with bedrock shaft dump E of central workings
EL-05 797710 8489140 sericitic bedrock with trace qtz shaft dump E of central workings
EL-06 797820 8489156 Qtz with altered slate and greywacke dump next to E most shaft, workings
EL-07 797880 8488981 massive oxide coated qtz incl asp dump Schoolteachers Adit
EL-08 797885 8488973 massive qtz incl asp, ch se altered bedrock dump Schoolteachers Adit
EL-09 797885 8488973 qtz with weathered ch se bedrock dump Schoolteachers Adit
EL-10 798333 8488251 fe qtz with altered bedrock Chinese workings, high Au area
EL-11 798270 8488244 Qtz with silicified greywacke Chinese workings, high Au area
EL-12 798114 8488302 moderately oxidized qtz with si greywacke Chinese workings, high Au area

TABLE 9-2 ROCK CHIP SAMPLING INFORMATION FOR ELIZABETH

Sample Au Ag Al As Ba Be Bi Ca Cu Fe Mn P Pb S Sb Zn
Description pp
m
pp
m

%

ppm
pp
m
pp
m
pp
m

%
pp
m

%

ppm
pp
m
pp
m

%
pp
m
pp
m
EL-01 0.01 3.5 0.62 89 230 1.3 <2 0.06 98 12. 7 16550 300 2050 <0.01 17 1600
EL-02 0.01 3.1 1.04 60 60 0.7 3 0.02 283 3.52 523 240 415 <0.01 126 683
EL-03 <0.01 0.3 3.38 55 90 0.9 <2 0.67 12 6.07 1245 790 23 <0.01 7 132
EL-04 0.07 3.1 0.33 940 40 <0.5 <2 0.01 27 1.5 199 90 307 <0.01 31 119
EL-05 0.1 5.4 0.72 1110 60 0.6 <2 0.01 47 1.8 2 41 160 1015 <0.01 41 125
EL-06 2.72 10.8 0.73 563 70 1 <2 0.02 107 5.1 4230 290 594 <0.01 42 749
EL-07 15.2 11.8 0.4 >10000 50 <0.5 4 0.29 424 2.81 109 870 6530 0.34 89 285
EL-08 0.03 1.2 0.56 574 50 0.5 <2 0.01 16 1.5 8 108 270 359 <0.01 9 130
EL-09 8.57 23.8 0.12 >10000 30 <0.5 5 0.01 111 5.9 2 98 280 2660 0.11 149 37
EL-10 0.05 <0.2 0.14 373 20 <0.5 <2 0.01 37 9.8 4270 120 255 <0.01 5 233
EL-11 <0.0 1 0.4 0.48 116 30 <0.5 <2 0.01 8 2.17 158 160 52 <0.01 4 28
EL-12 0.18 0.3 0.46 1250 180 <0.5 2 0.02 12 6.4 6 155 400 181 <0.01 21 48

TABLE 9-3 ROCK CHIP SAMPLING ANALYTICAL RESULTS FOR ELIZABETH

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9.3.2      STRUCTURAL STUDIES AND 3D MODELING

Esmeralda and Caroline (collectively under ML27999) and on the Mary River Pastoral Lease covers approximately 834 hectares and are situated 170km southeast of Darwin and 8km northeast of Pine Creek (Figure 9-20.) (J. Shaw 2005). It is located 6km down strike (southeast) of the Union Reefs processing facility. Underlying lithologies belong to the Mount Bonnie Formation. Mineralization has previously been described as being hosted within anticlinal shears and subsidiary fault splays in Zone A with anticlinal shears, analogous to the southern Union Reefs mineralization style within Zone B.

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Mapping Program - 2014

Crocodile Gold commissioned W.P. Karpeta of Bastillion Pty Ltd to conduct a mapping project over the Esmeralda project area for six weeks from April 15, 2014 to May 25, 2014.

The Esmeralda Area has three main areas of gold mineralization, Esmeralda A, B and C, together with a single known base metal occurrence, Caroline.

The aim of the mapping exercise was to establish the distribution and geometry of the mineralization in areas A, B and C (Caroline South). Grab samples were taken of key rock types during the mapping field work. The results of the mapping exercise were as follows:

Sedimentology & Palaeovolcanology

Five lithofacies (Lithofacies A to E) have been identified in the Mount Bonnie Formation in the Esmeralda area; slates, greywackes, cherts & massive and banded siliceous rocks. These are thought to represent a shallow, mostly muddy marine shelf environment (slates and greywackes) influenced by occasional submarine felsic volcanic eruptions (cherts, massive and banded siliceous rocks). These felsic volcanic systems may have produced VMS deposits, which could have had low-grade gold mineralization (~0.1g/t) that was remobilized during regional and/or thermal metamorphism and concentrated in suitable sites. The mechanism appears to be quite common elsewhere in the world.

Intrusives

A single example of an aplitic sill, 5cm thick, was seen in the Esmeralda A area in a greywacke/slate sequence approximately 100m away from the granite/sediment contact. This may be associated with the Allamber Springs Granite.

A deeply weathered, NW-SE striking 5m thick dolerite dyke was seen in the road cut in the southern part of Esmeralda and another deeply weathered NW-SE striking 2m thick dolerite dyke was seen in an old trench just north of the Caroline area. These may represent strike continuations of the same intrusive and can be traced from aeromagnetic results. They may represent Zamu type dolerites. Two other thin (<1m) NW-SE striking lamprophyric dykes were seen in the Esmeralda A area and south of Caroline Hill.

The large irregular intrusive body of the Allamber Springs Lobe of the Cullen Batholith marks the eastern boundary of the Esmeralda area. Its intrusion has hornfelsed the sediments in contact with it.

Structure

Bedding in the Esmeralda area is dominated by steep dips (>70°) towards the NE and SW, related to the tight to isoclinal NW-SE striking folding. Younging directions marked by graded bedding and ripple marks were observed on bedding planes enabling the fold facing to be ascertained and indicating that the western limbs of the major anticlines are mostly the right way up and the eastern limbs are overturned.

At least three foliations/cleavages have been observed, an earlier axial planar cleavage dipping steeply SW associated with the tight NW-SE striking folds, a NW-SE zonal anastomosing shear foliation associated with the strike-slip faulting and a later cross-cutting moderately eastward dipping fracture cleavage possibly associated with the Allamber Springs Granite intrusion.

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Three phases of folding were seen in the Esmeralda area, an early steeply plunging, asymmetric set of folds confined to Lithofacies E, interpreted as flow folds in felsitic lava flows, a subsequent upright, shallow plunging tight fold set formed under NESW compression and a later set of tight steeply plunging small scale folds associated with sinistral and occasionally dextral shear zones.

Early bedding plane parallel thrusting can be seen in the Esmeralda area marked by small duplexes with northeastwards vergences suggesting an early SW-over-NE thrusting event. This early thrusting may have been responsible for the initial folding of the rocks whereas later thrusting resulted in the displacement of fold axes to the NE, as at Esmeralda A.

NW-SE and NE-SW striking shear zones are found in the Esmeralda area marked by shear foliation, asymmetrical, steeply plunging folds and sulphide-mineralized brecciation in greywackes. A dominant early sinistral but subordinate later dextral sense of movement is indicated.

Several generations of quartz veins have been encountered: early barren, massive, white quartz veins, then subsequent barren, sheared white quartz veins and finally late massive and brecciated auriferous grey quartz -tourmaline veins. The early massive white quartz veins appear to represent saddle reefs on the culminations of second order upright folds (F2) of the D2 deformational event. The sheared quartz veins occur within and between the sinistral shear zones of the D3 deformational event. The late massive gold bearing quartz tourmaline veins are also found within and between the sinistral shear zones of the D3 deformational event but closer to the granite. Despite their differing parageneses, all these quartz vein generations appear to be parallel to bedding.

Mineralization and Alteration

Five alteration types were noted during mapping comprising silica, epidote, chlorite, tourmaline, and sericite.

Silica alteration involves the silicification of the massive and banded siliceous lithologies (Lithofacies D and E) and the country rocks around the quartz veins. The former is probably due to seafloor early diagenetic alteration/remobilization of silica in felsic volcanic rocks.

Pyrite alteration occurs in two settings: as randomly distributed euhedral-subhedral cubes up to 20mm across in the massive and banded siliceous lithologies (Lithofacies D and E) and as disseminated anhedral aggregates around quartz veins and in altered/sheared greywackes. The former may be due to early propylitic alteration and the latter due to alteration by regional and/or thermal metamorphic fluids.

Epidote alteration comprising irregular fine-grained patches of epidote was noted in the massive and banded siliceous lithologies (Lithofacies D and E), especially near the thicker developments of these lithologies. This alteration is possibly related to propylitic alteration produced by the migration of Fe- and S-bearing hydrothermal fluids.

Chlorite alteration was again noted in the massive and banded siliceous lithologies (Lithofacies D and E) associated with epidote-pyrite alteration.

Fine needles and rosettes of black tourmaline are found in the auriferous quartz veins at Esmeralda A but are rarer elsewhere in the meta-sedimentary rocks at Esmeralda.

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Sericite alteration comprising irregular disseminated books of sericite is found mainly around shears and quartz veins associated with pyrite and tourmaline alteration especially in greywackes.

Metamorphism of the rocks in the Esmeralda area has resulted in two types of metamorphic alteration, an earlier regional metamorphism and a later thermal metamorphism. The regional metamorphism is generally Upper Greenschist with the formation of biotite, chlorite, andalusite and garnet in the shales. The Zamu Dolerites also show Upper Greenschist metamorphism with the formation of amphibole, chlorite, sericite, biotite and albite. The thermal metamorphism associated with the intrusion of the Allamber Springs Granite has resulted in the formation of garnet, amphibolite and cordierite in a sillimanite to hornblende- hornfels 300m wide around the granite.

Gold Mineralization

Gold mineralization was detected in four areas, the already drilled Esmeralda A and B areas, an area of shearing south of Caroline (Esmeralda C) indicated by geochemical anomalies and rock chip samples and an area to the NW of Esmeralda A and to the NE of Esmeralda B (here termed Esmeralda D). Base metal mineralization has been noted at Caroline. The gold mineralization in Esmeralda A occurs in a series of NNW-SSE striking, bedding plane parallel quartz-tourmaline veins associated with pyrite-sericite alteration in a sequence of alternating slates and greywackes. The extent of this gold mineralized vein system is governed by a WNW-ESE striking cross fault to the north and the hornfelsed aureole of the Allamber Granite to the south. The gold mineralization at Esmeralda B again occurs in a series of NNW-SSE striking, bedding-plane parallel quartz veins in an alternating slate/greywacke sequence. This mineralization appears to be cut off to the north by the same WNW-ESE striking cross fault. The southern end of Esmeralda B is not constrained but disappears under a cover of siliceous rubble towards Caroline Hill. The gold mineralization at Esmeralda C occurs in a NNW-SSE striking sinistral 5m wide shear zone cutting through a 20m thick greywacke unit and is limited by cross faulting to the north but is unconstrained to the south. The gold mineralization at Esmeralda D was located by chip sampling (up to 0.3g/t Au) and comprises alteration in the culmination of a major NNW-SSE striking anticline. It is cut off to the south by the same WNW-ESE striking cross fault as Esmeralda A but is unconstrained to the north.

The base metal prospect at Caroline has been looked at in the field and comprises massive sulphides in an epidote-chlorite altered brecciated siliceous bed (Lithofacies D). It is thought to represent a poorly developed VMS body and may indicate that other better developed VMS bodies may be in the area. Some gold mineralization is also present.

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Along with the structural measurements, a total of 64 grab samples were collected from the Esmeralda deposit (Figure 9-24). These samples were tested for gold at North Australian Laboratories by FA50 fire assay. The results for these samples can be viewed in Table 9-4 below.

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Easting Northing Au   Easting Northing Au
805373.9 8477941.5 0.33   806079.0 8477487.0 0.15
805350.2 8477974.5 0.04   805988.0 8477525.0 -0.01
805339.8 8477996.1 0.13   806031.0 8477383.0 0.04
805340.5 8477995.5 0.01   806217.0 8477300.0 0.1
805537.6 8476986.8 -0.01   805395.0 8477802.0 -0.01
805563.1 8477025.4 0.22   805376.0 8477838.0 0.02
805565.7 8477023.2 0.04   805151.0 8477942.0 0.04
805661.6 8476999.8 0.08   806571.0 8476226.0 0.01
805889.9 8477538.6 -0.01   806732.0 8476090.0 0.22
805839.4 8477754.7 -0.01   807183.0 8475778.0 0.06
805802.9 8477748.2 0.01   805126.0 8477505.0 0.97
806158.0 8476654.0 -0.01   806963.0 8476089.0 0.03
806065.0 8476696.0 -0.01   806628.0 8476385.0 0.03
805314.6 8477405.1 2.34   806607.0 8476283.0 0.03
805729.9 8476940.5 0.11   805817.0 8476692.0 -0.01
805752.2 8476909.9 0.01   805820.0 8476675.0 0.1
806030.6 8476721.7 -0.01   805643.0 8476578.0 0.11
805845.0 8476925.1 0.18   805643.0 8476578.0 0.89
805207.0 8477504.0 0.17   806165.0 8476475.0 0.25
805202.0 8477499.0 1.04   806159.0 8476472.0 0.26
805085.0 8477601.0 0.09   806212.0 8476462.0 0.19
805267.0 8477448.0 0.03   806213.0 8476463.0 0.34
805222.0 8477657.0 0.01   806252.0 8476328.0 0.04
805305.0 8477565.0 -0.01   806331.0 8476443.0 -0.01
805205.2 8477233.0 0.06   806073.0 8476599.0 -0.01
805102.1 8477189.3 0.02   806314.0 8477236.0 0.11
805050.0 8477207.0 -0.01   806577.0 8477020.0 -0.01
805040.8 8477208.8 0.05   805570.0 8477849.0 -0.01
805030.3 8477242.3 -0.01   805082.0 8478093.0 -0.01
805032.0 8477550.0 0.09   805082.0 8478093.0 -0.01
805080.9 8477606.4 0.99   805082.0 8478093.0 -0.01
806205.0 8477384.0 0.19        

TABLE 9-4: 2014 ESMERALDA GRAB SAMPLE RESULTS AU G/T

Historical Exploration Work and Results

Prior to the 2015 RC and diamond drill program at Esmeralda Zone A and B, Esmeralda and Caroline had 179 RC holes, three diamond tails and two diamond holes to develop an Inferred mineral resource of 1.26Mt at 1.62g/t Au with a 0.7g/t Au cutoff. This is comprised of 550,000t of oxide at 1.58g/t Au, transition of 120,000t at 1.5 -2g/t Au and a fresh mineral resource of 590,000t at 1.67g/t Au, all Inferred (NB1). Twenty eight of these holes are within the Caroline leases, down strike from Zone B and have not been historically reported in conjunction to the Esmeralda holes and are not a part of the current mineral resource.

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves.

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Exploration over these leases is relatively mature with infill drilling within the current mineral resource and further drilling down strike of Zone B proposed to increase confidence and volume of what has been previously modeled.

No mining activities have been recorded on either Esmeralda or Caroline; however, some small workings and an in-filled shaft have been recorded at Caroline (Z. Bajwah 2007c).

The Armadeus Gas Pipeline crosses the east flank of Zone A, which has potential to effect project economics (B. Makar 2005a) (Figure 9-25).

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Geomin Exploration undertook stream sediment and follow up soil sampling and costeaning in 1968, defining the Caroline base metals anomaly (Z. Bajwah 2007c); (Butler 1991). Costean results adjacent to the old shaft included 3.64m with 5.2% Pb, 0.47% Zn and 40.3g/t Ag. Over a different anomaly a costean is reported to have intersected 1.83m with 2.9% Pb, 0.35% Zn and 102.3g/t Ag.

Cyprus established the Zone A and B deposits through a soil sampling program in 1990-1991. Zone A was found underlying a gold anomaly and Zone B underlying an arsenic anomaly. Follow up rock chip sampling confirmed the anomalies and drilling over the zones was undertaken. In 1992-1993 Cyprus undertook an IP/resistivity survey, which suggested that Zone A deposit was offset to the west at the south end and did not pass under the gas pipeline although drill results do not support this. In 1996 Acacia Resources commissioned an airborne magnetic and radiometric survey to be undertaken. From 1994 to 1999 Acacia Resources held the license to Esmeralda. Nine costeans were dug along grid northings of anomalous gold sites in 1996.

In 1997 50 RC holes and one re-entry were completed.

In 1998 Acacia Exploration Darwin completed a rock chip sampling program over potassium-altered zones between Zones A and B with no significant results.

In 1999 channel chip sampling was undertaken over a quartz tourmaline vein area with no results considered worthy of follow up from the 30 samples taken.

2000-2001 AngloGold took over Acacia Resources but did no fieldwork.

In July 2003 AngloGold closed Union Reefs and put the property up for sale; in 2004 Burnside Operations took over the leases. In 2005 Burnside Operations undertook a review of previous work done and created a work proposal for 2006.

9.4 PINE CREEK EXPLORATION

9.4.1      INTERNATIONAL DEPOSIT ENVIRONMENTAL DRILLING

Exploration activity over Pine Creek area since the mine closures and rehabilitation in the 1990’s has been very limited. Two environment assessment holes were drilled at the north end of the International pit (MLN1130) during November 2010 for the purpose of waste rock classification. The area was backfilled with waste rock after the pits closure in 1994. Five meter interval samples were collected from each hole and assayed for total sulphur, ANC (Acid Neutralizing Capacity) and NAG (Net Acid Generation). Figure 9-26 illustrates the holes drilled near the International pit.

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The results from the test work were to be used to confirm the acid base characteristics, existing acidity, elemental composition and solubility and provide data for developing and evaluating waste rock management options for re-establishment of mining operations within the International pit.

A reconnaissance of MLN1130 was done on June 2012. Access and the state of the pits were quickly assessed.

9.5 BURNSIDE AREA EXPLORATION

9.5.1      VTEM SURVEY

In 2011 a VTEM survey was flown over five separate areas within the Property area. A total of 2,811 line kilometers were flown at 150m line spacing covering a total area of 332km2. The five areas represented different geological environments with an emphasis on the same stratigraphy that hosts the Cosmo Mine. Hosting mineralization contains significant amounts of carbonaceous material as well as pyrrhotite, both of which are conductive. The northeast Burnside area was surveyed as it was indicated that here were extensive black hydromorphic soils in this area and it was felt that in all likelihood the VTEM survey could “see through” this cover. The geophysical signatures of deposits such as the Brock Creek deposits, Woolwonga, Glencoe, Mt Bonnie, Iron Blow, Princess Louise, Bon’s Rush, Mt Ellison, Rising Tide and strike extents of Cosmo Mine were determined.

9.5.2      COSMO SOUTH

9.5.2.1  Stream Sediment Survey

In order to quickly assess a large area of approximately 50km2 to the south and west of the Cosmo Mine it was decided to carry out a stream sediment survey. The program was designed to sample significant streams at 500m intervals along their course. Contractors Arnhem Exploration, out of Tenant Creek, carried out the sample collection. They were instructed to collect a minimum of 200 grams of -75 micron material, which was sieved in the field. All sample sites were photographed. A total of 69 sites were sampled. Figure 9-27

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Past surveys in the region have been carried out using BLEG sampling of -2mm material but it was thought that the sulphide associated style of mineralization found at Cosmo Mine would be more suited to sampling a finer grain size material.

Samples were submitted to ALS for multi-element analysis by package Gold Au-AA21, Multi-element ME-MS41. No QA/QC standards or duplicates were submitted in the sample stream. Site variability was addressed by collecting from 4-5 sites in close proximity at any particular sample location.

The most obvious interpretation is that the survey defines the Cosmo Mine. No doubt extensive mining activity over the past century or more has contaminated a fairly large area and this likely contributes to the anomaly in Au and As and other elements. Nevertheless, the stream draining the mine area is anomalous as it should be. As one proceeds downstream it crosses the same stratigraphy that hosts the Cosmo Mine and the anomalism in Au increases indicating that this horizon is likely favorable for gold mineralization. It is a priority target for further follow-up exploration work.

The streams to the east of the stream flowing out of the Cosmo Mine area are also anomalous in gold and arsenic. They drain areas of stratigraphy that host the Cosmo Mine and are considered areas warranting further exploration.

The Liberator deposit centered at 758000E 8501200N and extending along a NW-SE strike for approximately 600-800m is defined by the stream sediment survey. This may be lower in the stratigraphic sequence or conversely it may be emplaced due to east-west thrust faulting. Further investigation is required.

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Streams in the western part of the survey area are outside the Newmarket Gold tenement area but are underlain by favorable Koolpin Fm stratigraphy. Gold values are generally low and no priority areas for immediate follow-up exploration work are apparent.

9.5.2.2   Geological Mapping

In 2011 David Reid, a consulting geologist, was retained by Crocodile Gold to examine the strike extents of the Cosmo Mine horizon to the south. A number of weeks were spent geological mapping and investigating airborne EM responses. Unfortunately, this work was done before the VTEM survey results were available so only the broad line spacing 2009 government survey results were available. Unfortunately, the emplacement of conductive horizons was too imprecise to allow good ground follow up.

9.5.2.3   VTEM Survey

A VTEM survey has been flown over the strike extensions of the Cosmo Mine stratigraphic horizon to determine the location of potentially favorable targets that may host similar mineralization that is found within the Cosmo deposit. The actual Cosmo deposit was not flown, as there are too many cultural effects that would mask/interfere with the deposits geophysical signature.

It is interpreted that the Cosmo deposit’s geophysical signature will be a strong EM response with a coincident magnetic signature. Both graphite and pyrrhotite are very good conductors. Pyrrhotite is usually magnetic.

The VTEM survey has defined the host Cosmo horizon over many kilometers to the southwest with responses being variable but display increased strength at flexures and intersection points. The magnetometer survey indicates that only selected parts of the conductive horizon have a coincident magnetic anomaly. Particular attention should be paid to these areas. It is interesting to note that there is a Cosmo look-a-like target at the west end of the survey area. That being what is interpreted to be a NW trending anticlinal structure.

It is interesting to note the differences between the governmental published geology map and the VTEM results. Initial interpretations would indicate that there is more Koolpin Fm stratigraphy in the area than is shown on the published geology maps.

Figure 9-28 displays the VTEM conductor axis and the Channel 42 results. The stratigraphic horizon that hosts the Cosmo Mine is clearly displayed. The conductivity is in all likelihood caused by carbonaceous material that occurs stratigraphically within the Koolpin Fm. The aeromagnetic map indicates that parts of the conductive horizon defined by the VTEM survey are also magnetic. It is interpreted this magnetism is caused by either Zamu Dolerite or pyrrhotite (or both). Both are known to occur as stratigraphic horizons within the Koolpin Fm in close proximity to the carbonaceous horizons.

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Airborne radiometric data, available for the area, indicates that the VTEM conductor axis is anomalous in Uranium. Conceivably the carbonaceous material is acting as a catalyst and low quantities of uranium are being captured by this extensive horizon. The Uranium channel data provides a very good correlation with the VTEM conductor.

9.5.3      MT BONNIE EAST

Initial interpretations of the VTEM airborne data indicated a strong west dipping conductor with a direct magnetic correlation that seemed to sit in a stratigraphically similar position as the Iron Blow and Mt Bonnie deposits. Available geological maps indicated that the conductor was hosted by Gerowie tuff, lower in the sequence, but it was felt this was open to interpretation and that Mt Bonnie Fm could underlie the area. The Northern Territory Government’s mineral occurrence listing (MODAT) indicates a small copper showing (chalcopyrite, malachite) in the immediate area but a field inspection indicated only dolerite in the area. Subsequent field mapping located a small copper showing a few hundred meters to the northwest of the MODAT occurrence.

A preliminary field investigation indicated the conductor’s location to be on the western slope of a north trending ridge covered in open woodland. Regolith float on the slope was composed of slate (30%) and massive quartz (70%). The rock was determined to be in-situ massive lode quartz containing cryptocrystalline sulphide and locally strong box-works after coarse sulphides. Some areas showed accessory aphanitic permangano-jarosite gossanous precipitate with sparse box work texture and no relict sulphide. The vein strike is estimated at slightly east of MGA94 north and dipping at approximately 60 degrees to the west.

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Five grab samples were taken at sites along the ridge, comprising both rock types, for submission to NTEL for ICP analysis (Table 9-5).

Sample
ID
MGA94 N MGA94 E Description Au1 Total S Fe K Ag As Ba Bi Cd
MBEX001 778351 8503357 Quartzose manganiferous jarrositic gossan 0.04 0.1 334000 450 0.15 396 27.5 3.92 <0.05
MBEX002 778348 8503354 Reef quartz with micro- sulphidic inclusions <0.01 <0.05 7420 150 <0.05 55.5 5.7 0.22 <0.05
MBEX003 778346 8503346 Quartzose manganiferous jarrositic gossan 0.39 0.1 377000 150 0.55 640 11.7 121 <0.05
MBEX004 778365 8503300 Reef quartz with micro- sulphidic inclusions & boxwork texture <0.01 <0.05 7000 50 0.1 12 3.75 0.88 <0.05
MBEX005 778365 8503281 Quartz boxworks 0.01 <0.05 84300 150 0.15 60.5 38 0.28 0.05

TABLE 9-5 MT BONNIE EAST GRAB SAMPLE ICP RESULTS

The Pine Creek 1:100,000 scale map sheet shows the site to be on an anticlinal fold in the Gerowie Tuff. No evidence for this was seen during the site visit. The ridge immediately to the west also had a prominent outcrop, and investigation of this showed it to be composed of dolerite, as were outcrops to the east. This agrees with the Pine Creek map sheet.

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9.5.4      MT ELLISTON EAST

The Mt Ellison East area was chosen for ground exploration work in 2011 based on an interpretation of airborne magnetic data that indicated a highly folded magnetic unit was present in the area. See Figure 9-31. Regional structural interpretations indicate that the extensive NW trending Pine Creek Shear passes through the area.

Geological mapping and prospecting in 2011 revealed the following:

“Several hills in the center of the area provide intermittent bedrock outcrops, although lower slopes are obscured by colluvium. Elsewhere the terrain is flat, with few outcrops except for quartz veins. North of 8525500 N and west of 775500 E, there is a low area with fairly abundant exposures of a mafic intrusive rock.

Except for the mafic intrusion, the region is underlain by sedimentary rocks which regional geology maps indicate belong to the Mt Bonnie Fm. (however, large areas are poorly exposed, so other rocks could occur). By far the most common lithology is phyllitic mudstone, generally well foliated, but not noticeably bedded on the scale of a hand specimen. The mudstone sometimes carries narrow quartzite inter-beds and at least one bed of banded iron formation. Along the eastern margin of the area, the siliceous sediments are fine grained and were mapped as chert.

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The aeromagnetic data suggest the occurrence of two tight folds in the sedimentary rocks. One of these has been confirmed by mapping, but the other closure occurs in a region of little outcrop. The banded iron formation as mapped outlines a tight fold in the west central part of the area (the actual fold closure was not located, but it must occur about as shown). The chert on the eastern side of the area likely occurs in the limb of a second fold, based on magnetic data. The fold axial surfaces, and their limbs, strike north-northwest. Dips of foliation and bedding are consistently to the west-southwest, usually at 70° or more. This implies that the folds are overturned.

The largest concentration of quartz veins follows this same strike direction and is densest along the fold axis.

There were a total of 41 rock samples collected, of these, the three highest results returned 1.28g/t, 0.69g/t and 0.67g/t Au. There were a further five samples that analyzed between 0.2 and 0.4g/t Au. All of these were collected from quartz veins.

Two of the highest grade samples were collected from quartz veins in the west-central part of the area (no. 7583, 1.28g/t Au, and 7584, at 0.67g/t Au). The quartz veins from which they were collected are about 100m apart, and about on trend with each other, but from opposite sides of the banded iron formation. One other sample (ME 7554) collected from the northern vein returned only 0.03g/t Au. Box-works, hematite and traces of pyrite were reported from the anomalous samples. It should be noted that an anomalous sample result is not representative of the entire quartz vein, but only of carefully selected parts of it.

Sample 7581 (0.69g/t Au), was collected from quartz near the banded iron formation on the other limb of the fold (center of the map area). The other five samples that returned between 0.2 and 0.4g/t Au were collected from other scattered locations

One sample (no. 7592) was collected specifically from banded iron formation, but returned only 0.02g/t Au.

The map area was covered by the helicopter-borne VTEM (magnetic and electromagnetic) survey. Axes of linear magnetic anomalies, and the area of an EM anomaly, are indicated on the map.

The mafic intrusive body underlies the two magnetic anomalies in the northwest part of the map, which is the likely cause. These rocks are not magnetic in hand specimen, but such a large mafic body must have a relatively high magnetic susceptibility. The next anomaly to the east is probably underlain by the mafic intrusion as well, as there are two small exposures to the east of it.

The fold in the banded iron formation in the west-central part of the area is mirrored by a narrow magnetic anomaly. However, the anomaly is displaced to the southwest, by 50 to 200m. This may be explained by the southwest dip of the bedding and the fact that the airborne data is influenced by rocks deeper in the subsurface. Note that one dip measured on BIF was quite shallow at only 30°, although this is unusual.

Closure of the eastern fold occurs in a region of little bedrock exposure, so could not be mapped. It is believed that the intermittently magnetic cherts on the eastern side of the area occur along one limb of this fold, and that they are lateral equivalents of the banded iron formation in the western fold. The magnetics on the eastern fold may similarly be displaced to the southwest of the sub-crop of the magnetic rocks.

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There are at least two gaps in the magnetic anomaly that reflects the banded iron formation. This could be due to local facies changes in the rock. However, it may instead reflect alteration and destruction of magnetite, which would be of economic interest, particularly for gold.

There is a linear magnetic anomaly, extending almost due north-south, along UTM 776100E, between 8524200 and 8525400N. This anomaly cannot be easily explained on the basis of the mapped geology. It must be real, as it occurs on earlier aeromagnetic maps as well. Most of it occurs in an area of no bedrock exposure, but at one point it crosses a hill at an oblique angle; some rocks on the hill are magnetic, but appear to strike north-northwest. However, magnetite-bearing chert float occurs along this feature at about 776100E, 8525100 to 8525200N, in an area with no bedrock exposure; this suggests a possible explanation.

A large, sub-circular electromagnetic anomaly, about 1.2km in maximum dimension, occurs in the southern part of the map area. There is almost no bedrock exposure in this region. However, a large quartz vein in the southeast corner of the map is accompanied by graphitic mudstone. The graphitic rocks are fairly extensive there. This suggests that the EM anomaly may be caused, at least in part, by graphite.

The heart of the EM anomaly occurs in an area of little exposure, but does boast numerous quartz veins. At this very spot, the magnetic anomaly fades. It is possible that these phenomena represent an area of shearing and alteration, which may be of interest. Magnetite could have been destroyed by alteration. Graphite may be expected to occur in shear zones, and in this case would not be a detrimental factor.

Prospecting conducted during the mapping program succeeded in locating several quartz veins with anomalous levels of gold (the highest being 1.28g/t Au). Each sample is not representative of its entire quartz vein; nevertheless the results were encouraging. The rock sampling program was limited and a more detailed program would very likely locate further interesting mineralization. The three highest gold values came from quartz veins near the banded iron formation. This suggests the possibility of iron formation hosted gold veins, similar to the Lupin Mine in Canada.

Narrow magnetic anomalies in the central part of the area are caused by banded iron formation and by magnetite-bearing cherts. They coincide with a mapped fold on the ground, confirming both the existence of the fold and the correspondence of the anomaly with magnetite-bearing rocks. The anomalies are clearly displaced to the southwest of the mapped horizon of magnetic rocks; this may reflect the dip of the bedding in that direction.

Spotty magnetic anomalies in the northwest corner of the region are probably caused by the large mafic intrusion that underlies that area.

A linear, north-south magnetic anomaly that occurs along the eastern side of the map area cannot be explained on the basis of the mapped geology; most of it occurs in an area of no bedrock exposure. However, magnetite-bearing chert float occurs along it at one location.

The airborne electromagnetic anomaly in the southern part of the region occurs in a region of very poor bedrock exposure. However, one occurrence of graphitic mudstone was located on the margin of the anomaly. Therefore, the electromagnetic response may be due, at least in part, to the occurrence of graphitic rocks at depth.

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The central part of the AEM anomaly is characterized by a high density of quartz veins, and a fading of the magnetic anomaly that coincides with the iron formation. These phenomena may be caused by shearing and alteration, which would be of exploration interest. Alteration may destroy magnetite and hence, any related magnetic anomaly. Graphitic rocks in this situation would be a locus for any such shearing. Therefore the presence of graphite under these circumstances could be a positive factor in localizing gold mineralization.

The initial interpretation of the AEM data has indicated that all the responses in the area are due to surficial material. It is believed that this needs re-interpreting as some of the conductors appear to be of bedrock origin. As it currently stands the interpreted conductors seem to cross-cut the local stratigraphy. It is quite possible that several of the conductors have a bedrock source. Their strongest responses seem to occur in the area of iron formation and quartz.

Research of past work in the area revealed that Paladin and Acacia Resources had carried out work in the area in 1995 and 1996 consisting of field mapping, rock chip sampling, stream sediment sampling and soil sampling. Their soil sampling grid was extensive and covered the folded iron formation area. A reinterpretation of their results is required. The laboratory results from Paladin indicate they only analyzed for gold. Acacia indicate that they analyzed for multiple elements but only gold values were plotted. Their database needs to be obtained or re-created so that all element plots can be made.

Paladin took over 100 stream sediments from a widespread area surrounding the large Allamber Spring Granite. They analyzed for gold only on samples sieved to -200 mesh. The detection limit was 1ppb Au. Their base map indicates that large parts of the area were overlain by black soils. Their conclusions were that there were no significant anomalies defined by the survey. However, a re-interpretation indicates low order anomalism is present in the iron formation area from a cluster of stream sediment samples. Quite possibly this is as much as can be expected from areas covered with black soils.

Paladin’s soil sampling was extensive to the south and west of the Allamber Springs Granite. For the most part they sampled on lines 400m apart with samples at 40m spacing. Some lines at 200m spacing were also established. Samples were sieved to -45 mesh (-40 mesh in some instances) and were analyzed for gold with a lower detection limit of 1ppb. No statistical analysis of the results appears to have been carried out on the results. There is a low contrast between background and anomalous results, possibly due to the black soil cover.

There is no map which displays plotted gold values. There are, however, data sheets with sample numbers, UTM co-ordinates and gold values. If further work in the area is contemplated then this data should be re-entered and plotted so that basic statistical analysis can be carried out and any anomalous trends can be clearly defined.

Paladin also completed 220 RAB holes over a significant area partially underlain by banded iron formation. Again low order anomalism prevailed. It does not appear that Paladin let airborne magnetometer survey results guide them in their selection of sampling areas.

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The areas of interpreted magnetic destruction and AEM conductors need to be examined in more detail. If past soil results indicate any degree of anomalism then drilling a number of holes in the area should be considered.

Several rock chip samples from this area were collect with the results shown in Table 9-6 below. It can be seen that several samples anomalous in gold were collected. These results required follow up and it has been recommended that a more regional soil sampling campaign be conducted over this area to follow up on these results.

SampleID East North Au   SampleID East North Au
ME-7526 775978 8523815 <0.01   ME-7576 775938 8523920 0.02
ME-7527 775991 8523845 0.04   ME-7577 775954 8523892 0.03
ME-7528 775991 8523845 0.08   ME-7578 775854 8523946 0.02
ME-7529 775980 8523875 0.01   ME-7579 775270 8524166 0.03
ME-7530 775975 8523892 <0.01   ME-7580 775316 8524405 0.03
ME-7531 775976 8523813 0.02   ME-7581 775655 8524701 0.69
ME-7532 776129 8524173 0.03   ME-7582 775120 8524937 0.23
ME-7546 775947 8523842 0.03   ME-7583 775141 8524675 1.28
ME-7547 775947 8523842 0.36   ME-7584 775072 8524756 0.67
ME-7548 776037 8524235 0.03   ME-7585 775852 8523946 0.06
ME-7549 776171 8524373 <0.01   ME-7586 776167 8524151 0.07
ME-7550 776146 8524462 <0.01   ME-7587 776341 8524615 0.02
ME-7751 776022 8523290 0.23   ME-7588 775830 8524703 0.22
ME-7752 776022 8523290 0.01   ME-7589 775921 8525300 0.06
ME-7753 775313 8524384 0.01   ME-7590 775144 8525250 0.02
ME-7554 775082 8524746 0.03   ME-7591 775532 8525101 0.01
ME-7555 775190 8524889 0.02   ME-7592 775563 8524824 0.02
ME-7557 776147 8524123 0.02   ME-7593 775560 8524388 0.29
ME-7758 776174 8524636 0.02   ME-7594 775879 8524503 0.09
ME-7559 776175 8525182 <0.01   ME-7595 776092 8524223 0.02
          ME-7596 776343 8523695 0.07

TABLE 9-6 MT ELLISON EAST ROCK CHIP AU RESULTS IN PPM

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9.5.5      VTEM ANOMALIES BLT 18, 19, 20, 21, 23, 24, 26.

These are all good quality late time, short strike length conductors that have no distinct magnetic signatures. They are possibly hosted in a regional syncline and geology mapping indicates Mt Bonnie and Burrell Creek Formations underlie the area. The area is overlain by what is now seen to be a thin layer of black hydromorphic soils that overlie residual soils. There are no distinct uranium or potassium anomalies associated with the conductors. They are quite possibly massive sulphide targets. Weak but questionable formational looking conductors to the west may be a Union Reefs type environment.

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Southern Geoscience (Card 2012) modeled several of the conductors and their interpretation is as follows:

9.5.5.1   TARGETS BLT_021 AND BLT_022

These two targets lie in Block 1 of the Burnside VTEM survey. They are identified as anomalous responses in the latest VTEM channels, and appear near the eastern edge of lines 10600 and 10610. Each of the responses appears primarily on one line with a lesser (off end) expression on the adjacent line. The labeled axes of these conductors are shown in Figure 9-35 below together with the flight path and late time response profiles superimposed on a magnetic first vertical derivative image.

A re-interpretation of the VTEM conductors when directly compared with the magnetic data indicates that anomaly BLT_21 may be an extension of conductor BLT_24. Conductor BLT_18 may be a strike extent of this anomalous trend, which seems to have an association with a distinct magnetic “low”. A strong conductor located at the very east end of line 10610 needs to be investigated.

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9.5.5.2   BLT_021

This target is characterized by a discrete, non-migrating response in a region of low magnetic response. The area does display some characteristics of polarized ground; however, this response does not appear to be attributable to this.

The VTEM responses shown in Figure 9-35 are the modeled response (red) and the observed response (black). Plan and section views of models BLT_021 and BLT_022 are presented in Figure 9-36.

The modeled conductor is 500S and dips at around 40° to the southeast. The model has a strike length of 215m at Azimuth 027°. It is only 40m in depth extent; so careful drill targeting will be necessary. Full model parameters are supplied in Table 9-7.

It is recommended to target this plate along line 10600, as this is where data directly over the conductor, is available making this the best-constrained part of the plate. It is also recommended that the plate be drilled 1/4 to 1/3 of the way down its depth extent. Plate parameters are provided below. The suggested intersection target for drilling is at:

9.5.5.3   BLT_022

This target is characterized by a discrete, non-migrating response in a region of low magnetic response. The area does display some characteristics of polarized ground; however, this response does not appear to be attributable to this.

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The modeled conductor is 200S and dips at around 35° to the northwest. The model has a strike length of 215m at Azimuth 017°. It is only 55m in depth extent, so careful drill targeting will be necessary. Full model parameters are supplied in Table 9-7.

It is recommended to target this plate along line 10610, as this is where there is data directly over the conductor, making this the best-constrained part of the plate. It is also recommended that the plate be drilled 1/4 to 1/3 of the way down its depth extent.

The suggested intersection target for drilling is at:

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9.5.5.4   TARGETS BLT_020 AND BLT_026

These two targets lie in block 1 of the Burnside VTEM survey. They are identified as anomalous responses in the latest VTEM channels, and appear near the eastern edge of lines 10470 and 10480. Each of the responses appears primarily on one line. The labeled axes of these conductors are shown in Figure 9-37 below together with the flight path and late time response profiles superimposed on a magnetic first vertical derivative image.

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9.5.5.5   BLT_020

This target is characterized by a discrete, non-migrating response in a region of low magnetic response.

The VTEM responses shown in Figure 9-38 and Figure 9-39 are the modeled response (red) and the observed response (black). Plan and section views of models BLT_020 and BLT_026 are presented in Figure 9-3740.

The modeled conductor is 450S and dips at around 70° to the west. The model has a depth extent of 30m and a strike length of 70m, but this is poorly constrained, as the response is only clearly evident on one line. Full model parameters are supplied inTable 9-8.

It is recommended to target this plate along line 10480, as this is where we have data directly over the conductor, making this the best-constrained part of the plate. It is also recommended that the plate be drilled 1/4 to 1/3 of the way down its depth extent. Plate parameters are provided in below. The suggested intersection target for drilling is at:

9.5.5.6   BLT_026

This target is characterized by a discrete, non-migrating VTEM response in a region of low magnetic response.

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The modeled conductor is 120S and dips at around 50° to the east. The model has a depth extent of 65m and a strike length of 80m, but this is poorly constrained, as the response is only clearly evident on one line. Full model parameters are supplied in Table 9-8.

It is recommended to target this plate along line 10470, as this is where there is data directly over the conductor, making this the best-constrained part of the plate. It is also recommended that the plate be drilled 1/4 to 1/3 of the way down its depth extent. Plate parameters are provided in Table 9-8 below.

The suggested intersection target for drilling is at:

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9.5.5.7   GEOLOGY

Most of the area is covered with unconsolidated Quaternary material in which the weakly magnetic dolerite represents the most prominent outcrops as ridges, dykes and scattered patches. The dolerite was not recorded in previous mapping projects.

Three different sets of dolerite dykes were identified in the field. In the south, a general E-W trend was observed. The largest dolerite outcrops were recorded in the central part of the area intermittently transitioning into NW-SE and E-W striking dykes. There were no coincident magnetic or EM anomalies associated with the dolerites but a close special relationship may exist. A dolerite ridge NE of anomaly BLT_023 might explain the local magnetic high. In the north, a dolerite swarm running N-NW-SSE was recorded close to BLT_026.

Vein quartz appeared as being opaque white with minor amounts of iron oxides and no sulphides (possibly oxidized sulphides). The quartz outcrop in the central part displays a stock work of four parallel vein sets with a total length of 25m and 50m across strike. This system appears to follow an E-W strike. The contact between veining and bedrock was recorded further to the east where strongly silicified meta-sediments are bedding concordant indicating the presence of at least two different vein sets in the area.

9.5.5.8   SOIL SAMPLE SURVEY

Initial interpretations of the overburden cover indicated that the area as covered with significant thicknesses of black soils. These do not react well to standard soils sampling and analytical techniques so it was decided to attempt one of the more innovative geochemical means of seeing through exotic overburden. ALS Chemex’s Ionic Leach was chosen.

This sodium cyanide leach is buffered to pH 8.5 using the chelating agents ammonium chloride, citric acid and EDTA. Ionic Leach enables the detection of buried mineralization through the dissolution and subsequent measurement of weakly bound ions loosely attached to surface particles. The ability of this innovative leach to give close to true background detection limits makes it effective in providing geochemical contrast and targeting buried mineralization. The method requires 50 grams of sample. Instrumental analysis is carried out using ICP-Mass Spectrometry.

The application of ionic geochemistry in deeply weathered lateritic terrains for gold exploration, has seen significant development over the last 12 years as scientific understanding, sampling requirements, extractant solution chemistry and instrumentation have all made significant progress. As the knowledge has developed, so to interpretation of the data has improved providing not simply single element target anomalies, but multi-element patterns and associations linked to specific mineralization styles and their settings. In landforms dominated by lateritic regolith units, either in-situ, partially dismantled or transported, element associations become increasingly significant for interpretation of the data.

Selective leach geochemistry seeks to extract only the mobile portion of metals from a soil sample. Typical soils contain high concentrations of subsurface metals (i.e., that are an intrinsic component of the soil) and since many soil parent materials have been transported the endogenic metal signal has no chemical relationship to that of the underling mineralization. Selective leach techniques are very weak extractions that attempt to selectively dissolve the ‘exogenic’, or mobile component of metals that stick to the outside of the mineral grain. Since metals originating from mineralization buried by exotic overburden are most likely to be part of the mobile component, selective leaches are more successful at detecting a signal from mineralization than are conventional strong acid digestions that dissolve most of the soil sample.

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The acidification of overburden over massive sulphide deposits can produced strong halo anomalies in elements such as Ca, Mg, Sr, Cl and Br. Zinc-Copper-Lead mineralization in drier terrain produces wide, apical, commodity element responses in Zn, Cd and Pb but notably not in Cu.

In the VTEM anomaly areas a normal soil profile was observed with an approx. 10cm lacustrian clay top layer, which was underlain by what was interpreted to be residual soils. Selective lines over all the strong VTEM anomalies were soil sampled. A total of 384 samples were collected. Fifteen of these were duplicates taken for QA/QC purposes.

Initial interpretations indicate that the sample duplicates for most elements correlated quite well and within reasonable limits. However, the gold results in some instances display a significant amount of variability. Possibly the low detection limits of the samples selected for duplicate analysis plays a role.

Soil samples taken at the southeast end of the grid are anomalously high in a number of elements including Ag, Mg, Ca, Au, Cu and Ba, and anomalously low in Pb. It is suspected that the proximity to a north south flowing stream immediately to the east implies that there is contamination from overbank sediments. Why the stream sediments are anomalous in multiple elements remains to be determined.

With respect to the remainder of the VTEM anomalies no particular element clearly defines any individual conductor. There are implications that some elements display anomalism in close proximity to some conductor axis but not necessarily along its entire length. One exception is the conductor at the east end of line 10610 which does appear to exhibit anomalies in Ag, Pb, Ba, Cu, Mg and Ca.

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9.5.6      SNAKEBITE

The Snakebite anomaly is situated between latitudes 13°24’S and 13°25’S and longitudes 131°32’26E and 131°32’49E. The anomaly is situated within EL25748 and is located approximately 130km southeast of Darwin, and 55km northwest of the Pine Creek Township. The area is situated within Perpetual Pastoral Lease No. 1111, held by Ban Springs Station Pty Ltd.

EL25748 surrounds the Burnside Granite, with the Snakebite prospect lying immediately southwest of the Woolwonga Pit by 700m. Snakebite is identifiable as being a conspicuous ridge in an otherwise low lying terrain; and straddles the Margret River and its floodplain (Figure 9-42).

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Snakebite is located in close proximity to the previously producing Woolwonga mine, within the Burnside deposits area. Limited historical and modern-day exploration has been conducted at Snakebite. Evidence of historical workings exists along the Snakebite ridge as two small 1.5m depth excavations.

Snakebite lies within the northern extension of the Pine Creek Shear Zone, located between the Glencoe and Woolwonga anticlines. Lithology is dominated by the greywackes, siltstones and shales of the Mount Bonnie Formation. Extensive hydrothermal alteration is evident as sericite and chlorite in quartz veins, stockwork and surrounding sediments (Masterman 2007).

Lithologies of the Snakebite deposit are dominated by greywackes, siltstones and shales of the Mount Bonnie Formation. Mineralization is identified as series of approximately 10cm width quartz veining. Structural analysis has defined quartz veins striking 330-350° and dipping either vertically or towards the northeast. Extensive hydrothermal alteration is present within mineralized zones, characterized as sercite and chlorite in quartz veins, stockwork and surrounding sediments. Pyrite, arsenopyrite and galena are visible within alteration zones. The extent and width of mineralization is unknown, however, the ridge is 850m in length and 350m in width (Masterman 2007).

In 2007 GBS Gold Australia Pty Ltd conducted a rock chip sampling program inclusive of 13 samples; the samples were collected along the center of the ridge. The average grade of the rock chip samples was 0.7g/t Au; however, with individual samples recorded results as high as 3.14g/t Au. Following the results of the rock chip program a soil sampling program was also proposed. The soil sampling program consisted of 119 samples and was also carried out in 2007 by GBS. The soil sampling program consisted of collecting samples every 50m along 50m traverses, within a 350m by 850m grid containing the ridge. The soil sampling program provided promising results, inclusive of 5,090ppb Au and 935ppb Au. Overall the sampling program results provided an average of 0.12g/t Au. The soil sampling program also included analysis for a suite of other elements including As, Cu, Pb and Zn. Gold results obtained from the soil sampling were used to create a detailed geochemical map for the Snakebite ridge. The geochemical map provided insight into the possible orientation of the Snakebite gold anomaly. Based on the results from the geochemical sampling an RC drilling program was proposed for 10 holes, each drilling to a provisional depth of 100m, totaling 1,000m of RC drilling. The RC program was not carried out and an interpretation of the geochemical data and map was not conducted.

The basis for exploration interpretation was the geochemical soil sampling. The objective was to gauge whether there were any underlying relationships between gold (Au) and arsenic (As), copper (Cu), lead (Pb) or zinc (Zn). Table 9-9 details the correlation matrix between the suites of elements. Iron and zinc were the only elements to show any significant distribution similarities.

  Au ppb As ppm Cu ppm Pb ppm Zn ppb
Au ppb 1        
As ppm -0.002 1      
Cu ppm 0.06 -0.07 1    
Pb ppm -0.11 0.13 0.17 1  
Zn ppb -0.10 0.04 0.56 0.68 1

TABLE 9-9 SNAKEBITE - SOIL SAMPLING CORRELATION MATRIX

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The geochemical map previously created is the basis for the proposed RC drill program. Manipulations of the soils data were conducted to determine if any other elements provided further insight into the possible orientation of the anomaly, no identifiable relationships were determined. Basic map interpretation was conducted with the results correlating to the position of the proposed drill holes, indicating a possible northwest-orientated gold rich zone.

9.5.6.1   GEOPHYSICS

A merging of Crocodile Gold’s VTEM data with past operators close line spacing aeromagnetic / radiometric data indicates that there are no VTEM conductors associated with the Snakebite occurrence although a weak surficial looking anomaly 1.0km to the west northwest should be investigated.

The aeromagnetic data would indicate that the Snakebite area is underlain Mt Bonnie Fm (possibly in contact with Burrell Creek Fm) with distinct magnetic horizons that are interpreted to be folded. The occurrence is in all likelihood to be within the hangingwall influence of the regional Pine Creek Shear Zone. Radiometric data indicates a weak but distinct potassium anomaly in what could be interpreted as a fold nose trending N-NW.

9.5.7      NORTH CULLEN VTEM TARGETS

This area was named due to the location compared to the regional Cullen Granite. Due to a lack of historic exploration drilling and thicker soil cover it was decided to cover this area with the VTEM survey to try and identify prospective targets.

Seven late time VTEM targets were identified for follow up. Unfortunately, by the end of the field season this area had only been visited with several rock chip samples collected. No mapping was conducted but several quartz veins were noted close to surface. These were sampled with one returning results of +1g/t gold. This suggests follow up work is required in this area.

A summary of the seven VTEM targets as noted by Card (Card 2012)) are noted below: -

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The majority of the VTEM conductors are interpreted to be hosted by Koolpin Fm. Cosmo type mineralization is expected, which in all likelihood will consist of graphite and pyrrhotite, both of which are highly conductive.

Results of the rock chip sampling can be seen in Table 9-11 below. It has been noted that two samples from surface gave sub-economic to economic gold grades, which require further testing and mapping to identify the potential in this area

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SampleID Type North East Au
NBURK2077 Rock  8528666      762919 -0.20
NBURK2078 Rock  8528667      762920 -0.20
NBURK2079 Rock  8528728      763044 -0.20
NBURK2080 Rock  8528728      763044 -0.20
NBURK2081 Rock  8528762      763104 -0.20
NBURK2082 Rock  8528775      763098 0.50
NBURK2083 Rock  8528953      763438 -0.20
NBURK2084 Rock  8528933      763442 -0.20
NBURK2085 Rock  8529104      763088 -0.20
NBURK2086 Rock  8529104      762906 -0.20
NBURK2087 Rock  8529140      762782 -0.20
NBURK2088 Rock  8528752      759910 -0.20
NBURK2089 Rock  8528589      759773 1.40
NBURK2090 Rock  8528724      759800 -0.20

TABLE 9-10 NORTH CULLEN - RESULTS OF ROCK CHIP SAMPLING, AU G/T

9.5.8      JENKINS AREA VTEM TARGETS

Jenkins is a previously un-named prospect approximated 4.5km to the east of the Iron Blow deposit on EL23540. A review of this work is outline below and was taken from an internal report, which has been reviewed by the Authors and agreed with the outcomes by Jenkins (Jenkins 2012)

The VTEM anomalies BLT 114, 154, 155 and 156 are located approximately 4.5km east of the Iron Blow deposit and about 2.2km south of Goldfields Rd. The area can be accessed by a track that heads south from Goldfields Rd, 5km east of Grove Hill. This track leads to within 500m of the eastern part of the conductor axes. The anomalies are in the exploration lease EL23540.

Previous exploration conducted in the area has targeted Mount Bonnie/Iron Blow style stratiform Au and base metal mineralization, Cosmo/Golden Dyke hydrothermal Au mineralization and Yam Creek/Union Reefs style anticline hosted Au mineralization.

A soil sampling program conducted in 2006 by GBS Gold returned some significant results including 1,330ppb Au within 250m of the conductor BLT154 and 111ppb within 50m of BLT 154. Lines were spaced 800m apart and samples spaced 50m apart along the lines.

In 2011 a diamond drill hole, MBEXD001, was completed by Crocodile Gold approximately 1.2km to the west of conductor BLT154. Only three samples from the diamond hole returned results above the lower detection limit of 0.01g/t and these were 0.03g/t or less. According to the 1987 1:100K Geology map of the area the hole was drilled in an area dominated by the Gerowie Tuff and Zamu Dolerite.

A 1:10K scale geology outcrop map of the area has been created.

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The conductor axis BLT114 is short, about 150m in length, that occurs just to the north of the much longer BLT154. The axis follows a ridge comprised of silicified meta-sediments, on the western side a large dolerite sill. There was no outcrop found on the eastern side of the ridge. Five rock samples of meta-sediments, quartz and laterite at the base of the ridge were collected

An area of alluvial workings was observed along one of the creeks in the field area. One small excavation was also observed just up slope of the alluvial workings. Another small possible alluvial excavation was observed along the northeastern portion of the anticline.

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The conductor axis BLT154 has been identified as the most likely to be related to gold mineralization because of its geometry and spatial relationship with anomalous soil sample results. This axis, along with the others should be further examined

It is recommended that mineralization be further tested in the target area with a program of soil sampling. Sample lines should be un-evenly spaced between 100 and 400m apart, depending on previous sampling results and field observations. Samples should be spaced 25m apart along lines. This program would have a total of 532 samples. Once collected the samples should be dispatched for a multi element analysis.

The areas further south of the soil program should be tested and this could be done using stream sediment sampling. Several small creeks cut the conductor axis that extends further south.

The geology and structures mapped and observed in the field resemble a Cosmo/Golden Dyke style deposit. The VTEM conductor axis seems to follow bedding and lines up with anomalous gold assays in several places. These factors make BLT 114, 154, 155 and 156 promising exploration targets.

9.5.9      BAN BAN AREA

In late 2011, with the results of the VTEM survey available, a number of targets were noted in a new prospective area, which was called Ban Ban after the local station.

The VTEM survey and analysis had noted 5 significant late time targets in this area, (BLT 74, 77, 78, 80 and 178). These targets were recommended for further investigation due to the similarity to the nearby Mt Ellison deposit, which was a high grade/low tonnage copper mine that produced in the early 1900’s.

The summary of the Ban Ban targets are noted below;

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Published geological maps indicate Koolpin Fm along with Zamu Dolerite underlies the majority of the area. Lower sequences of the Gerowie Tuff and Mt Bonnie Fm are likely also present. Of prime importance is that the regional Pine Creek Shear Zone traverses the area in a NW-SE direction immediately south of VTEM conductor BLT_076. The geological and structural settings are very intriguing for Cosmo Mine type gold deposits. Additional work is recommended.

9.5.10      BON'S RUSH DEPOSIT

Northern Gold defined the Bon’s Rush deposit in the 1990’s when they tested a bedrock gold anomaly (through RAB drilling of a gold in soil anomaly) 900m long and 180m wide. Northern Gold subsequently drilled 20RC holes on 100m spaced section lines delineating significant near-surface high-grade gold over a strike extent of at least 400m. Two diamond drill holes were drilled to determine the orientation and controls on the mineralization and confirmed the results from the RC drilling.

In 2008 Northern Gold calculated an Inferred mineral resource of 540,000t at a grade of 2.51g/t Au (43,300 oz/Au). Crocodile Gold (F. E. Muller 2011)reported an Inferred mineral resource at Bon’s Rush, with a 0.7g/t Au lower cut-off, of 805,000t at a grade of 2.3g/t Au (60,400oz gold). (NB1)

NB: (1) The mineral resource estimate cited is sourced from the reference indicated, is believed to be a historical estimate, not prepared in accordance with currently accepted guidelines for the preparation of mineral resources and mineral reserves, may not comply with NI43-101 and is not considered by either the Authors or Newmarket Gold, as current mineral resources or mineral reserves, as the Authors have not done sufficient work to classify historical estimates as current mineral resources or mineral reserves.

Crocodile Gold flew a VTEM survey over the Bon’s Rush area in 2011. Line spacing was 150m Survey results indicate that a moderate strength VTEM anomaly clearly defines the mineralization at the Bon’s Rush deposit, paralleling the projected to surface mineralized shear and possibly delineating additional mineralization to the north and south of the currently known drill defined mineralization.

High-grade gold in bedrock anomalies have been identified by Northern Gold’s RAB drilling along and around the western limb, fold nose and eastern limb of a parasitic fold of the Howley Anticline. This RAB defined gold in bedrock anomaly is 1,000m long and 100m wide. Subsequent limited RC drilling confirmed and identified significant zones of gold mineralization within sheared Zamu Dolerite. Higher-grade zones have the same association with quartz-carbonate veining, pyrite, arsenopyrite and pyrrhotite similar to the mineralization at the Bon’s Rush deposit.

A second 500m long, northeast trending, gold in bedrock anomaly was identified by the Northern Gold RAB drilling located immediately east of the western limb mineralization, within a similar structural setting. The mineralization is located within a Zamu Dolerite sill stratigraphically below the Upper Zamu Dolerite. A moderate VTEM conductor to the NW and a strong VTEM conductor to the SE bracket this zone. Ground investigations by Crocodile Gold indicated there were no outcrops located throughout this area.

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Results from the Northern Gold RC drilling indicate the potential for delineating high-grade, north plunging lodes in and around the fold hinge analogous to mineralization at the Bridge Creek deposit to the south.

On the Bon’s Rush East Limb high-grade gold in bedrock mineralization has been identified by Northern Gold’s RAB drilling on the Eastern Limb of the anticline within Zamu Dolerite over a strike length of 200m that is open to the north. There has been no RC drilling follow-up completed. A weak to moderate VTEM anomaly is coincident with this zone of potential gold mineralization.

At Bon’s Rush South, Northern Gold defined a base metal intersection of 1.634% Pb and 0.145% Zn over 30m in RAB hole BRRB-380, hosted by the dolerite stratigraphically below the Bon’s Rush Dolerite sill. Northern Gold reported additional Pb/Zn dolerite hosted base metals outlined by previous explorers located 800m to the north within the same dolerite sill.

Significant gold in siltstone/graphitic shale, analogous to the Bridge Creek mineralization, was intersected 70m to the east in old FSDC RC holes (FSDC-057, 58 and FSDC-059) along with weak gold mineralization along a dolerite/siltstone contact. Northern Gold never followed up this area. This mineralization occurs immediately south of the strong VTEM anomaly BLT-173.

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9.5.11      BURNSIDE AREA CONCLUSIONS AND RECOMMENDATIONS

The Bon’s Rush deposit currently has an inferred Mineral resource. Numerous detailed maps (Geological, Drill holes locations and results, Au and As in soil analytical results, structural interpretation) are available in the database. This data should be compiled, which along with the newly acquired VTEM data, would serve to guide future exploration programs. Additional detailed geological mapping or soil sampling would not be fruitful, as 1:5000 scale programs have already been completed by previous operators. Following up on the VTEM conductors should be completed and grab samples collected if warranted.

Following are additional recommendations presented in point form:

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10 DRILLING

The drilling completed for Newmarket Gold can be broken into two main categories;

Diamond – Traditionally NQ (47.6mm) core is used for diamond drilling, however, at times this is increased to HQ (63.5mm) to assist with sample quality. The majority of underground diamond drilling is NQ while the majority of surface drilling is HQ.

RC –Reverse Circulation (RC) drilling is used as it can be a quicker and cheaper alternative to diamond drilling. This drilling was used in areas where diamond drilling was not required or appropriate. The RC drilling at Cosmo, for example, was designed to test the potential mineralization to the south of the Cosmo pit where limited drilling was identified. RC Drilling at Cosmo used one contractor who had a cone splitter attached to the rig. Generally RC drilling procedures were used including having a staff member on the rig at all times to maintain sample security and quality.

10.1 COSMO MINE DRILLING

10.1.1      DIAMOND DRILLING

Crocodile Gold/Newmarket Gold conducted a significant amount of drilling over the past five years, mainly targeting the expansion of mineral resources and the development of the mine.

Capitalized mineral resource definition (non-grade control) diamond drilling has been the focus at Cosmo over the past four years for a number of reasons;

During 2015, Newmarket Gold embarked on the first surface diamond drilling program at Cosmo Mine since 2012. This drilling was designed to test the down plunge targets of the mineralized system to allow for better information in the designing of underground infrastructure and drilling. This program has not defined any new mineral resources due to the amount of step out (over 100m from deepest drilling in this part of the deposit), but has identified the geometry of the mineralization. This program was not complete as of December 31, 2015 but only those meters completed at this time have been included in this summary.

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Deposit Diamond First Drilled Last Drilled
Holes Meters Size
Cosmo Surface 13 5,358 NQ-HQ 19-Jun-11 2-Dec-12
Cosmo Underground 360 46,632 NQ-HQ 1-Jan-11 31- Dec-12
Cosmo Underground 142 22,693 NQ-HQ 1-Jan-13 31- Dec-13
Cosmo Underground 264 40,057 NQ-HQ 1-Jan-14 31- Dec-14
Cosmo Surface 3 2,174 NQ 7-Oct-15 31- Dec-15
Cosmo Underground 231 46,820 NQ-HQ 1-Jan-15 31- Dec-15
Total 1,013 163,734      

TABLE 10-1 DIAMOND DRILL STATISTICS FOR THE COSMO MINE

Specific drill programs are summarized below:

Several exploration diamond drill programs have occurred on the deposit over the past five years. For more information on the drilling that has occurred prior to 2015 please see (Gillman, et al. 2009) (Smith and Pridmore 2014). The information below is a summary of the exploration drilling conducted during 2015 at the Cosmo Mine.

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10.1.1.1   300 Lode Target – Southern Cosmo Mine:

This target is the downward, and southern extensions to the 300 Lode adjacent to the F8 and F8a Faults. Results proved to be of lower grade or thickness and only a minimal amount of additional mineral resource was added from this drill program. No firther drilling is planned on this target.

10.1.1.2   Western Lode Deeps:

This target was developed in 2013-14 as a possible fold repetition of the Eastern Footwall Lodes on the northwestern flank of the Cosmo Anticline. Existing surface drill results in CP003W1 of 11.1g/t Au over 4.8m and 2.5g/t Au over 9.5m was followed up in late 2014 with two successful holes CW69001 (3.5g/t Au over 20.7m) and CW69002 (2.6g/t Au over 6.6m) completed (Smith and Edwards 2015).

Gold intersections were within the Pgt/Pgtb greywacke host sequence close to the graphitic shale Pmc contact as similarly known for the 100 Lode mineralization in the Eastern Lodes, and the validity of this target was also confirmed by the (J. Miller 2014)structural review (Target 1 in Figure 9-1).

During 2015 further underground diamond drilling phased programs were completed on the Western Lode target to scope for additional near mine mineral resources. Significant drill results for the Western Lodes Target included 6.59g/t Au over 6.4m in hole CW69006 and 4.54g/t Au over 11.85m in hole CW69009, with most gold intersections within approximately 160m of current underground infrastructure.

Additional close spaced underground drilling is planned in 2016 after the newly developed extension to the 640 drill drive is completed. A small number of holes are designed to test the main high-grade plunging gold shoot identified from structural studies in late 2015 (Beeson 2015) and those outlined in section 1.6.1. With success, a 50m x 25m spacing would be achieved for subsequent mineral resource assessment.

10.1.1.3   Cosmo 2200N Surface Drill Section:

With the down plunge continuation of the Sliver mineral resource, a four hole surface drill section began in October 2015 to provide new geological information and mineralization results on the 2200mN section approximately 160m north of the most northern previous Sliver underground drill intersection. The program was completed in January 2016, generating the following outcomes;

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One hole on the section was drilled steeper so as to pass through the F1 Fault and test the Eastern Footwall Lodes orthogonal to a small number of very successful underground holes. These had established that the Eastern Lodes continued north, but with more gold mineralization found in the 300 Lode than the 100 and 200 Lodes, stratigraphy closer to the eastern Pmc contact.

Whether the better gold assay intersections are found on the 300 Lode due to an as yet un-modelled NW cross fault remains unknown. There may be potential for further areas to the north to see the 100 and 200 Lodes return to the stronger gold grades typical of the Cosmo Mine.

There is an emerging interpretation that the positive gold results to the west of the 300 Lode in the Eastern Lode Footwall Deeps may be folded portions of the 100, 200, 300, or 400 Lodes and thus represent mineralization in the Cosmo Fold Hinge below the F1 Fault (Figure 10-2).

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10.1.2      RC DRILLING

Also during 2015, a small surface exploration program was completed to the south of the main Cosmo open pit. This program was designed to target small oxide mineralization as defined in a structural study. Generally RC drilling at Cosmo Mine has been limited due to the depth of potential mineralization.

Deposit RC First Drilled Last Drilled
Holes Meters Size
Cosmo Surface 3 249 5" 19-Jun-11 2-Dec-12
Cosmo South Surface 29 2,720 5" 1-Jun-15 30-Sep-15
Total 32 2,969      

TABLE 10-2 RC DRILL STATISTICS FOR THE COSMO MINE

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10.1.3      HISTORIC DRILLING

Within the historic databases that Newmarket Gold has inherited since operations began there are several different sets of data that are available for the Cosmo Mine.

All of this data has been reviewed and entered into a database for use in model updates. Within the data that has been made available are some of the original QA/QC data as well as the original hard copy assay files. These have been audited by Company staff prior to entering the drilling data into the database. Several thousand meters of RC and diamond drilling is available within the Newmarket Gold database. This data is captured by drilling type, assay method, QA/QC, drill date, project and other relevant data types. Some of these are:

Diamond Drilling 
Company Number Meters
DML 133 13,793
NGNL 20 10,034
GBS GOLD 10 4,380
BMR 15 1,755
GEOPEKO 23 8,771
HOMESTAKE 30 3,587
Total 231 42,321

TABLE 10-3 HISTORIC DRILLING BY COMPANY – DIAMOND DRILLING

RC Drilling
Company Number Meters
NGNL 66 4,399
GBS GOLD 22 2,224
DML 503 39,235
BOPL 1 78
CROCGOLD 62 6,097
GEOPEKO 1 241
Total 655 52,274

TABLE 10-4 HISTORIC DRILLING BY COMPANY – RC DRILLING

  Diamond Drilling  
Year Number Meters   Year Number Meters
Pre-1977 17 2,961   1993 2 1,548
1977 15 1,847   2004 12 6,433
1978 6 1,070   2005 8 3,601
1980 8 630   2006 10 4,380
1981 1 40   2009 2 1,851
1982 4 2,359   2010 16 10,732
1983 15 3,658   2011 9 4,322
1985 62 5,976   2012 351 42,310

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Diamond Drilling
Year Number Meters   Year Number Meters
1987 34 3,132   2013 142 22,693
1988 28 2,822   2014 264 40,057
1989 3 361   2015 231 46,820
1991 3 403   Total 1,246 211,106
1992 3 1,099        

TABLE 10-5 COSMO MINE HISTORIC DRILLING BY YEAR – DIAMOND DRILLING

The historic drilling is all surface drilling, a combination of RC and diamond holes. The diamond holes have provided the most valuable data to use as most of the drilling has focused on areas beneath the Cosmo open pit. The historic drill holes (not prefix “CP”) have been downhole surveyed using single shot and multi-shot cameras. When combined with the recent underground diamond drilling the exact location in some cases, is not accurate based on the geology correlation. A total of 24 diamond holes, due to the downhole location inaccuracies have been removed from the estimation.

The historic holes are also high angle to the mineralization lodes, in some cases +75°, which may cause some bias with the estimation process. This issue has been resolved somewhat by ensuring that underground diamond definition drilling targets these areas as part of the infill drill program. A lot of the historic holes including the Crocodile Gold drilled “CP prefix holes” are drilled at an azimuth that does not intersect the mineralization perpendicular to the strike of mineralization. This problem is exacerbated even more with the high angle drilling, which has caused issues with interpreting mineralization continuity in the fold hinge.

With the greater geological understanding of the deposit, over 80% of the historic diamond drill holes that could be found and the more recent Crocodile Gold “CP prefix” holes were re-logged. This program was instigated to ensure lithological continuity when interpreting. This re-logging has been a vital process when modeling the mineralization at depth and particularly on the hangingwall of the Eastern Limb when combining the data with the underground definition holes.

For more information about geological interpretation please refer to Section 14.

No drilling, sampling or recovery factors have been noted (other than currently noted in this technical report), which could materially impact on the accuracy and reliability of the results.

10.2 UNION REEFS DRILLING

The majority of recent drilling at the Union Reefs deposit was completed during 2011 and early 2012. However, an RC and Diamond drilling campaign was completed at the Esmeralda deposit during 2015, which is summarized below.

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10.2.1      DIAMOND DRILLING

Generally diamond drilling was used to test for underground mineralization at Prospect and Crosscourse deposits. Some small drilling campaigns were also completed for other smaller deposits in the Union Reefs area during 2011-12.

In 2015 a small diamond drilling program was completed at Esmeralda, mainly for the geotechnical purposes, however, once logging was complete these holes were sampled and used in the mineral resource estimation. A total of eight holes were drilled at Esmeralda between November and December, 2015.

Deposit Diamond First Drilled Last Drilled
Holes Meters Size
Union Reefs 91 22,392 NQ-HQ 27-Jan-11 08-May- 12
Esmeralda 8 567 HQ 15-Nov-15 15- Dec-15

TABLE 10-6 DIAMOND DRILL STATISTICS FOR UNION REEFS AREA

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Hole ID Drill Type Deposit Grid Coordinates Survey Data Interval Grade
Local
Easting
Local
Northing
Azimuth Dip From
(m)
To
(m)
Interval
(m)
True
Width
(m)
(g/t
Au)
URNDD21 Diamond Prospect 4875.0 7299.8 265.3 -58 153.4 162.2 8.8 5.7 3.5
including   158.9 159.4 1.5 1.0 7.4
including   161.4 162.2 0.75 0.5 24.4
URNRC22 RC Alta 4,828 8,664 270 -60 15 17 2   1.3
URNRC23 RC Alta 4,811 8,660 270 -60 35 37 2   0.8
URNRC24 RC Alta 4,789 8,660 270 -60 34 38 4   1.6
URNRC27 RC Alta 4,852 8,575 270 -60 2 4 2   0.6
URNRC31 RC Alta 4,846 8,529 260 -60 no intercept
URNRC32 RC Alta 4,830 8,525 270 -60 86 88 3   1.6
URNRC33 RC Alta 4,810 8,525 270 -60 3 6 3   2.3
and   28 33 5   1.1
URNRC36 RC Prospect 4,800 7,706 270 -60 0 3 3   0.9
and   9 11 2   1.2
and   48 50 2   0.6
URNRC37 RC Prospect 4,823 7,708 270 -60 0 3 3   0.8
and   34 38 4   4.9
and   44 48 4   1.6
URNRC38 RC Prospect 4,840 7,710 270 -60 23 32 9   1.1
URNRC49 RC Crosscour 5,010 7,084 90 -60 no intercept
URNRC50 RC Crosscour 5,031 7,083 90 -60 no intercept
URNRC53 RC Crosscour 5,065 6,992 270 -60 65 70 5   1.7
and   82 94 12   1.0

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Hole ID Drill Type Deposit  Grid Coordinates Survey Data Interval Grade
Local
Easting
Local
Northing
Azimuth Dip From
(m)
To
(m)
Interval
(m)
True
Width
(m)
(g/t
Au)
URSRC1 RC Millars  5,021 5,179 270 -60 100 103 3   0.9
and   106 108 2   1.5
URSRC5 RC Millars  4,900 5,080 90 -60 75 77 2   1.6
and   83 85 2   0.5
URSRC6 RC Millars  4,901 5,060 90 -60 88 94 6   0.5
URSRC7 RC Temple  4,860 5,300 270 -60 57 59 2   1.8
URSRC10 RC Millars  5,037 5,320 270 -60 no intercept
URSRC13 RC Millars  5,020 5,220 270 -60 87 89 2   0.8
URSRC13A RC Millars  5,020 5,220 270 -60 114 118 4   1.1
and   129 131 2   0.6
URSRC32 RC Crosscour  4,901 5,972 270 -60 92 104 12   1.1
and   113 116 3   2.1
URSRC34 RC Crosscour  4,873 5,969 90 -60 1 3 2   0.8
and   31 33 2   2.3
and   46 48 2   0.5
and   51 58 7   0.6
and   74 76 2   0.5
and   85 90 5   2.2
and   96 98 2   0.9

TABLE 10-7 SUMMARY OF 2011-12 DRILLING AT UNION REEFS AREA

Hole ID Location Local Grid Depth From To Width  True Width   Grade
East North RL (m) (m) (m) (m) (m) (g/t Au)
URNDD15 Millars/Big Tree 5056 5599 1194 225.0 143.4 144.6 1.3 0.9 9.4
and 159.0 160.2 1.2 0.9 27.8
and 176.4 179.6 3.2 2.4 1.4
URSDD12 Millars/Big Tree 5037 5300 1187 195.0 149.4 151.2 1.8 1.4 1.1
and 179.0 183.4 4.4 3.3 1.0
URSDD14 Millars/Big Tree 5031 5560 1195 155.3 77.6 79.9 2.4 1.8 7.5
Incl.           1.7 1.3 10.3
and       86.4 90.4 4.0 3.0 1.4
and       112.4 124.6 12.2 9.1 1.1
and       128.6 132.5 3.9 2.9 1.8
URNDD03 Cross- course 5125 6800 1195 248.8 244.5 248.4 3.9 2.9 2.5
Incl.           1.5 1.1 5.4
URSDD17 Cross- course 5067 5900 1196 201.1 129.8 134.5 4.7 3.5 2.2
and         155.3 157.0 1.7 1.3 4.5
and         163.1 164.6 1.5 1.1 1.1

TABLE 10-8 SUMMARY OF 2011-12 DRILLING AT UNION REEFS AREA

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The results as outlined in the tables above represents some of the first pass drilling completed at Union Reefs to test mineralization below historic pits. Little or no historic drilling has been completed to great depths excluding drilling beneath the historic Crosscourse Pit. The recent drilling was designed to determine if there is potential for deeper deposits for underground mining or potentially larger scale open pit mines.

1.

Drill samples were assayed at Northern Australian Laboratories (NAL).

   
2.

Assay results are based on 50g fire assays.

   
3.

All intersections from diamond samples are HQ half-core.

   
4.

Diamond sample intervals are determined using lithological controls or 1.0m, whichever is larger.

   
5.

Mean grades have been calculated on a minimum of 1.0m interval, 0.5g/t Au lower cut -off and maximum 3m internal dilution.

   
6.

All intersections are down-hole intervals.

   
7.

All deviations have been verified by down hold camera.

   
8.

QA/QC for these holes has been checked and verified through blind standards, field duplicates, lab repeats and barren flushes. QA/QC results are within expected limits.


Hole_ID Location Local Grid Depth
(m)
From
(m)
To
(m)
Width
(m)
True
Width
(m)
Grade
(g/t Au)
East    North  RL
URNDD03b Crosscourse 5133.8    6797.8  1195.2 598.7 234.7 241.15 6.45 5.16 2.83
incl.           239 241.45 2.45 1.96 5.43
and           274 277.1 3.1 2.5 1.4
and           345.55 347.3 1.8 1.4 15.09
and           353.7 354.8 1.1 0.9 1.26
and           404 405.9 1.9 1.5 1.42
and           409.4 418.8 9.4 7.5 1.38
incl.           414.6 415.3 0.6 0.5 5.72
and           427.8 432.1 4.3 3.5 1.89
incl.           431.8 432.1 0.3 0.24 17.1
and           459.1 461.1 2 1.6 2.89
incl.           460.1 461.1 1 0.8 5.25
and           470.2 471.8 1.5 1.2 1.39
and           490 492 2 1.6 1.45
and           497.9 506.8 8.9 7.1 1.12
and           540.7 541.9 1.2 0.9 1.05
and           551.6 554.2 2.6 2.1 3.23
incl.           552.7 553.3 0.6 0.5 6.75
and           560.1 561.4 1.3 1 5.23
and           588 591.5 3.5 2.8 3.1
incl.           588 589.6 1.6 1.3 5.78
URNDD29 Prospect 4939.1    7268.1  1207.5 434.4 278.4 279.4 1 0.8 1.26
and           283.7 286.9 3.2 2.6 1.01
and           290 294.7 4.6 3.7 1.56
incl.           294.1 294.7 0.5 0.4 7.97
and           369.1 370.4 1.2 1 1.11
and           376 381 4.9 3.9 2.81
incl.           379 381 2 1.6 5.57
URNDD29W1 Prospect 4912.5      7268 1110.8 395.4 164.6 165.8 1.2 1 7.83
and           195.1 200.4 5.3 4.2 2.12
incl.           195.1 196.1 1 0.8 5.3

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Hole_ID Location Local Grid Depth
(m)
From
(m)
To
(m)
Width
(m)
True
Width
(m)
Grade
(g/t Au)
East North RL
URNDD30 Prospect 4927.1 7367.4 1208.6 494.4 408.3 409.7 1.4 1.1 2
and 415.7 418.8 3.1 2.5 4.83
incl. 415.7 418.1 2.4 1.9 6.04
URNDD30W1 Prospect 4904.4 7367 1114.7 292.8 219.5 226.8 7.3 5.84 3.99
incl. 223.7 226.8 3.1 2.5 7.7
URNDD28 Prospect 4928.9 7166.9 1191 472 371.65 374.93 3.28 2.62 12.51
incl. 373.36 373.9 0.54 0.43 65
and 425.78 427.83 2.05 1.64 27.96
incl. 426.8 427.11 0.31 0.25 165
URNDD28W1 Prospect 4909.8 7166.8 1093.1 289.6 196.6 197.6 1 0.8 21.98
incl. 196.9 197.3 0.4 0.32 52.3
and 254.44 259.65 5.21 4.17 3.38
incl. 259.35 259.65 0.3 0.24 20.4
URNDD34 Prospect 4946.6 7548.1 1199 276   No Significant Intercept      

TABLE 10-9 SUMMARY OF DRILL RESULTS FROM PROSPECT AND CROSSCOURSE DEPOSITS -2012

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10.2.2      RC DRILLING

Generally, RC drilling has been little utilized by Crocodile Gold/Newmarket Gold at Union Reefs due to the focus being on higher grade underground style of deposits. Some RC drilling was used for close to surface programs such as the drilling at the Orinoco deposit. However, in 2015 a large program of RC drilling was completed at the Esmeralda deposit, where over 5,000 meters was completed. This drilling has been used to update the mineral resource for the deposit, which is summarized in Section 14.

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Deposit RC First Drilled Last Drilled
Holes Meters Size
Union Reefs 49 4,434 5" Jan-27-11 May-08-12
Esmeralda 72 5,174 5” Oct-28- 15 Nov-30- 15

TABLE 10- 10 RC DRILL STATISTICS FOR UNION REEFS AREA

10.2.3      HISTORIC DRILLING

Within the historic databases that Newmarket Gold has inherited since operations began there are several different sets of data that are available for the Union Reefs deposit. This includes some of the grade control data used in the mining period by AngloGold.

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All of this data has been reviewed and entered into the Newmarket Gold database for use in model updates. Within the data that has been made available is some of the original QA/QC data as well as the original hard copy assay files. These have been audited by Crocodile Gold/Newmarket Gold staff prior to entering into the database.

Several thousand meters of RC and Diamond drilling is available within the Newmarket Gold database. This data is captured by drilling type, assay method, QA/QC, drill date, project and other relevant data types. Some of these are;

Diamond Drilling
Prospect Number Meters
Elizabeth 1 61
Esmeralda 5 361
Lady Alice 7 519
Orinoco 2 323
Prospect Claim 18 1,848
Union Reefs 13 2,069
Total 46 5,182

TABLE 10-11 HISTORIC DRILLING BY PROJECT – DIAMOND DRILLING

  RC Drilling   
Prospect Number Meters   Prospect Number Meters
Bungo 17 1,685   Orinoco 70 6,622
Caroline 34 2,720   Prospect 2,348 41,527
Culvain 5 402   Rosalie 6 582
Dam A 55 6,520   Snaddens Creek 20 1,394
Elizabeth 49 4,360   Tobermoray 14 1,176
Ennis 5 290   Tomsk 3 182
Esmeralda 149 11,230   Tomsk North 7 524
First Bite 15 976   Union Reefs 126 12,166
Great Uncle Bulgaria 15 1,306   Wellington 7 520
Lady Alice 246 27,504   Wimbledon 4 316
Lady Alice Contin 3 181   Total 3,204 122,770
Northern Belle 6 586        

TABLE 10-12 UNION REEFS HISTORIC DRILLING BY PROJECT – RC DRILLING

10.3

PINE CREEK AREA

10.3.1      DIAMOND DRILLING

Five holes were drilled at the International deposit in 2012. A total of 465.29m was drilled and is summarized in Table 10-13 below.

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Deposit  Diamond First Drilled Last Drilled
Holes Meters Size
International 5  465 NQ-HQ 30- Sep-12 31-Oct-12

TABLE 10-13 2012 DIAMOND DRILL STATISTICS FOR INTERNATIONAL DEPOSIT

There were several significant intercepts recorded from this drilling which are shown in Table 10-14 below. These intercepts show the number of mineralized lodes located in the deposit. These results highlighted the importance of re-interpreting the mineralization for a model up-date.

All holes were lithologically and geotechnically logged for use in future reporting.

Hole ID Intercept g/t Au
PCIDD_001 3.6m @ 1.33g/t from 17m  
PCIDD_001 7.66m @ 1.72g/t from 27.61m  
PCIDD_001 4.88m @ 0.77g/t from 39.07m  
PCIDD_001 4.87m @ 0.65g/t from 68.76m  
PCIDD_002 2.65m @ 0.62g/t from 12.8m  
PCIDD_002 15.45m @ 1.12g/t from 27.55m Incl. 1.92m @ 2.91g/t
PCIDD_002 5.4m @ 3.34g/t from 54m  
PCIDD_002 5.1m @ 1.02g/t from 75.1m  
PCIDD_002 7.59m @ 2.14g/t from 84.55m  
PCIDD_006A 16.88m @ 0.65g/t from 31.72m  
PCIDD_006A 10.8m @ 0.58g/t from 52.7m  
PCIDD_006A 2.53m @ 3.87g/t from 70.45m  
PCIDD_006A 5.56m @ 1.72g/t from 85.05m  
PCIDD_006A 3.38m @ 0.82g/t from 95.62m  
PCIDD_005 9.18m @ 0.77g/t from 30.12m  
PCIDD_005 1.07m @ 0.55g/t from 43.21m  
PCIDD_005 8.33m @ 0.96g/t from 49.03m Incl. 2.92m @ 1.66g/t
PCIDD_005 13.6m @ 1.67g/t from 65.36m Incl. 7.81m @ 2.42g/t

TABLE 10-14: SIGNIFICANT INTERCEPTS FROM INTERNATIONAL DRILLING

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10.3.2      RC DRILLING

In January 2011 RC holes were drilled into the International deposit to drill test historic drill results and to provide the Environmental team samples of the backfilled waste material. The drilling was successful in providing confidence in the historic drilling and in also providing environmental samples for future test work.

The details of this drilling can be seen in Table 10-15 below:

Deposit RC First Drilled Last Drilled
Holes Meters Size
International 2 152 5" 10-Jan- 11 13-Jan- 11

TABLE 10-15 RC DRILL STATISTICS FOR PINE CREEK AREA

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10.3.3      HISTORIC DRILLING

RC, open percussion (RAB) and diamond (DDH) drilling have been conducted at Pine Creek in the past. Over 50% of the holes drilled have been RC with a total of 79,694m. There were more RAB holes drilled than diamond holes.

A total of 2,491 holes were drilled in the Pine Creek Goldfields. The overwhelming majority of drilling was conducted between 1980 and 1994.

Burnside Operations drilled a total of 4,455m in 51 RC holes in the South Enterprise and Czarina Deposits in 2004/2005. This was the last drilling to have occurred at Pine Creek before Crocodile Gold took over ownership in 2009.

Below is a list of the prefixes, drilling methods and meters for the 2,491 holes drilled in the Pine Creek area.

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Period Drilling Type Number of Holes Meters Drilled Hole Prefix
Pre-1994 RC 138 7,015 GRC
Pre-1994 RC 110 6,497 IRC
Pre-1994 RC 8 458 MRC
Pre-1994 RC 49 3,372 PCPH1
Pre-1994 RC 1,027 59,674 PCRC
Pre-1994 RC 11 332 THRC
Pre-1994 DDH 31 2,879 GDD
Pre-1994 DDH 28 2,126 IDD
Pre-1994 DDH 332 32,870 PCDH
Pre-1994 RAB 66 2,727 B
Pre-1994 RAB 32 1,284 C
Pre-1994 RAB 116 5,311 CX
Pre-1994 RAB 29 928 E
Pre-1994 RAB 414 13,697 PCPH
2004 RC 16 1,694 CZRC
2004 RC 5 652 SERC
2005 RC 30 2,109 SERC

TABLE 10-16 HISTORIC DRILLING TYPES AT PINE CREEK

Wherever possible the details of the historic drilling were identified and noted in the database. This included trying to identify original files or details from governmental files. Drilling from 2000 onwards has all original files stored on site and can be used for all QA/QC requirements. Some pre-1994 drilling has limited QA/QC data but where possible a new hole will be used to twin the historic holes.

10.4 BURNSIDE AREA

10.4.1      DIAMOND DRILLING

Over all its properties Crocodile Gold conducted a significant amount of drilling in 2011-12, mainly targeting the expansion of mineral resources. Some regional “green fields” exploration drilling was conducted over selected targets in late 2011.

A summary of the Diamond Drilling completed for specific areas is shown in Table 10-17 below.

Deposit Diamond First Drilled Last Drilled
Holes Meters Size
Yam Creek 3 236 HQ-HQ 13-Sep-11 11- Nov-11
Rising Tide 3 200 HQ-HQ 03-Jul-11 16- Oct-11
Exploration 2 184 NQ-NQ 03- Nov-11 06- Nov-11
Total 8 620      

TABLE 10-17 2011-12 DIAMOND DRILL STATISTICS FOR BURNSIDE AREA

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10.4.1.1   Mt Bonnie Northeast

The 2011 VTEM survey defined a strong conductor with a coincident magnetic anomaly located to the northeast of the Mt Bonnie sulphide deposit. The interpretation of the VTEM results indicated that the stratigraphy in the area might be somewhat different to that which is shown on published geology maps. It is interpreted that the Mt Bonnie NE conductor is hosted by Gerowie Tuff, which is lower stratigraphically than the Mt Bonnie sulphide horizon. Essentially, this newly defined conductor may be on a previously unmapped embayment of the Margret syncline. The signature of the conductor was very similar to that produced by the Mt Bonnie and Iron Blow deposits

A preliminary field investigation indicated significant quartz veining on surface some of which had appreciable gossanous material associated.

The NT Government mineral occurrence records indicated an unnamed copper showing in the area of the conductor. A field investigation indicated that at the coordinates provided there was only outcrop of Zamu Dolerite. Subsequent fieldwork found a small copper showing several hundred meters to the north and west.

It was decided to drill test the conductor in 2011 and one shallow drill hole (93.1m) was collared drilling to the east at 95°. The hole encountered significant quartz veining with appreciable amounts of pyrite and pyrrhotite with very minor arsenopyrite and chalcopyrite from 43.7 to 93.1m. The VTEM conductor and magnetic anomaly were adequately explained. Younging directions were up-hole, indicating the sequence is on the east limb of a syncline as the VTEM results indicated.

The disseminated magnetite seen in the hole over narrow intervals is of some interest, possibly indicating that the drill hole intersected the periphery of a sulphide system. Additional work should be considered. No significant assays resulted from the samples submitted.

Hole ID Orig. East Orig. North Orig. RL Local Azimuth Dip Max Depth (m)
MBEXD001 12206.096 9039.390 1140.761 95.610° -60.8° 93.1

TABLE 10-18 DRILL HOLE CO-ORDINATES MT BONNIE EAST

10.4.2      RC DRILLING

No “greenfields” RC drilling has been completed for the Burnside area, however, a significant amount of drilling was completed into the Yam Creek and Rising Tide deposits during 2011, these are summarized below.

Deposit RC First Drilled Last Drilled
Holes Meters Size
Yam Creek 35 2,321 4.5-5" 13-Sep-11 11-Nov- 11
Rising Tide 88 6,591 4.5-5" 03-Jul-11 16-Oct-11
Exploration - -   03-Nov- 11 06-Nov- 11
Total 123 8,912      

TABLE 10-19 2011- 12 RC DRILL STATISTICS FOR BURNSIDE AREA

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10.4.2.1   Rising Tide

In early 2011 it was decided to assess areas that could be quickly brought into mine production. During this review it was noted that the Rising Tide area still had an active mine permit so should be assessed for mining potential. During 2011 Crocodile Gold drilled 88RC holes for 6,591m of drilling and three diamond holes for 200m of HQ core.

This drilling data was then used to develop a new mineral resource model. Rising Tide entered into mine production from December 2011 and finished the stage one pit in June 2012.

With local infrastructure still in place from previous mining activity on this deposit, as well as the adjacent Brocks Creek Mine facilities, very little capital was required to put this deposit back into commercial production.

The deposit is hosted by the same stratigraphic horizon as the Cosmo deposit with mineralization in close proximity to sheared sediments in contact with a dolerite sill, which acts as a basal sequence in the area. Mineralization is interpreted to be hosted by multiple sub-parallel pyrrhotitic, carbonaceous units within the sediments. Both normal and reverse faults bound the mineralization. Mineralized horizons vary between 2.0 and 10.0m thick. Minor mineralization extends into the dolerite sill. The deposit is in very close proximity to a large granitic unit located to the north, which has created some contact metamorphic assemblages and structural complexities.

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10.4.2.2   Yam Creek

During 2011 it was decided to drill around the past mining centers of North Point and Princess Louise as any future discoveries in this area could be mined in conjunction with these pits. It was recognized that significant past drilling results were located close to surface and could potentially be exploited.

Drilling at the Yam Creek consisted of 35RC holes (2,321m) and three diamond holes (236m) for a total of 2,557m. Work on the Yam Creek deposit focused on in-filling historic RC drilling to test the continuity of grade along strike from current open pits. Further mineralization has the potential to extend the open pits. Drilling at the Temperance pit aimed at testing the continuity of grade of the deposit both along strike and below the pit.

Results from the drilling at the Yam Creek deposit highlighted some significant intercepts including 2m @ 4.41g/t Au (YCRC111 66m), 7m @ 5.66g/t Au (YCRC115 52m), 1m @ 28.1g/t Au (YCRC 21m), 6m @ 3.23g/t Au (YCRC125 33m), 3m @ 4.44g/t Au (YCRC142 33m) and 5m @ 8.48g/t Au (YCRC137 49m). These intersections have now been geologically modeled.

Drilling at the Temperance deposit intercepted some narrow mineralized zones including 0.95m @ 1.87g/t Au and 0.85m @ 1.35g/t Au (YCDD101 2.78m and 7.6m respectively), 1m @ 1.79g/t Au (YCRC100 14m), 1m @ 1.52g/t Au and 2m @ 1.85g/t Au (YCRC103 87m and 95m respectively). While these results were seen as promising they do not require further work.

While the drilling was completed in 2011 no further work has been completed as this deposit is ranked lower in priority than the other deposits worked on during 2012.

In 2011 two exploration holes were drilled targeting VTEM anomalies around the two polymetallic deposits within the Burnside area. These two drillhole locations are shown in Figure 10-12 below.

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10.4.3      HISTORIC DRILLING

While Newmarket Gold completed significant amounts of drilling in the Cosmo Mine area, there has been a reliance on some historic drilling datasets. Generally, Newmarket Gold still holds onto all the original hard copy data including assay reports, which can be used to cross-reference digital data. This process will require a significant amount of man-hours but that which has already been done has shown the digital data to be correct and consistent.

With this, a significant amount of the assay results for this historic drilling have been completed using the Australian Laboratories in Pine Creek, Northern Territory (NAL), or another company managed by the same laboratory manager as is present in the current NAL. On speaking with the manager it was noted that the laboratory procedures are the same now as they were historically including the same internal QA/QC processes. This gives confidence that a standard approach has been completed even on this historic data.

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While these statements above are generic in nature, more detailed information on each deposit is outlined below.

10.4.3.1   Rising Tide

Drilling using either RC or diamond core has been recorded in the company database since 1990. Several companies have drilled the deposit with the first significant drilling being completed by Solomon Pacific between 1994 and 1996 as part of the development of the Brocks Creek mining center. Over the years a total of 140RC holes for 10,407m have been recorded including the work done by Crocodile Gold in 2011. This drilling was used in the mineral resource up-date included in this report. A summary of the drilling is below.

Company Year Number of Holes Meters
Cyprus 1993 3 349
SolPac 1994-96 23 2,452
UNK 1990 5 182
BOPL 2006 21 833
Newmarket Gold 2011 88 6,591
Total   140 10,407

TABLE 10- 20: HISTORIC RC DRILLING AT RISING TIDE DEPOSIT

Company Year Number of Holes Meters
Newmarket Gold 2011 3 200.3
BOPL 2006 2 161.27

TABLE 10-21 HISTORIC DIAMOND DRILLING AT RISING TIDE DEPOSIT

The laboratories used for this work include ALS, Amdel and AssayCorp (a predecessor to NAL). The method used is most commonly Fire Assay with AAS. Some ICP work has also been recorded but mostly looking at base metals and not gold.

Drilling methodology is also generally captured in the database with diamond drilling mostly HQ and 5.5” RC hammer with riffle splitter for sample collection for RC.

10.4.3.2   Yam Creek

Yam Creek, as with Rising Tide, has been developed since the mid-1990 with modern drilling. Several companies have drilled the deposit over the years using both RC and diamond methodology. The summary of the drilling used is outlined in the tables below. As before the majority of assaying was completed at Amdel, ALS or AssayCorp/NAL using a fire assay technique.

Company Year Number of Holes Meters
DML 1994 153 7,919
UNK   26 724
NGNL 1996 26 1,996
Newmarket Gold 2011 36 2,381
Total   241 13,020

TABLE 10-22 HISTORIC RC DRILLING AT YAM CREEK DEPOSIT

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11 SAMPLE PREPARATION, ANALYSIS AND SECURITY

This section specifically summarizes the sampling, and analysis of that sampling, completed by Crocodile Gold/Newmarket Gold since 2009. A review of historic data (drilled prior to 2009) used in the reported models is outlined in Section 12 below. This includes a review of the historic data that is housed in the company database and a review of hard copy results and reports that are still available. This will also include the review of previously reported QA/QC results from these drilling programs to add confidence in the data used in the mineral resource estimates.

For this section Newmarket Gold/Crocodile Gold has completed drilling and sampling on the deposits outlined below:

11.1 REVERSE CIRCULATION DRILLING SAMPLING

The geologist sieves and washes a portion of each 1.0m RC sample interval. The sample is then inspected to determine its geological attributes. Geological descriptions are entered directly onto standard logging sheets in either a hard copy or digital form via a portable computer, using standardized geological codes. Each washed sample is then stored in a chip tray, which is stored on shelving at the exploration yard for future reference if required.

RC drillholes are typically sampled on 1.0m intervals the drill cuttings are riffle or cone split to produce a final sample of approximately 2 to 3kg. There is a systematic submission of duplicates, barren flushes, standards and blanks into the sample stream. At the completion of each hole samples for assay are collected in large plastic bags in short intervals. These are sealed on site and stored ready for dispatch to the laboratory.

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11.2 DIAMOND DRILLING SAMPLING

Geologists log each hole, paying particular attention to the degree of weathering, lithological contacts, structural contacts, alteration, mineralization, and geotechnical information. Drill core was oriented based on the orientation marks on the core made during drilling. Core orientations were marked for the bottom side of the hole using a Camteq Orishot tool at the end of each core rod. Zones of core loss are identified and marked by inserting marker blocks recording the exact length of the core loss.

At the completion of logging, the geologist marks the core ready for sampling and a photo was taken of each tray, as a means of checking the intervals as well as geological logs if required. Sample intervals are chosen based on lithological contacts or where there are significant changes in the nature of the gold mineralization with no overlaps over geological boundaries. Sample boundaries are often pre-existing breaks; otherwise the half core was cut perpendicular to the core axis.

A minimum sample size of 0.3m and a maximum size of 1.5m were cut using an Almonte automated diamond saw. The core was cut so as to divide the mineralization in half whilst preserving the orientation line. Some drillholes were sampled over their entire length whereas other drillholes were sampled from 20-50m into the hangingwall, through to the end of hole.

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The underground diamond definition drill programs are intense infill programs drilling the deposit on nominal 15 m x 20 m spacing. 80% of the Underground holes are full core sampled with the remaining 20% half core sampled for future record checks and reference.

Each sample was placed into pre numbered calico bags with standards, blanks, barren quartz flush material and duplicates placed within calico bags during this stage. Samples are then loaded into green plastic bags with the sequence of samples in the bag labeled to assist sorting at the lab. The green plastic bags are then placed into dispatch cages and dispatched at the end of each hole either by Newmarket Gold staff or by courier directly to the Laboratory.

At the completion of each hole, the core is moved to a secure site and the trays stored for future retrieval, if warranted.

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It is believed that all analytical work, including sample preparation, analytical procedures, QA/QC measures and associated security and chain of custody procedures have been completed in accordance with the established protocols routinely used by Newmarket Gold. All analytical work for the reported drilling was completed at either NAL, at the Northern Territory Environmental Laboratories laboratory in Darwin (NTEL), or the Australian Laboratory Service (ALS) laboratory in Brisbane or Townsville. NAL is a non-certified laboratory while ALS is ISO 90001 certified. Regular lab visits by the Authors and other Company staff to meet with the management of the laboratories and inspect the facilities. All laboratories used are independent to Newmarket Gold and are well known to the Authors as competent assayers. The Authors consider that these procedures and protocols are of acceptable quality and are broadly consistent with international “best practice” standards.

11.3 COSMO MINE FACE SAMPLING PROCEDURE

The following outlines the process for collecting underground samples for individual faces and the walls of drives. The direction of the bedding strike dictates whether wall sampling or face sampling is required. Sampling is ideally conducted as perpendicular to the bedding as possible, as opposed to sampling parallel to the bedding where one bedding layer is in effective being sampled.

Face sampling (mineralization development drives):

Checking there is adequate ground support and that the face is fully scaled.
   

Spray one line horizontally across the face, the height of this line is depended upon the height of the sampler and if the drive has been adequately bogged out. This line is the delineator between the upper and lower samples.

   

A quick log of the face is then conducted and vertical lines are then applied to separate the different lithological units. These lines will not be perfectly vertical as they will be along the bedding plains. Units that are thinner than the minimum sample size of 0.3m will be included in another sample interval.

   

Any lithological units, which are wider than 1.4m are then split up with further lines as 1.4m is the maximum sample interval.

   

The intervals are then measured from left to right; the distance from the left hand wall is sprayed next to the sample interval vertical line.

   

The face location, level and heading, along with the date are sprayed on the wall.

   

A photo is taken before chipping begins.

   

The face is then mapped. Bedding direction, structures, locations of veins and massive sulphides are drawn to scale. The sample intervals are recorded and sample bag identification numbers are assigned. A rock description of each interval is then recorded including rock type, percentage sulphides and a description of the sulphides, any veins and their composition, alteration and any other significant details.

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Sampling is then conducted with a rock hammer; the rock is chipped directly in to the sample bags. The sample will only come from the area of the face, which has been outlined by the spray paint lines. In the ideal situation 3 to 4kg of rock will be collected evenly from across the interval ensuring a representative sample of the sample interval is obtained. If the rock is particularly silicified then a hand- held percussion drill can be used. This process is then repeated for all the sample intervals across the face.

   

The distance from the center of the face to the nearest survey station will then be measured and recorded on the face sheet Figure 11-3.

Wall sampling (access drives)

Wall sampling is conducted normally on crosscuts where development is running perpendicular to the mineralization. The process of sampling is identical to the face sampling process in the mineralization development drives. The difference is only one line of sampling is conducted on both the left hand wall and the right hand wall, not upper and lower. The starting position of the mapping is carefully recorded using a combination of measurements and sketches in plans.

11.4 SAMPLING PREPARATION

Three commercial laboratories were used throughout the past drilling campaigns with North Australian Laboratories (NAL) in Pine Creek, Northern Territory being the primary laboratory for the Cosmo and Esmeralda mineral resource drilling. Northern Territory Environmental Laboratories (NTEL), (now Genalysis) was used in the past for some drilling programs (Howley for example). Australian Laboratory Services (ALS) in Darwin acted as an umpire lab for the drilling at Cosmo and Esmeralda deposits. Some samples sent to ALS were prepared in Darwin and then sent to either the ALS laboratory facilities in Perth, Brisbane or Townsville for analysis. Some primary samples were sent to ALS due to restrictions on NAL in Pine Creek.

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11.5 SAMPLE SECURITY

A Company staff member is stationed on the RC drill rig while samples are being drilled and collected; this is generally a Company or contract geologist. At the end of shift samples were generally transported to the sample collection area where they are stored in crates as they await transportation the lab. For some drill holes at Esmeralda the samples were shipped at the end of each shift to the lab in Pine Creek. Samples are shipped in regular intervals so they are not in crates for any length of time. These samples are located at the Brocks Creek exploration office, which can be secured if no staff member is on site.

In terms of diamond drilling, the core is collected daily from the rig and transported to the geology office near the Cosmo Mine or at Brocks Creek exploration yard. The drill core is then stored in the core shed for logging and sampling. Both core shed processing facilities (Brocks Creek and Cosmo Mine) are located in compounds with security fencing. These locations have limited access when no Newmarket Gold staff member is present. Samples are cut at this location and samples loaded into lab crates as they await collection. These samples are then transported directly to the lab for analysis.

Once assaying is complete the results are returned in digital format to the data entry personnel. Cosmo Mine Geology results are loaded directly into an acQuire database. Validation takes place via a visual comparison of expected values of standard and blanks against received assay values. Any questionable results are demoted in priority, not to be used in mineral resource modeling, and are investigated further. Once investigation is complete the priority is adjusted if the original assays were considered correct, or any re-assay work imported and promoted in priority for use in mineral resource modelling work. Exploration results are imported into a DataShed database and checked visually against the expected values. If the results are considered incorrect further testwork is completed on the samples, with any results considered correct are imported into the database overwriting the original results. Any incorrect assays and re-assays are noted in the digital sampling logs and results sheets.

The Cosmo Mine Geology AcQuire database is located at the Cosmo Mine office and the Exploration DataShed database is located at the Union Reefs office. Both databases software utilizes SQL database systems with in-built security limiting access to people outside the Company network or without sufficient login access.

11.5.1      NAL

NAL is an independent laboratory based in Pine Creek. The relationship between NAL and Newmarket Gold is on a client/supplier arrangement with a contract in place for services.

Upon arrival at the laboratory, samples are sorted, reconciled against the accompanying paperwork and dried on racks in the oven. Each sample is initially crushed in a jaw crusher to the size of 10mm. following the jaw crusher, each sample is passed through a roll crusher to the size of 2mm. Samples are riffle split into two sub-samples - one sample is milled, whilst the other is retained as a coarse reject and returned to Newmarket Gold. The sub-sample retained for analysis is milled to 100µm in a Keegor mill. Each milled pulp sample is further split to provide 50g for fire assay (FA50). The remaining sample is kept as a pulp sample for future analyses and returned to Newmarket Gold. After firing, samples are analyzed using AAS, with results reported in ppm.

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Quality control procedures include blasts of compressed air to clean jaw and roll crushers between samples, a barren flush of river sand to clean bowls in between samples, laboratory duplicate samples undertaken at a rate of 1 in 10 and the insertion of NAL internal standards intermittently at a rate of 1-2 times on samples.

11.5.2      ALS

Pulp samples are sent to the ALS preparation facility in Darwin where samples are reconciled against the accompanying paperwork, barcoded for tracking and sent to the analytical lab in Perth where 30g of pulp is weighed off for fire assay with an AAS finish (AA26). Results are reported in ppm. During late 2015 some original samples were transported directly to the ALS lab in Townsville for analysis due to restrictions at the main NAL lab in Pine Creek. These samples were prepared using the same methodology as used in the facility based in Darwin.

ALS laboratories are certified using the ISO9001:2008 accreditation (“Quality Management Systems – Requirements”). They also hold the NATA Technical accreditation under ISO17025:2005. They are a commercial laboratory based in Brisbane and Perth who supply an assaying service to the Company under contract rates.

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11.5.3      NTEL

NTEL is an independent laboratory based in Darwin (now called Genalysis). The relationship between NTEL and Newmarket Gold is on a client/supplier arrangement with a contract in place for service.

Upon arrival at the laboratory, samples are sorted, reconciled against the accompanying paperwork and an average sample weight for the submission is taken. Each sample is then dried at 105ºC until fully dry. Each sample is initially crushed in a jaw crusher to the size of 2mm. Following the jaw crusher, each sample is rotary spilt with 300g being taken for milling and assay and the remainder being set aside as a coarse reject and returned to Newmarket Gold. The 300g sample is then milled to pass through a roll crusher to the size of 2mm. Samples are riffle split into two sub-samples - one sample is milled, whilst the other is retained as a coarse reject and returned to Newmarket Gold. The sub-sample retained for analysis is milled to 85% passing 75µm with 1 in 20 samples wet screened to check for compliance. Each milled pulp sample is further split to provide 25g for fire assay (FA25) with <1g used for multi element analysis if requested. The remaining sample is kept as a pulp sample for future analyses and returned to Newmarket Gold. After firing, samples are analyzed using AAS, with results reported in ppm.

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11.6 QUALITY ASSURANCE/ QUALITY CONTROL

11.6.1      COSMO MINE

Samples for the Cosmo Mine block model have had their validity monitored by consistent quality assurance and quality control (QA/QC) checks.

Surface drilling QA/QC procedures include insertion of control samples (standards) and barren quartz flushes (BQF), inter-laboratory pulp checks and submission of sample duplicates. RC duplicates are produced from splitting the original sample straight from the rig mounted cone splitter. Diamond core duplicates are sent as a quarter core sample cut from the original sample interval. Blanks and standards are generally inserted in the sample stream every 30 samples.

Underground drilling QA/QC procedures include insertion of control samples (standards), inter-laboratory pulp checks and the assaying of field duplicates during the various drill programs. Internal laboratory repeats provide an indication of the laboratory precision. Common standards were included with the inter-laboratory check samples to allow the performance of both laboratories to be gauged. Local blank (un-mineralized) dolerite samples were submitted to assess laboratory hygiene. For the year 2015 a total of 4,609 QA/QC samples were taken, including blanks, Company standards, lab standards, barren quartz flushes and umpire lab repeats.

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The following section describes the QA/QC procedures adopted by Newmarket Gold for the entire data collection period for the deposit model.

Diamond % of
total
RC % of
total
Samples 162,446 71% 83,814 87%
Blanks 3,683 2% 14 0%
BQFs 455 0% 0  
Duplicates - - 52 0%
Repeats 48,151 21% 12,098 13%
Standards 2,733 1% 97 0%
Lab standards 10,444 5% 101 0%
Umpire lab repeats 1,409 1% 0 0%

TABLE 11-1 RATE OF QA/QC SAMPLING FOR COSMO OPERATION 1 JANUARY 2010 TO 31 DECEMBER 2015

11.6.1.1   Standards

Certified standards are submitted to the laboratory on a regular basis. A standard is inserted into every batch periodically throughout diamond drilling sampling usually every 30 samples and similar insertion for RC drilling programs.

The lab standards total of 2,505 for 2015 is not split on hole type because there are some Lab Job Numbers containing both drill holes and face/ROM sampling.

Drilling by Newmarket Gold at the Cosmo Mine has seen certified standards submitted to NAL on a regular basis. A range of 45 different certified standards have been used since the start of the operation and these are summarized in the table below. Standards used in 2015 are highlighted in yellow.

Standard Au (ppm)   Standard Au (ppm)   Standard Au (ppm)
ST02 2.37   ST274/5358 5.96   ST499 0.4
ST02/5355 2.37   ST28/6366 34.5   ST504 1.42
ST04 4.87   ST28/9489 34.2   ST508 3.29
ST05 2.58   ST335 13.65   ST535 0.97
ST07/8441 0.23   ST347 9.6   ST559 0.52
ST07/9258 0.22   ST383 7.24   ST576 1.37
ST08 0.32   ST39 1.67   ST590 0.22
ST08/6342 0.32   ST39/6167 0.87   ST603 0.38
ST08/8225 0.33   ST39/9420 0.89   ST605 0.42
ST09/3320 1.93   ST43/7370 3.37   ST622 2.04
ST09/7382 1.93   ST48/9278 4.55   ST631 2.43
ST10 2.94   ST482 1.94   ST684 0.75
ST14 0.405   ST487 0.49   ST698 0.65
ST15/6138 0.022   ST49/6403 1.99   ST70 0.099
ST16/5357 0.52   ST493 0.119   ST73/7431 1.54

TABLE 11-2 LIST OF STANDARD SAMPLES USED AT COSMO MINE

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The graph above shows the standards sent by Newmarket Gold to the NAL lab since the last technical report period. Labeling errors were likely the cause as to why standards ST508 and ST605 returned some results well outside their expected ranges. ST08/6342 (low grade Au 0.32g/t) was found to consistently grade above the acceptable maximum Au value of 0.36g/t reporting a mean value of 0.36g/t. Standards ST631, ST535, ST508 and ST504 also had multiple results outside the acceptable range. An example of the chart and statistics used to chronologically check the standard results is shown below (Figure 11-8 , Table 11-3).

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STANDARD ST535 NAL
Recommended Value g/t Au 0.97
Standard Deviation 0.040
Number of Assays 99
Mean Result g/t Au 1.0041
Minimum g/t Au 0.79
Maximum g/t Au 1.08
% Outside Error Limit 11.1%
Outside 2SD 11
Outside 3SD 1

TABLE 11-3 COSMO MINE STANDARD SR535 COMPLIANCE TABLE

11.6.1.2   Blanks

Blank materials included in the sample stream were derived from several sources including barren core (Dolerite core drilled during the program), barren coarse rejects, crushed Bunbury Basalt (from Gannet Holding Pty Ltd, referred to in this report as “blank”). Blank results above 0.02ppm Au are queried and any issues resolved.

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Anonymous blank samples were submitted to NAL on a regular basis by Newmarket Gold. A blank half core sample was often submitted strategically placed to follow a high-grade sample or zone; otherwise blanks were distributed randomly through each hole. In this way blanks were used to check the sample preparation and to identify possible contamination from the high-grade samples.

Blanks, particularly the dolerite core submitted, tended to return more assay grade than expected although the amount of contamination was considered to be minimal and associated with the original host rock material. For the year 2015 un-mineralized dolerite core was used for blank material. The chart above (Figure 11-9) shows the results of all samples during this period. Of note is one high value, which is likely due to a mineralized piece of core (probably containing coarse gold) being selected accidentally. Investigation of these samples showed that they were not preceded by high-grade samples so the result could not be due to poor assay hygiene/gold smearing. Some other higher than expected results were also seen but these are also believed to be the result of mineralized blank core being selected. The insertion of blanks after high-grade sample intervals has allowed the situation of sample contamination to be monitored effectively, and highlights the “cleansing” effect of strategic placement of blanks and barren quartz flush, preventing contamination of subsequent samples.

The average grade of all blank dolerite core samples for the year was 0.08ppm, which is in line with background value in the mine area. The median value for all blank core samples for the year 2015 was 0.05ppm.

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11.6.1.3   Laboratory Duplicate Assays

Relative precisions have been used to analyze the accuracy of duplicate samples. The relative precision is a measure of dissimilarity, that is, if both distributions are exactly the same, this value will equal zero and increases as the distributions become more dissimilar.

Relative precision has been calculated using all data pairs for the ranges of below detection.

Newmarket Gold for the period of 2015 has undertaken 52 duplicate assays from RC drilling.

11.6.1.4   Inter-Lab Repeats

Inter-lab repeats were taken for diamond drilling programs with pulp material sent to umpire lab ALS in Perth for assay. Results were compared to original assay results for each area and each umpire lab separately. Two examples of the tables and charts used to analyze each drill type and umpire lab are below.

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INTER-LAB REPEATS: NAL ALS
Number of assays 209 209
Minimum value g/t Au 0.01 0.03
Maximum value g/t Au 21 21.8
Mean g/t Au 2.1878 2.1832
Median g/t Au 0.905 0.8
Variance 10.0587 10.2824
Standard deviation 3.1715 3.2066
Coefficient of variation 1.4497 1.4687
Correlation co-efficient R 0.899  
Coefficient of determination R2 0.808  

TABLE 11-4 STATISTICAL RESULTS FOR COSMO MINE INTER-LAB REPEATS - 100, 200 & 300 LODES

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Type of Repeat NAL ALS
Number of assays 285 285
Minimum value g/t Au 0.02 0.02
Maximum value g/t Au 26.5 30.2
Mean g/t Au 2.7271 3.0286
Median g/t Au 1.155 1.07
Variance 16.1562 22.126
Standard deviation 4.0195 4.7038
Coefficient of variation 1.4739 1.5532
Correlation co-efficient R 0.7413  
Coefficient of determination R2 0.5496  

TABLE 11-5 STASTISTICAL RESULTS FOR COSMO MINE INTER-LAB REPEATS - 101, 400, 500, 600 AND WESTERN LOADS

11.6.1.5   Analysis of Inter-Lab Repeats

For the year 2015, a total of 494 original pulps samples were sent to ALS for inter-lab fire assay check against NAL. The data has been split into two groups, which represent contrasting areas of the mine.

The first group of samples was taken from mineralization intervals associated with the 100, 200, and 300 Lodes (Figure 11-10). The statistical results of the data set indicate a strong correlation between the assays received with a Correlation Co-efficient of 0.899 for the entire data set. The repeat assay mean is 0.0046g/t Au lower than the original assay (Table 11-4).

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The Coefficient of Determination represents the percentage of the data that is closest to the line of best fit. The R2 value of 0.808 indicates 81% of the values for the 100, 200 & 100 300 Lodes have a strong linear relationship, which means that independent umpire lab (ALS) strongly agrees with the original lab (NAL). The poor relationship of 19% of samples can be explained by the moderate relative nugget range of 21% to 29% in the modeled variography.

The second group of samples was taken from mineralization intervals associated with the 101, 400, 500, 600 and Western Lodes (Figure11-11). The statistical results of the data set indicate a weaker correlation between the assays received with a Correlation Co-efficient of 0.7413 for the entire data set. The repeat assay mean is 0.3015g/t Au higher than the original assay (Table11-5).

The R2 value of 0.5496 indicates that 55% of the values for the 101, 400, 500, 600 and Western Lodes have a strong linear relationship and that the independent umpire lab (ALS) shows moderate variation with the original lab (NAL). The poor relationship of 45% of samples could be explained firstly by the sample set, which was taken from more structurally complex areas of the mine with a higher instance of coarse gold and vein related gold. Secondly the relative nugget range is higher in these areas with between 22.7% and 32.3% in the modeled variography.

The Q-Q Plot for all inter lab repeat samples (Figure 11-12) indicates that pulp samples with grades between 0g/t Au and 8g/t Au repeat very closely. In the 8–12g/t Au range the repeatability tends to decrease, with NAL appearing to report slightly higher grades than ALS. Above 12g/t Au ALS reports higher than the NAL original, but in this range samples are likely to contain coarse gold and therefor repeatability is decreased due to the nugget effect.

11.6.1.6   Opinions on Sampling, Security and Ananlyis

The three laboratories that have been used by Newmarket Gold for the entire Cosmo Mine QA/QC program offer different preparation techniques with a 50g fire assay by NAL, a 30g fire assay by NTEL and a 30g and 50g fire assay by ALS. During 2015 only NAL and ALS labs were utilized, both using the 50g fire assay technique.

The following summarizes findings with respect to assay work from the two independent laboratories:

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11.6.1.7   Recommendations

The results from the QA/QC analysis of drill samples has indicated a change in the level of confidence in assay grades for use in the mineral resource model due to inclusion of structurally complex areas of the deposit with higher instances of coarse gold. The following recommendations for improvements in the current procedures are:

Standards continue to be inserted on site and independently of the assaying laboratories. The procedure for selecting and sending standards from site needs to be improved to stop incorrect standard sample labeling.

   

Standards that return assay values outside the minimum/maximum range need to be rejected and re- tested by the lab. Re-testing will give information about lab testing precision. Lab jobs containing significant standard assay inconsistencies need to be reviewed.

   

Bunbury Basalt coarse Blank material to be used for all Blanks moving forward, with the aim of eliminating gold assay values that come from the local dolerite material due to free gold and selection errors.

   

It is recommended that inter-lab checks be undertaken every quarter with pulps selection being taken from the main mineralization horizons.

   

Further inter-lab test work using quarter cored diamond holes to investigate analytical processes and data correlation between labs.

   

NAL appears to be reporting variable values when compared to the ALS repeats. For Au values 0.1- 2.0g/t the original NAL samples are consistently higher than the ALS check value, for Au values 2.0– 35.0g/t the NAL samples are consistently lower than the ALS check value. This needs to be monitored closely with inter-lab checks done every quarter on both pulps and original quarter cored samples.

   

Additional grind size test work for all diamond holes to investigate the effects of grind size on gold values, particularly when coarse gold is present.

   

Campaigns for re-assaying the pulps at the same lab (in this case NAL) should be undertaken on a more regular basis. The sample dispatches should be identified using the QA/QC analysis data from the Standards and Blanks charted plots.

   

Procedures should be put in place to ensure the high standard of QA/QC is maintained. The acQuire database functionality is sufficient for assessing QA/QC and needs to be utilized more. This database has been set up to allow greater control and geostatistical analysis of the QA/QC data being imported into the database.

   

Investigate whether tighter drill spacing will increase statistical confidence & repeatability of assay values for the 500, 600 and 101 Lodes where the nugget effect is higher.

   

Monthly QA/QC reports on Diamond assay results, standards and blanks need to be developed.

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11.6.2      ESMERALDA

Samples for the Esmeralda deposit block model have had their validity monitored by consistent quality assurance and quality control (QA/QC) checks. The historic drilling used in the mineral resource model wisas considered to have passed all QA/QC requirements and the assay results were accepted as true and correct.

The 2015 RC Surface drilling QA/QC program included the insertion of control samples (standards) and barren quartz flushes (BQF), inter-laboratory pulp checks and submission of sample duplicates. RC duplicates are produced from splitting the original sample straight from the rig mounted cone splitter. Blanks and standards are generally inserted in the sample stream every 30 samples.

  Diamond % of total RC % of total
Samples 245 65.33% 5,190 85.60%
Blanks 26 6.93% 166 2.74%
Duplicates 0 0.00% 156 2.57%
Repeats 46 12.27% - 0.00%
Standards 26 6.93% 167 2.75%
Lab standards 32 8.53% 302 4.98%
Umpire lab repeats 0 0.00% 82 1.35%
Total QA/QC 130 34.67% 873 14.40%
Total Samples 375 100.00% 6,063 100.00%

TABLE 11-6 ESMERALDA DEPOSIT QA/QC SAMPLING RATES

11.6.2.1   Standards

Certified standards were submitted to the laboratory on a regular basis. A standard is inserted into every batch periodically throughout sampling usually every 30 samples.

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All standards that were submitted as part of the 2015 drilling completed at the Esmeralda deposit were normalized around their standard deviation values and plotted relative to their sample ID numbers as shown in Figure 11-13. The graph shows some significant variation in the expected values sometimes approaching 50 standard deviations from the expected value. The majority of the errors are attributable to the incorrect standard being submitted during sampling.

  NAL On Site  
Std_ID Gold Grade
(g/t Au)
Gold Grade
(g/t Au)
Difference
ST16 0.52 0.56 0.04
ST09 1.93 1.86 0.07
ST08 0.32 0.32 0
ST622 2.04 2.04 0
ST622 2.04 2.04 0
ST684 0.75 0.75 0
ST16 0.52 0.51 0.01
ST559 0.52 0.53 0.01
ST08 0.32 0.34 0.02
ST559 0.52 0.56 0.04
Blank 0 0.01 0.01
ST09 1.93 2.05 0.12
ST684 0.75 0.78 0.03

TABLE 11-7 ESMERALDA DEPOSIT INTER-LABRATORY STANDARD PERFORMANCE CHECK DATA

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During the drilling program the performance of the standards was questioned, some standards were thought to be old and may have degraded. A program of inter laboratory checks of standards was conducted with standards being tested at On Site Laboratory Services in Bendigo, Victoria. On Site tested the standards through a 25gm charge fire assay using a similar method to NAL. The performance of the standards at both labs was very close, confirming the accurate performance of the standards.

The overall standard performance throughout the drilling was of an acceptable standard, with no bias observed in the reported results ignoring the obvious errors of misplaced standards.

11.6.2.2   Blanks

During the 2015 drilling program Barren Quartz Flushes were utilized as blank material to test for sample preparation cross contamination.

Several instances of sample contamination were observed, however only at low levels. The majority of the large assay values returned from blank samples can be attributed to sample placement errors. Cross contamination is very unlikely in these cases with no significant mineralisation surrounding the blank sample to cause the contamination.

11.6.2.3   Laboratory repeat Assays

During the 2015 drilling the laboratory’s routinely conducted repeat firings of original samples, which showed elevated gold results. The repeat firings showed a good correlation between the original and repeat results showing a correlation co-efficient value of 0.8536, and show no bias across the repeat sampling.

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Correlation (Original Vs Repeat)
Range (ppm)
Combined 0.92
<0.20 0.71
0.21 - 0.50 0.78
0.51 - 0.70 0.68
0.71 - 1.00 1.00
1.01 - 1.40 -1.00
1.41 - 5.00 0.61
>5.01 -

TABLE 11-8 ESMERALDA DEPOSIT ORIGINAL ASSAY VS REPEAT ASSAYS CORRELATION BETWEEN GRADE RANGES FOR LAB REPEATS, AU G/T

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              Absolute  
            Average Average  
                 
R - CHART         % Diff. % Diff % Diff  
Range
(ppm.)
# of
Assays
% of
Total #
Mean
Original
Mean
Repeat
Between
Means
Between
Assays
Between
Assays

S.D.
<0.20 28 50% 0.07 0.07 -3.23% -6.04% 39.66% 0.06
0.21 - 0.50 11 20% 0.35 0.35 -0.44% -25.24% 34.89% 0.09
0.51 - 0.70 5 9% 0.58 0.56 -4.18% 1.48% 4.36% 0.07
0.71 - 1.00 2 4% 0.81 0.85 5.08% -12.17% 22.37% 0.10
1.01 - 1.40 2 4% 1.16 1.16 0.00% -21.75% 27.06% 0.13
1.41 - 5.00 7 13% 2.41 2.87 16.03% -3.87% 30.39% 0.72
>5.01 0 - - - - - - -
TOTAL 56 100% 0.57 0.63 10% -2.39% 44.02% 0.82

TABLE 11-9 ESMERALDA DEPOSIT TABLE OF STASTICS FOR 2015 LAB REPEATS, AU G/T

Tables 11-8 and 11-9 summarize the performance of the laboratory repeats across selected grade ranges.

11.6.2.4   Inter-Lab Repeats

A series of inter laboratory repeats was conducted from the RC drilling completed in 2015. A selection of 82 samples including standards were prepared and assayed at NAL, with the pulps sent to ALS for re-assay. Figure 11-16 summarizes the results from the check sample program.

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A very good correlation can be seen between the original NAL samples and the repeated ALS samples, with a correlation coefficient of 0.983 giving confidence to the accuracy of the assay results received from both labs.

Correlation (Original Vs Repeat)
Range (ppm)
Combined 0.992
<0.20 0.899
0.21 - 0.50 0.802
0.51 - 0.70 -0.866
0.71 - 1.00 -1.000
1.01 - 1.40 0.991
1.41 - 5.00 0.908
>5.01 -

TABLE 11-10 ESMERALDA DEPOSIT ORIGINAL ASSAY VS REPEAT ASSAYS CORRELATION BETWEEN
GRADE RANGES FOR INTER -LABRATORY CHECKS

              Absolute  
            Average Average  
R - CHART         % Diff. % Diff % Diff  
Range # of % of Mean Mean Between Between Between  
(ppm.) Assays Total # Original Repeat Means Assays Assays S.D.
<0.20 18 30% 0.12 0.09 -41.14% -55.79% 59.71% 0.04
0.21 - 0.50 24 39% 0.34 0.31 -9.27% -21.62% 24.83% 0.09
0.51 - 0.70 3 5% 0.58 0.57 -1.16% -15.33% 15.33% 0.02
0.71 - 1.00 2 3% 0.77 0.98 21.94% -13.47% 13.47% 0.04
1.01 - 1.40 4 7% 1.05 1.15 8.48% 0.68% 3.75% 0.06
1.41 - 5.00 8 13% 1.93 2.01 4.29% 4.46% 7.64% 0.40
>5.01 0 - - - - - - -
TOTAL 61 100 0.56 0.45 -24% -4.06% 42.83% 0.61

TABLE 11-11 ESMERALDA DEPOSIT SUMMARY OF STASTICS FOR 2015 INTER-LABRATORY CHECK SAMPLES

Table 11-10 and Table 11-11 summarize the performance of the inter-laboratory check samples across selected grade ranges.

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11.6.2.5   Opinions on Sampling, Security and Analysis Procedure

It is the opinion of the author that all sampling and QA/QC from the Esmeralda deposit is of a sufficient standard for use in these mineral resource and mineral reserve estimates.

There are several issues observed in the standard and blank performance, however, it is not expected that it will have a material impact on the performance of the mineral resource estimation.

11.6.2.6   Recommendations

It is recommended that several areas be improved in the future regarding the use of QA/QC from the performance of the drilling at the Esmeralda deposit:

11.6.3      UNION REEFS

Quality Assurance and Quality Control (QA/QC) procedures for Union Reefs include insertion of control samples (standards) and barren quartz flushes (BQF), inter-laboratory pulp checks and submission of sample duplicates. RC duplicates are produced from splitting the original sample straight from the rig mounted cone splitter. Diamond core duplicates are sent as a quarter core sample from the original sample interval. Blanks are generally inserted in the sample stream every 20 samples.

  Diamond % of total RC % of total
Samples 12,806 67.41% 4,418 64.63%
Blanks 621 3.27% 85 1.24%
BQFs 167 0.88% 231 3.38%
Duplicates 75 0.39% 163 2.38%
Repeats 3,083 16.23% 1,100 16.09%
Standards 408 2.15% 52 0.76%
Lab standards 979 5.15% 353 5.16%
Umpire lab repeats 406 2.14% 434 6.35%
Screen fire assay 31 0.16% 0 0.00%
Total QA/QC 6,191 32.59% 2,418 35.37%
Total Samples 18,997 100.00% 6,836 100.00%

TABLE 11- 12 RATE OF QA/QC SAMPLING FOR UNION REEFS RC AND DIAMOND DRILLING

11.6.3.1   Standards

Certified standards are submitted to the laboratory on a regular basis. A standard is inserted into every batch every 117 samples during the RC program and every 26 samples during the diamond drilling samples (or less). There were initially three standards across all ranges used during the RC drilling program and five standards across all ranges used during the diamond drilling program.

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Each standard for each drill type is charted chronologically to check for compliance and any progressive trends, which may be apparent. An Example of the chart used to chronologically check the standards is below.

Overall, for Union Reefs 408 standards were inserted for the 12,806 samples taken for the diamond with 52 standards inserted into the RC program total of 4,418 sample program. 979 lab standards were included in the diamond program with 353 lab standards in the RC program. This equates to around 1 in 13.6 samples for the diamond program being a standard.

Standard: ST48/9278
Recommended value: 4.55
Mean Result: 4.53
% diff std v RV: -0.4
Standard Deviation: 0.11
Number of Assays: 24
No > -2SD: 0
No > +2SD: 1
% + - SD: 95.8

TABLE 11-13 UNION REEFS STANDARD ST48/9278 COMPLIANCE TABLE

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11.6.3.2   Blanks

Blank materials included in the sample stream were derived from several sources including barren core, barren coarse rejects, crushed Bunbury Basalt (from Gannet Holding Pty Ltd), referred to in this report as “blank”). Blank results above 0.02ppm Au are queried and any issues resolved. Results are chronologically charted to visually check compliance. Below is an example of the blanks charts used.

In the Union Reefs program it was recognized in the early days that the blank material was returning higher results than was ideal. It was determined this was due to the high-grade gold samples previously sampled smearing across the sample. Therefore through the program a blank quartz flush (BQF) sample was always submitted after an expected high-grade intersection, this was then followed by a blank sample. This quartz flush improved the sample quality for the program.

A total of 622 blanks were taken for Union Reefs with 82.6% of sample results returning at or below 0.02ppm Au (Figure 11-18). This is actually a poor result but was the factor of inserting blanks after every higher grade hit. Once the BQF was regularly inserted the overall results of improved. A total of nearly 96% of all sample reults returned were below 0.1g/t Au, which is more acceptable. This issue needs constant monitoring as it could have a significant impact of the mineral resource estimate.

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11.6.3.3   Laboratory Duplicate Assays

Relative precisions have been used to analyze the precision of duplicate samples. The relative precision is a measure of dissimilarity, that is, if both distributions are exactly the same, this value will equal zero increases as the distributions become more dissimilar.

In this report, relative precision has been calculated using all data pairs for the ranges of below detection; <0.01ppm to 0.2ppm Au, 0.21 to 0.5ppm Au, 0.51 to 0.7ppm Au, 0.71 to 1ppm Au, 1.01 to 1.4ppm Au, 1.41 to 5ppm Au and >5ppm Au. This is to isolate the large conditional variance of errors associated with assay determinations near both lower and upper analytical detection limits and to selectively analyze results within these set ranges.

An example of the analysis tables is given in Table 11-14.

Type of Repeat Union Reefs DD NAL
FA50 Total Program
Mean Original Results 2.17
Mean Repeat Results 1.44
No of Assays 75
Sd: 5.56
Sum of Differences 54.62
Sum of Diff*Diff: 2288.13
Mean Difference 0.73
% Results within + or - 2 SD : 99
Results within 10% Precision Level 32
Ave. Absolute % Difference 111
% Assays Original < or = Repeat: 64

TABLE 11-14 UNION REEFS DIAMOND DUPLICATE ANALYSIS TABLE

CORRELATION Original vs Repeat
Range (ppm)
Combined 0.589
<0.20 0.411
0.21 - 0.50 -0.710
0.51 - 0.70 1.000
0.71 - 1.00 0.716
1.01 - 1.40 0.101
1.41 - 5.00 0.693
>5.01 0.071

TABLE 11-15 UNION REEFS DIAMOND PROGRAM DUPLICATE CORRELATION TABLE

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              total error  
                 
            bias Absolute  
                 
          bias Average Average    
R - CHART         % Diff. % Diff % Diff  
Range # of % of Mean Mean Between  Between  Between  
(ppm) Assays Total # Original Repeat Means Assays Assays S.D.
<0.20 46 61 0.06 0.21 -258.4 -99.25 122 0.94
0.21 - 0.50 4 5 0.36 0.97 -173.2 -238.94 256 1.43
0.51 - 0.70 2 3 0.53 0.18 66.7 66.7 67 0.49
0.71 - 1.00 4 5 0.77 1.18 -52.8 -50.43 69 0.81
1.01 - 1.40 4 5 1.25 1.60 -27.9 -28.04 74 1.21
1.41 - 5.00 9 12 2.68 1.83 31.8 37.62 52 1.56
>5.01 6 8 20.87 11.09 46.8 17.28 89 21.05
TOTAL: 75 100 2.17 1.44 33.6 -18.9 111 5.56

TABLE 11- 16 DUPLICATE R TABLE UNION REEFS DIAMOND PROGRAM

A chart (Figure 11-19) for each correlation range is produced to visually assess any correlation bias.

One hundred and sixty three RC duplicate samples were taken with 91 returning results below the detection limit. Fifty eight percent of all samples were within the 30% precision level, a reflection of the nuggetty nature of the deposit. Seventy-five diamond duplicate samples were taken with 20 returning results below the detection limit. Forty six percent of samples fell within the 30% precision level.

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11.6.3.4   Inter-Laboratory Repeats

Inter-laboratory repeats were taken for both RC and diamond drilling with pulp material sent to umpire lab ALS in Perth for assay. Results were compared to original assay results for each area and each umpire lab separately. An example of the tables and charts used to analyze each drill type and umpire lab is given in Table 11-17, Table 11-18, Table 11-19 and Figure 11-20.


Type of Repeat
Union Reefs DD NAL:
ALS June 2011 to End of
Program
Mean Original Results 14.94
Mean Repeat Results 7.84
No of Assays 186
Sd: 84.2.9
Sum of Differences 1321.92
Sum of Diff*Diff: 1314295.56
Mean Difference 7.11
% Results within + or - 2 SD : 99
Results within 10% Precision Level 37
Ave. Absolute % Difference 38
% Assays Original < or = Repeat: 56

TABLE 11-17 UNION REEFS DIAMOND NAL:ALS INTER- LABORATORY REPEAT SUMMARY TABLE

CORRELATION Original vs Repeat
Range (ppm)
Combined 0.928
<0.20 0.498
0.21 - 0.50 0.519
0.51 - 0.70 0.324
0.71 - 1.00 0.359
1.01 - 1.40 -0.133
1.41 - 5.00 0.830
 >5.01 0.611

TABLE 11-18 UNION REEFS DIAMOND INTER-LABRATORY REPEATS CORRELATION TABLE NTEL: ALS

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              total error  
                 
            bias Absolute  
                 
          bias Average Average    
R - CHART         % Diff. % Diff % Diff  
Range # of % of Mean Mean Between  Between Between   
(ppm) Assays Total # Original Repeat Means Assays Assays S.D.
<0.20 13 7 0.13 -0.02 114.5 -162.99 200 0.08
0.21 - 0.50 24 13 0.39 -0.05 112.8 -8.99 26 0.34
0.51 - 0.70 11 6 0.69 -0.07 109.9 -10.51 29 0.31
0.71 - 1.00 17 9 0.86 -0.01 100.9 0.08 35 0.45
1.01 - 1.40 14 8 1.26 -0.01 101.0 -2.15 18 0.34
1.41 - 5.00 54 29 3.14 -0.33 110.6 -11.10 21 0.95
>5.01 53 28 23.34 25.33 -8.5 -2.05 29 158.98
TOTAL: 186 100 14.94 7.84 47.6 -11.5 38 84.29

TABLE 11-19 NAL:ALS INTER-LABRATORY REPEAT R TABLE UNION REEFS DIAMOND PROGRAM

11.6.3.5   Screen Fire Assays

A selection of samples for Screen Fire assays were taken for the Union Reefs area to assess the variable character of the mineralization and to assess why samples logged with visible gold were returning unexpectedly low gold results. Thirty one samples were taken with the results (Figure 11-21) of the screen fire assays demonstrating no systematic underestimation of grade. The samples showed high variability across all grade ranges.

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11.6.3.6   Opinions on Sampling, Security and Analysis Procedure

Due to the nugget nature of the Prospect mineralization it is important that the quality of the sampling and analysis is monitored closely. The use of predominately diamond drilling in the mineral resource estimate will assist with any estimation process. There were several issue noted during the program which were identified and rectified before data was entered into the database. Another advantage of using diamond drilling was that if there were issues with the sampling the second half of core could be used to validate data.

Some of the issues noted during the drilling program are noted below:

There are some errors in the datasets. Typographical errors, wrong standards, recorded/sent to the lab, obvious swaps in the databases. These errors are collaborative from both the laboratories and the database operator.

   

ALS and NAL report lab standards while NTEL do not.

   

Outright errors should not be appearing in the database (standards and replicates).

   

Proper control charting methods should be applied to fire assay batches that indicate standards outside proper control limits.

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The use of blanks in the Prospect mineralization is incredibly important and needs to be monitored regularly.

   

A significant amount of historic data exists in both digital and hard copy format. This material needs to be checked against the database to ensure accuracy. Reviews have been completed on previous QA/QC reports giving the Author confidence in the previous work completed.

It is of the opinion of the Authors that the sampling preparation, analysis and security procedures are all adequate for use in these mineral resource and mineral reserve estimates.

11.6.3.7   Recommendations

The results from the QA/QC analysis of drilling have indicated a good level of confidence in assay grades for use in the mineral resource model. The following recommendations for improvements in the current procedures are:

Regular site visits to the laboratory as required.

   

Monitor the performance of the Blanks and Quartz flushes to ensure mineralization is not smeared through the mineralization.

   

An immediate follow up with the laboratory when controls fail.

   

Inter-laboratory repeats to meet or exceed a rate of 1:20 to original samples. These are to be performed on a regular basis and not at the end of a program.

   

Assay results to be thoroughly assessed for errors prior to loading.

   

Conducting an analysis on barren core that is re-used to serve as blanks for future batches.

   

Regular tracking of QA/QC compliance.

   

Monthly reporting of QA/QC performance and data.


11.6.4      INTERNATIONAL DEPOSIT

Quality Assurance and Quality Control (QA/QC) procedures for the International drilling programs include insertion of control samples (standards) and barren quartz flushes (BQF), inter-laboratory pulp checks and submission of sample duplicates. RC duplicates are produced from splitting the original sample straight from the rig mounted cone splitter. Diamond core duplicates are sent as a quarter core sample from the original sample interval. Blanks are generally inserted in the sample stream every 20 samples.

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  International
Diamond
% of total X: Samples
 
Samples 440 62.86% 1:1
Blanks 26 3.71% 1:17
BQFs 0 0% 0
Duplicates 9 1.29% 1:49
Repeats 149 21.29% 1:3
Standards 16 2.29% 1:27
Lab standards 33 4.71% 1:13
Umpire lab 27 3.86% 1:16
Screen fire 0 0 0
Total QA/QC 260 37.14% 1:1.7
Total 700 100.00%  

TABLE 11-20 RATE OF QA/QC SAMPLING FOR INTERNATIONAL DEPOSIT DIAMOND DRILLING

11.6.4.1   Standards

Certified standards are submitted to the laboratory on a regular basis. A standard is inserted into every batch every 26 samples during the diamond drilling samples (or less). There were five standards across all ranges used during the diamond drilling program.

Each standard for each drill type is charted chronologically to check for compliance and any progressive trends, which may be apparent. An Example of the chart used to chronologically check the standards is below.

Overall, for International 16 standards were inserted for the 440 samples taken for the sample program. 33 lab standards were included in the diamond program. This equates to around 1 in 9 samples for the diamond program being a standard.

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Standard: ST08/8225
Recommended value: 0.33
Mean Result: 0.32
% diff std v RV: -3.8
Standard Deviation: 0.01
Number of Assays: 4
No > -2SD: 0
No > +2SD: 0
% + - 2SD: 100.0

TABLE 11-21 INTERNATION STANDARD ST08/8225 COMPLIANCE TABLE

11.6.4.2   Blanks

Blank materials included in the sample stream were derived from several sources including barren core, barren coarse rejects, crushed Bunbury Basalt (from Gannet Holding Pty Ltd, referred to in this report as “blank”). Blank results above 0.02ppm Au are queried and any issues resolved. Results are chronologically charted to visually check compliance. Below is an example of the blanks charts used.

As only a limited number of samples were submitted for the International drill program only a limited number of blanks we submitted and reported. In the results seen below a few blanks came back with elevated grades with 4 returning results higher than 0.03g/t Au, but not were higher than 0.1g/t Au. It is recommended that this is continually monitored to ensure higher grade intercepts do not contaminate subsequent samples.

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11.6.4.3   Laboratory Duplicate Assays

Relative precisions have been used to analyze the precision of duplicate samples. The relative precision is a measure of dissimilarity, that is, if both distributions are exactly the same, this value will equal zero increases as the distributions become more dissimilar.

Newmarket Gold reports relative precision as being calculated using all data pairs for the ranges of below detection <0.01ppm to 0.2ppm Au, 0.21 to 0.5ppm Au, 0.51 to 0.7ppm Au, 0.71 to 1ppm Au, 1.01 to 1.4ppm Au, 1.41 to 5ppm Au and >5ppm Au. This is to isolate the large conditional variance of errors associated with assay determinations near both lower and upper analytical detection limits and to selectively analyze results within these set ranges.

An example of the analysis tables for each deposit and drill type is given in Table 11-22.

Type of Repeat International DD
NAL Duplicate FA50
Mean Original Results 0.24
Mean Repeat Results 0.25
No of Assays 8
Sd: 0.12
Sum of Differences -0.13
Sum of Diff*Diff: 0.10
Mean Difference -0.02
% Results within + or - 2 SD : 34
Results within 10% Precision Level 15
Ave. Absolute % Difference 38
% Assays Original < or = Repeat: 24

TABLE 11- 22 INTERNATIONAL DIAMOND DUPLICATE ANALYSIS TABLE

 Correlation Original vs Repeat
Range (ppm)
Combined 0.959
<0.20 0.694
0.21 - 0.50 -
0.51 - 0.70 -
0.71 - 1.00 -
1.01 - 1.40 -
1.41 - 5.00 -
>5.01 -

TABLE 11-23 INTERNATION DUPLICATE TABLE

A chart for each correlation range is produced to visually assess any correlation bias.

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Eight diamond duplicate samples were taken with three being below the detection limit. Seventy five percent of all samples were within the 30% precision level, a reflection of the disseminated nature of the deposit as well as the small number of samples taken.

11.6.4.4   Inter-Lab Repeats

Inter-lab repeats were taken for the International diamond drilling with pulp material sent to umpire lab ALS in Perth for assay. Results were compared to original assay results for each area and each umpire lab separately. An example of the tables and charts used to analyze each drill type and umpire lab is given below.

TYPE OF REPEAT: International DD
NAL: ALS 2012
Mean Original Results 1.15
Mean Repeat Results 1.30
No of Assays 25
Sd: 0.65
Sum of Differences -3.76
Sum of Diff*Diff: 10.16
Mean Difference -0.15
% Results within + or - 2 SD : 96
Results within 10% Precision Level 56
Ave. Absolute % Difference 20
% Assays Original < or = Repeat: 56

TABLE 11-24 INTERNATION DIAMOND INTER-LAB REPEAT SUMMARY TABLE NAL:ALS

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Correlation Original vs Repeat
Range (ppm)
Combined 0.928
<0.20 1.000
0.21 - 0.50 0.713
0.51 - 0.70 0.824
0.71 - 1.00 0.996
1.01 - 1.40 0.455
1.41 - 5.00 0.809
>5.01 -

TABLE 11-25 INTERNATION DIAMOND NTEL:ALS INTER- LAB REPEATS CORELATION TABLE

              total error  
                 
            bias Absolute  
                 
          bias Average Average    
R -                
CHART         % Diff. % Diff % Diff  
   Range # of % of Mean Mean Between  Between  Between  
   (ppm.) Assays Total # Original Repeat Means Assays Assays S.D.
<0.20 2 8 0.15 0.21 -41.4 -49.24 49 0.09
0.21 - 0.50 7 27 0.33 0.37 -11.5 -10.18 17 0.09
0.51 - 0.70 4 15 0.59 0.69 -16.0 -15.07 20 0.18
0.71 - 1.00 4 15 0.87 0.88 -0.3 0.00 6 0.04
1.01 - 1.40 3 12 1.27 1.19 6.3 6.53 18 0.31
1.41 - 5.00 6 23 2.80 3.34 -19.4 -15.15 21 1.40
>5.01 0 0 - - - 0.00 - 0.00
TOTAL : 26 100 1.15 1.30 -13.1 -1.0 20 0.64

TABLE 11-26 NAL:ALS INTER-LABS REPEAT R TABLE INTERNATIONAL DIAMOND PROGRAM

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11.6.4.5   Opinions on Sampling, Security and Analysis Procedure

The two laboratories used by Newmarket Gold for International analytical work offer different preparation techniques with the 50g fire assay by NAL and a 30g fire assay by ALS. The following summarizes findings with respect to assay work from the three independent laboratories:

It is of the opinion of the Authors that the sampling preparation, analysis and security procedures are all adequate for use in these mineral resource and mineral reserve estimates.

11.6.4.6   Recommendations

The results from the QA/QC analysis of drilling have indicated a good level of confidence in assay grades for use in the mineral resource model. The following recommendations for improvements in the current procedures are:

11.6.5      RISING TIDE AND YAM CREEK

Quality Assurance and Quality Control (QA/QC) procedures for the Rising Tide and Yam Creek drilling programs include insertion of control samples (standards) and barren quartz flushes (BQF), inter-laboratory pulp checks and submission of sample duplicates. RC duplicates are produced from splitting the original sample straight from the rig mounted cone splitter. Diamond core duplicates are sent as a quarter core sample from the original sample interval. Blanks are inserted in the sample stream every 20 samples.

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Rising Tide RC % of
total
Yam Creek RC % of
total
Yam Creek
Diamond
% of
total
Samples 6,568 69.42% 2,380 72.47% 361 77.80%
Blanks 146 1.54% 52 1.58% 11 2.37%
BQFs 0 0.00% 0 0.00% 0 0.00%
Duplicates 278 2.94% 101 3.08% 0 0.00%
Repeats 1,415 14.96% 490 14.92% 66 14.22%
Standards 262 2.77% 88 2.68% 10 2.16%
Lab standards 475 5.02% 173 5.27% 16 3.45%
Umpire lab repeats 317 3.35% 0 0.00% 0 0.00%
Screen fire assay 0   0   0 0.00%
Total QA/QC 2893 30.58% 904 27.53% 103 22.20%
  Total  Samples  9,461 100.00%    3,284  100.00%   464   100.00%

TABLE 11-27 RATE OF QA/QC SAMPLING FOR RISING TIDE & YAM CREEK RC AND DIAMOND DRILLING

11.6.5.1   Standards

Certified standards are submitted to the laboratory on a regular basis. A standard is inserted into every batch every 117 during the RC program and every 26 samples during the diamond drilling samples (or less). There were initially three standards across all ranges used during the RC drilling program and five standards across all ranges used during the diamond drilling program.

Each standard for each drill type is charted chronologically to check for compliance and any progressive trends, which may be apparent. An example of the chart used to chronologically check the standards is below.

Overall, for Rising Tide, 262 standards were inserted for the 6,568 samples taken for the RC program. A total of 317 lab standards were inserted.

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   Standard: ST43/7370
   Recommended value g/t Au: 3.37
   Mean Result g/t Au: 3.32
   % diff std v RV: -1.5
   Standard Deviation: 0.06
   Number of Assays: 6
   No > -2SD: 0
   No > +2SD: 0
%=-2SD:  100.0

TABLE 11-28 RATE OF QA/QC SAMPLING FOR RISING TIDE & YAM CREEK RC AND DIAMOND DRILLING

11.6.5.2   BLANKS

Blank materials included in the sample stream were derived from several sources including barren core, barren coarse rejects, crushed Bunbury Basalt (from Gannet Holding Pty Ltd, referred to in this report as “blank”). Blank results above 0.02ppm Au are queried and any issues resolved. Results are chronologically charted to visually check compliance. Below is an example of the blanks charts used.

A total of 146 blanks were taken for Rising Tide with 99.3% of samples returning values at or below 0.02ppm Au. A total of 52 blanks were taken for the Yam Creek RC program with all samples returning at or below 0.02ppm. A total of 11 blanks were taken for the diamond program with one sample recording above 0.02ppm Au (0.03ppm) .

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11.6.5.3   Laboratory Duplicate Assays

Relative precisions have been used to analyze the precision of duplicate samples. The relative precision is a measure of dissimilarity, that is, if both distributions are exactly the same, this value will equal zero increases as the distributions become more dissimilar.

In this report, relative precision has been calculated using all data pairs for the ranges of below detection (<0.01ppm) to 0.2ppm Au, 0.21 to 0.5ppm Au, 0.51 to 0.7ppm Au, 0.71 to 1ppm Au, 1.01 to 1.4ppm Au, 1.41 to 5ppm Au and >5ppm Au. This is to isolate the large conditional variance of errors associated with assay determinations near both lower and upper analytical detection limits and to selectively analyze results within these set ranges.

An example of the analysis tables for each deposit and drill type is given in Table 11-29, Table 11-30, Table 11-31 and Figure 11-28:

ANALYSIS OF REPEAT SAMPLES  
   
   Rising Tide NAL FA50 Total  
TYPE OF REPEAT: Program
MEAN ORIGINAL RESULTS  
g/t Au: 0.25
MEAN REPEAT RESULTS g/t  
Au: 0.17
NO of ASSAYS: 278
Sd: 1.46
SUM of DIFFERENCES: 22.62
SUM of DIFF*DIFF: 592.87
MEAN DIFFERENCE: 0.08
% RESULTS within + or - 2 SD : 99
RESULTS within 30%  
PRECISION LEVEL: 71
AVE. ABSOLUTE %  
DIFFERENCE: 35
% ASSAYS ORIGINAL < or = REPEAT: 84

TABLE 11-29 RISING TIDE RC DUPLICATE ANALYSIS TABLE

CORRELATION Original vs Repeat
Range (Au ppm)
Combined 0.484
<0.20 0.821
0.21 - 0.50 0.826
0.51 - 0.70 -0.329
0.71 - 1.00 -
1.01 - 1.40 0.670
1.41 - 5.00 0.710
>5.01 1.000

TABLE 11-30 RISING TIDE RC DUPLICATE CORRELATION TABLE

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              total error  
                 
            bias Absolute  
                 
          bias Average Average    
R - CHART         % Diff. % Diff % Diff  
Range # of % of Mean Mean Between  Between Between   
(Au ppm.) Assays Total # Original   Repeat Means Assays Assays S.D.
<0.20 250 90 0.02 0.00 101.4 -17.32 36 0.02
0.21 - 0.50 12 4 0.33 -0.02 104.7 -1.17 27 0.11
0.51 - 0.70 4 1 0.70 -0.11 116.1 -20.49 36 0.28
0.71 - 1.00 1 0 0.41 0.34 17.1 45.33 45 -
1.01 - 1.40 3 1 1.16 0.03 97.1 2.62 5 0.11
1.41 - 5.00 6 2 4.41 -0.86 119.4 -16.46 40 2.82
>5.01 2 1 1.87 14.01 -651.2 86.55 87 23.50
TOTAL: 278 100  0.25  0.17  33.0  -15.5  35  1.46 

TABLE 11-31 RISING TIDE DUPLICATE TABLE

 A CHART FOR EACH CORRELATION RANGE IS PRODUCED TO VISUALLY ASSESS ANY CORRELATION BIAS.


Rising Tide

A total of 278 duplicate samples were taken for Rising Tide with 167 returning values below the detection limit. A total of 71% of duplicate samples were within the 30% precision limit.

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Yam Creek

A total of 101 duplicate samples were taken from Yam Creek with 53% falling within 30% of the precision level, reflecting the coarse gold nature of the deposit. Forty-four of the original samples returned values below the detection limit. No duplicates were taken for the two Yam Creek diamond holes drilled.

11.6.5.4   Inter-Lab Repeats

Inter-lab repeats were taken for both RC and diamond drilling with pulp material sent to umpire lab ALS in Perth for assay. Results were compared to original assay results for each area and each umpire lab separately. An example of the tables and charts used to analyze each drill type and umpire lab is given in Tables 11-32, Table 11-33, Table 11-34 and Figure 11-29.

   ANALYSIS OF INTERLAB REPEAT SAMPLES
   
   TYPE OF REPEAT: Rising Tide RC NAL FA50 ALS Au-AA25
   MEAN ORIGINAL RESULTS g/t Au: 1.17
   MEAN REPEAT RESULTS g/t Au: 1.12
   NO of ASSAYS: 219
   Sd: 0.01
   SUM of DIFFERENCES: 10.51
   SUM of DIFF*DIFF: 0.03
   MEAN DIFFERENCE: 0.05
   % RESULTS within + or - 2 SD : 36
   RESULTS within 5% PRECISION LEVEL: 15
   AVE. ABSOLUTE % DIFFERENCE: 28
  % ASSAYS ORIGINAL < or = REPEAT: 45

TABLE 11-32 RISING TIDE RC INTER-LAB REPEAT SUMMARY TABLE NTEL:ALS

CORRELATION Original vs Repeat

Range (Au ppm)
Combined 0.718
<0.20 0.724
0.21 - 0.50 0.706
0.51 - 0.70 0.249
0.71 - 1.00 0.427
1.01 - 1.40 0.044
1.41 - 5.00 0.788
>5.01 -0.149

TABLE 11-33 RISING TIDE INTER-LAB REPEATS CORRELATION TABLE NTEL:ALS

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              total error   
            bias Absolute  
                 
          bias Average Average    
   R - CHART         % Diff. % Diff % Diff  
Range # of % of Mean Mean Between  Between Between    
(ppm.) Assays Total # Original Repeat Means Assays Assays S.D.
<0.20 30 14 0.13 0.14 -5.2 -1.8 30 0.06
0.21 - 0.50 51 23 0.35 0.34 5.0 5.9 20 0.10
0.51 - 0.70 30 14 0.60 0.71 -18.8 -17.1 44 0.73
0.71 - 1.00 25 11 0.86 0.77 10.9 11.8 30 0.37
1.01 - 1.40 24 11 1.17 1.31 -12.0 -12.8 29 0.57
1.41 - 5.00 54 25 2.35 2.32 1.2 0.8 21 0.70
>5.01 5 2 7.80 5.26 32.6 28.1 63 7.12
TOTAL 219 100 1.17 1.12 4.1 -0.1 28 1.08 

TABLE 11-34 NTEL: ALS INTER-LAB REPEAT R TABLE RISING TIDE RC PROGRAM


11.6.5.5   Opinions on Sampling, Security and Analysis Procedures

The laboratories used by Crocodile Gold for the Burnside deposits offer different preparation techniques with the 30g fire assay by NTEL and a 30g fire assay by ALS and 50g fire assay for NAL. The following summarizes findings with respect to assay work from the three independent laboratories:

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It is of the opinion of the Authors that the sampling preparation, analysis and security procedures are all adequate for use in these mineral resource and mineral reserve estimates.

11.6.5.6   Recommendations

The results from the QA/QC analysis of drilling have indicated a good level of confidence in assay grades for use in the mineral resource model. The following recommendations for improvements in the current procedures are:

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12 DATA VERIFICATION

Newmarket Gold utilize specialized industry computer software to manage its drill hole and assay database and employ dedicated personnel to manage the database and apply appropriate QA/QC procedures to maintain the integrity of the data. Data is assessed for errors against standards and blanks prior to loading into the AcQuire™ database software. Data is then spatially assessed in commercially available mining software package Surpac™ for any other questionable results.

Previously, independent consultants have completed various database checks, which have not identified any reportable errors, which would have raised any concerns about the integrity of the data. During the preparation of this technical report, which has included search and lookup of assay results, generation of plans and sections and estimation of mineral resources, the Authors did not encounter any difficulties with the database; hence the Authors believe the historical data/database has been verified to a sufficient level to permit its use and have confidence in its reliability.

Wherever possible the Company has also conducted on ground checks of data, this includes the resurveying of historic drill collars and previously mined open pits. The checking of the open pits has involved the use of a surveyor with a depth sounder to test the bottom of the pit against previous pit pickups as all previously mined pits contain some surface water. This was done to ensure an accurate depletion of the mineral resource.

During the past 3-4 years a large amount of time and money reviewing all historic data in both hard and soft copy forms. This has given the Company a much better understanding of the original data that is available for cross checking and review.

In conjunction with the review of historic data, a detailed review has commenced on the QA/QC results for historic drill campaigns on currently reported mineral resources. This has included a review of the assay results and QA/QC processes for the Western Arm, Bon’s Rush and Kazi mineral resources. While these deposits were drilled more than 10 years ago there is sufficient data available to check against the information stored in the Company database. While more work is required to validate this data, through twinned holes and resampling of existing diamond core, it is of the opinion of the Author that this drilling data fulfils the requirements for reporting mineral resources. Further work is planned to add more confidence to the historic drilling data. This is the case for all Newmarket Gold mineral deposits, regardless of the generation of data used in the estimation process.

There were no limitations or failure by the Authors to verify the data in this technical report. In the opinion of the Authors such data is adequate for the purposes of this technical report.

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13 MINERAL PROCESSING AND METALLURGICAL TESTING

13.1 UNION REEFS PROCESSING FACILITY

Acacia Resources Ltd, an entity spun out of the Shell/Billiton group, commissioned the Union Reefs facility in 1994. AngloGold Australia Ltd acquired the mine through its successful takeover bid for Acacia in December 1999. Until late 2000, Union Reefs formed one half of what was known as AngloGold’s Pine Creek operations, which also included the smaller Brocks Creek project.

The Union Reefs CIL treatment plant was commissioned with a throughput capacity of 1.25Mtpa. It included a gravity circuit to extract coarse gold. It was designed by Kinhill and commissioned in December 1994. JR Engineering carried out an upgrade in 1998 that involved the installation of a tertiary crusher, second ball mill, and two additional leach tanks.

The plant currently has a maximum capacity (depending on mineralization type) for 2.5Mtpa and is configured with three-stage crushing and two single-stage milling circuits. Prior to the plant being placed on care and maintenance in 2003, the milling rate at Union Reefs was typically 335tph at a P80 of 75µm. Plant availability was typically 96-98%.

In August 2004, and before they were acquired by GBS Australia, the Burnside JV partners purchased the Union Reefs Gold project for A$4 million on a walk-in, walk-out basis.

In August 2006, GBS Australia re-commissioned the Union Reefs plant on the larger of the two mills while leaving the other smaller ball mill in a care and maintenance state. The first source of feed material was low-grade stockpiles from Cosmo Mine and an alluvial tailings deposit from the Union Reefs site. Following commissioning, mineralization was sourced from a blended mix of oxidized and fresh underground and open pit mines.

In June 2010 Crocodile Gold announced commercial production for the Union Reefs plant, which has continued to operate since commissioning.

13.1.1      UNION REEFS PLANT OPERATIONS

Mineralization is broken to minus one meter by blasting. Any larger rocks produced from the blasting process are subsequently broken to suitable size by rock breaker. Run of Mine (“ROM”) ore is transported by truck directly to the ROM feed bin or the ROM stockpile for storage before subsequent processing.

ROM ore is crushed at a rate of up to 2.5Mtpa in a three stage crushing circuit incorporating a primary jaw crusher operating in open circuit and a secondary and tertiary cone crushers operating in closed circuit with a double deck banana screen. Crushing circuit product, at a nominal size of 12mm is conveyed to the grinding circuit via the Fine Ore Bin (“FOB”).

The FOB, with a live capacity of 2,500 tonnes, provides a buffer of approximately seven to eight hours between the crushing and grinding circuits. Ore is reclaimed via a slot feeder at a variable rate and is conveyed to the grinding circuit.

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As the crushing circuit capacity exceeds that of the milling circuit, crushed ore is stockpiled and fed back into the grinding circuit using a front end loader into an emergency feed hopper and feeder arrangement.

Crushed ore is ground in the grinding circuit consisting of two ANI single stage rubber lined ball mills operating in closed circuit with a nest of Warman cyclone classifiers. A proportion of ball mill discharge is directed to the gravity circuit incorporating four Knelson concentrators, two per mill. Knelson concentrator tailings report back to the mill discharge stream whilst the concentrated coarse gold is sent to the gold room for further processing.

The remainder of the mill discharge and the Knelson concentrator tailings are pumped to cyclone classifiers. The coarse underflow fraction reports back to the ball mill for further grinding whilst the fine overflow fraction (P80 75-106µm) gravitates to a single high rate thickener for density control before being pumped to the first of two leach tanks.

Cyanide is added into the two leach tanks to put the gold into solution before the material gravitates into the CIL circuit. High purity oxygen is added into the leach tanks from the Pressure Swing Absorption plant on site.

The CIL circuit, comprising seven leach/adsorption contactors is gravity fed through open launders. All tanks are agitated and aerated and are fitted with hollow shaft mechanical agitators. Barren slurry exits from the last CIL tank and gravitates to the residue treatment circuit. Activated carbon is pumped counter current to the process slurry to recover gold from solution, achieving the highest gold on carbon loading in CIL tank 1. Carbon from tank 1 is pumped to the elution circuit.

Tailings slurry is pumped to the Crosscourse pit tailings facility. Process water is recycled back from the Crosscourse pit.

The loaded carbon recovered from the CIL circuit is screened to remove pulp and subjected to a desorption stage (split AARL, 4t capacity) to remove gold as an auriferous caustic-cyanide solution from which the gold is recovered by electro winning. The stripped carbon is reactivated in a vertical kiln and returned to the CIL circuit for reuse.

Gravity gold recovered in the Knelson concentrators is periodically discharged to a settling cone located in the gold room. The gold is then intensively leached in an Acacia reactor and the pregnant solution electro-won onto steel wool.

The electro-won gold and the gravity won gold are calcined in an electric oven and smelted separately in a gas-fired furnace into doré bullion. Bars are stamped for identification and dispatched via security service to AGR at Perth International Airport.

Water is supplied from various dams strategically located to maximize catchment of run-off drainage. A dam constructed on the nearby McKinlay River provides make up water if required.

The whole plant is controlled by a CITECT process control system over an Allen Bradley PLC.

A schematic flow sheet of the plant, as currently configured, is shown below.

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13.2 METALLURGY

Sampling for metallurgical test-work is completed on an as needs basis following the direction given by the Company’s metallurgists. This may mean detailed sampling of specific ore types and at others it may be required to supply a sample that represents the overall ore type. At all times the sampling is done using procedures required by the metallurgical team and the samples would represent the material required.

While test-work is required to determine the processing ability of all ore types, the test-work is also designed to determine if any deleterious elements are within the various ore types. From all test-work completed to date on the various deposit no deleterious elements have been identified that would have a significant effect on the economic extraction of gold.

13.2.1      COSMO MINE METALLURGICAL TEST WORK

Date Report No. Description
2011 Ammtec A13451 Bond Work Index, Abrasion Index, Leach/Gravity Recovery, Oxygen Uptake, Mineralogy
2012 Ammtec A13605 Confirm effect of preg-robbing using various ratios of F10 Fault carbonaceous shale. Head assays, Direct and CIL cyanidation leach test work
2012 Ammtec A14523 Head analysis, gravity/cyanidation leach test, gravity/CIL leach test work
2014? ALS A16107 Carbon composite sample analysis, head analysis, oxygen uptake, gravity/direct cyanidation leach test work

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Date Report No. Description
2014 ALS A15850 Head Assays, Gravity and Direct Cyanidation Leach Extraction and preg- robbing tests, Oxygen uptake test work
2015 ALS A16598 Bond Abrasion, Grind Establishment, Gravity/Direct Cyanidation and Oxygen Uptake test work

TABLE 13-1 SUMMARY OF REPORTS AVAILABLE FOR COSMO MINE METALLURGICAL TEST WORK

Since taking over the NT properties in 2010 the Company’s metallurgists have requested that ALS Metallurgy (formerly ALS Ammtec) conduct various programs of metallurgical test work on samples originating from the Cosmo Mine and other gold deposits in the area. A summary of that test work follows:

13.2.1.1   AMMTEC Study A13451- 2011

A program of metallurgical test work was carried out on two samples of gold mineralization from the Cosmo Mine. Salient test results are summarized below:

Head Assays

Comprehensive head analysis was carried out on a sub-sample of the Mill Feed Sample.

Analyte Unit Sample –GFF001 Sample CP011
Au Fire Assay g/t 6.15 9.66
Au Fire Assay (duplicate) g/t 5.5 7.61

TABLE 13-2 HEAD ASSAY RESULTS FOR COSMO MINE SAMPLES

Bond Ball Mill Work Index Determination

Each sample was tested using the standardized procedure detailed by F.C. Bond to determine the Bond Ball Mill Work Indices of the samples at a closing screen size of 106µm.

Sample Micrometers Gbp
(g/rev)
Test Aperture
Pi (•m)
Bond Ball Mill Work
Index (kWh/t)
F80 P80
GFG001 3001 81 0.752 106 22.9
CP011 2925 78 0.839 20.4

TABLE 13-3 COSMO MINE RESULTS OF BOND WORK INDEX TEST WORK

Bond Abrasion Index Determination

Each sample was tested to determine the abrasion index value using the standard procedure developed by F.C. Bond. A summary of results is presented below:

Sample Feed Particle
Size (mm)
Bond Abrasion
Index (Ai)
GFG001 -19.0+12.7 0.2136
CP011 0.2535

TABLE 13-4 COSMO MINE BOND ABRASION INDEX RESULTS

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Gravity Separation/Cyanidation Time Leach Test Work

A gravity separation/direct cyanidation leach test was conducted on the samples, at the nominated grind P80 80µm.

Composite Identity Test No. Grind Size
P80 (µm)
Direct Cyanidation Extraction
% Au/hr
@ Hours
Consumption
(kg/t)
Gravity
%
4 8 12 24 36 Lime NaCN
CFG001 MA102 80 63.6 91.6 94.7 95.5 95.3 96.3 0.79 0.83
CP001 MA103 75.5 91.8 94.4 94.4 94.4 94.6 1.21 0.65

TABLE 13-5 GRAVITY/DIRECT CYANIDATIOIN LEACH TEST WORK ON COSMO ORE

Direct Cyanidation Time Leach Test work

A single direct cyanidation time leach test was conducted on the samples, at the nominated grind of P80 80µm.

Composite Identity Test No. Grind Size
P80 (µm)
% Au Direct Cyanidation Extraction % Au/hr
@ Hours
Consumption
(kg/t)
 2 4 8 12 24 36 Lime NaCN
CFG001 MA114 80 48.3   72.0 87.3 91.4 88.6 95.8 0.95 0.76
CP001  MA115 53.5   75.5 93.3 97.5 98.1 97.5 1.22 0.83

TABLE 13-6 DIRECT CYANIDATION TIME LEACH RESULTS ON COSMO ORE

Mineralogy

Quantitative automated mineralogical investigations were conducted on gravity concentrates and tailing for each composite. The small number of detected grains of interest doesn’t allow a comprehensive characterization of gold-silver mineralization of the analyzed samples.

 Summary of the Detected Gold-Silver Minerals in the Concentrates 
Product No. of
Particles
No. of
Grains
Dominant
Mineral Phase
Dominant
Liberation Type by Mass
Qualitative
Sizing
CPO11
Knelson Con
2 2 (Argentian) Native
Gold
Free (Liberated) Coarse
CGF001
Knelson Con
1 1 Argentian
Native Gold
Encapsulated Very Fine

TABLE 13-7 MINERALOGY OF COSMO ORE

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Oxygen Uptake Rate Determination

A 2.0kg sub-sample of each sample was ground to P80 of 80µm. The freshly milled slurry sample was utilized to determine the oxygen uptake characteristics of the sample. The test procedure described by G.M. Fraser was utilized.

Time
(hours)
Oxygen Uptake Rate* (mg/l/min)
GFG001 CP011
 0** -0.1770 -0.0505
1 -0.0723 -0.0694
2 -0.1831 -0.0470
3 -0.0507 -0.0726
4 -0.1341 -0.0469
5 -0.0845 -0.0520
6 -0.0828 -0.0328
24 -0.0371 -0.0311

* Ambient temperature **Baseline data prior to aeration

TABLE 13-8 RESULTS OF OXYGEN UPTAKE TESTING COSMO MINE ORE

13.2.1.2   AMMTEC Study A13605 - 2012

A program of metallurgical test work was carried out on composites created from ore samples, originating from Cosmo Mine. Two carbonaceous shale samples were combined with a retrieved Cosmo sample (ALS Ammtec Test Program No. A13451), which displayed varying amounts of carbon.

Head Assays

Sample Au
(g/t)
Au Repeat
(g/t)
CTotal
(%)
COrganic
(%)
A72026 - Fault 10 Shale 0.09 0.08 4.89 4.62
A72027 – Cosmo 3.95 - 5.34 4.17
5% Fault 10 Shale/95% Cosmo 8.39 - 1.26 0.93
10% Fault 10 Shale/90% Cosmo 15.40 - - -
5% A72027/95% Cosmo 8.85 - 1.29 0.93
10% A72027/90% Cosmo 6.87 7.65 - -
25% Fault 10 Shale/75% Cosmo 4.92 4.80 2.19 1.98
25% A72027/75% Cosmo 5.04 4.63 2.22 1.86

TABLE 13-9 HEAD ASSAYS FOR COSMO MINE SAMPLES

Direct and CIL Cyanidation Test work

Each composite was submitted for Direct and CIL cyanidation test work at the received grind size.

Composite Identity Test
No.
Test Type % Au
Extraction
Au Residue
(ppm)
Consumption
(kg/t)
Lime NaCN
5% Fault 10 Shale/ MA393 Direct 93.15 0.47 0.69 0.58
95% Cosmo MA397 CIL 95.63 0.38 0.61 0.88

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Composite Identity Test
No.
Test Type % Au
Extraction
Au Residue
(ppm)
Consumption
(kg/t)
Lime NaCN
10% Fault 10 Shale/
90% Cosmo
MA394 Direct 93.91 0.48 0.70 0.47
MA398 CIL 95.14 0.39 0.54 0.81
5% A72027/
95% Cosmo
MA395 Direct 93.56 0.51 0.79 0.76
MA399 CIL 94.79 0.43 .71 0.76
10% A72027/
90% Cosmo
MA396 Direct 92.24 0.52 0.96 0.65
MA400 CIL 94.75 0.40 0.71 0.83
25% Fault 10 Shale/
75% Cosmo
MA437 Direct 91.68 0.58 0.89 0.44
MA439 CIL 95.48 0.28 0.77 1.02
25% A7027/
75% Cosmo
MA438 Direct 85.40 0.74 0.92 0.49
MA440 CIL 94.99 0.33 0.72 1.15

TABLE 13-10 DIRECT AND CIL CYANIDATION TESTWORK RESULTS FOR COSMO MINE ORE

13.2.1.3   AMMTEC Study A14523 - 2012

A program of metallurgical (preliminary extraction) test work was carried out on a gold composite sample originating from the Cosmo Mine.

Head Assays

Element Unit Assay
Au1 g/t 2.25
Au2 g/t 2.18
Ag ppm 0.80
As ppm 5070
CTOTAL % 2.64
CORGANIC % 2.31
Cu ppm 164
Fe % 10.10
STOTAL % 5.88
SSULPHIDE % 5.20
SiO2 % 49
Zn ppm 166

TABLE 13-11 HEAD ASSAY RESULTS FOR COSMO MINE ORE

The gold grades suggest the presence of coarse-grained gold in the mineralization. Base metals (Cu, Ni and Zn) are present in moderate concentrations, limiting the possibility of excess cyanide consumption through preferential complexing with these metals. The high content of arsenic suggests the probability that the gold can be locked within any arsenic complex mineralization. The presence of organic/graphitic (2.31%) suggests that preg-robbing can occur during the cyanidation leach process. Additionally a high content of sulphur sulphide can indicate the possibility of excess lime and cyanide consumption.

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Gravity/Cyanidation Leach Test work

Gravity separation and subsequent cyanidation leach test work was carried out on a representative sub-sample of the Cosmo Mine Composite at a grind size of P80 75um. The test work was conducted to determine a baseline gold extraction level.

Composite Identity Test No. Grind Size P80 (µm) % Au Cyanide Extraction @ hours Consumption
(kg/t)
Gravity
%
1 2 4 6 12 24 36 Lime NaCN
Cosmo
Underground
DM1669
DM1670
75 64.76 80.89 85.66 87.85 88.85 88.45 88.25 88.85 0.39 0.37

TABLE 13-12 GRAVITY/CYANIDATION LEACH TESTWORK RESULTS COSMO ORE

Overall gold extraction was relatively high with 88.85% of gold recovered. The test results indicate that some preg-robbing is occurring. Reagent consumptions were relatively low.

After reviewing the results, a CIL cyanidation leach test was requested to ascertain the degree of preg-robbing occurring during cyanidation.

Gravity/CIL Cyanidation Leach Test work

Gravity separation and subsequent CIL cyanidation leach test work was carried out on a representative sub- sample of the sample at a grind size of P80 75um.

Sample
Identity
Test No. Grind Size
P80 (µm)
% Au Extraction
@ Hours
Consumption
(kg/t)
Gravity 24 Lime NaCN
Cosmo DM2073 75 75.28 95.57 0.32 0.42
Composite DM2074

TABLE 13-13 COSMO MINE GRAVITY - CIL CYANIDATION LEACH TEST RESULTS COSMO MINE ORE

The results indicate excellent gold extraction was achieved at 95.57% . The ore contained a significant proportion of gravity recoverable gold at 75.28% . Most of the remaining gold was recovered in 24 hours of CIL cyanidation leaching. The CIL cyanidation improves the overall gold recovery from 88.25 to 95.57% . Reagent consumption levels were low.

The high percentage of gravity-recovered gold may have masked the CIL/cyanidation extraction. Typical gravity extraction at the Union Reefs mill is about 50% of the total.

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13.2.1.4   ALS Study A16107

Head Analysis

Analyte Unit Carbon Comp #1
CE74006
Carbon Comp
#1 CE74007
Carbon Comp
#2 CE74006
Carbon Comp
#2 CE74007
Au1 g/t 0.41 0.37 3.22 2.45
Au2 g/t 0.48 0.43 3.55 2.46
As ppm 2030 2040 2680 5180
Ctotal % 0.78 0.57 0.39 0.57
Corganic % 0.42 0.30 0.42 0.57
CO3 2- % 1.80 1.35 < 0.03 < 0.03
Stotal % 1.12 1.16 2.02 3.30

TABLE 13- 14 COSMO MINE HEAD ANALYSIS

For Carbon Composite #2, the arsenic levels are elevated, increasing the probability of refractory gold locked in solid solution with minerals such as arsenopyrite.

Gold grades showed a high degree of variability, increasing the probability of coarse gold grains, typically suited to gravity separation.

For Carbon composites #2, #3 and #4, the carbon is present entirely as organic/graphitic carbon, increasing the probability of preg-robbing gold from solution during cyanidation.

An ICP scan of each composite did not reveal any further elements detrimental to gold cyanidation in significant quantities.

Oxygen Uptake Rate Determination

Two-kilogram sub-samples of each Carbon Composite were ground to P80 75μm. The freshly milled slurry samples were utilized to determine the oxygen uptake characteristics of each composite at ambient temperature, and at 45ºC (to simulate Northern Australian summer temperatures). The test procedure described by G.M. Fraser was utilized.

The test work results indicate relatively low rates of oxygen uptake.

The rate of oxygen uptake generally increased at higher temperatures.

Oxygenation of the leach pulps is warranted to increase the kinetics of precious metal dissolution.

Gold Extraction Test work

Gravity/direct cyanidation (bottle roll) test work was conducted on sub-samples of each of the Carbon composites in order to determine extraction characteristics.

For each composite, a single test was conducted. The tests were conducted at a grind size of P80 75μm.

A summary of the extraction results is presented in the following table.

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Sample ID Target
P80
(µm)
% Gold Extraction
@ hours
Au grade
(g/t)
Consumption
(kg/t)
Gravity 2 4 8 48 Calc’d
Head
Leach
Residue
NaCN Lime
Carbon Comp
#1
75 45.33 88.27 89.78 91.25 89.78 0.49 0.05 0.29 0.32
Carbon Comp
#2
75 30.74 82.89 85.08 85.35 95.17 2.69 0.13 0.25 0.81
Carbon Comp
#3
75 16.58 92.63 91.23 89.85 95.22 0.52 0.03 0.22 0.31
Carbon Comp
#4
75 25.1 84.37 83.98 87.83 93.06 1.87 0.13 0.25 0.28
Carbon Comp
#5
75 47.73 86.45 87.67 91.27 95.84 0.6 0.03 0.32 0.25

TABLE 13- 15 COSMO MINE GOLD EXTRACTION RESULTS

Gravity recovery of gold was high for Carbon composites #1 and #5, ranging from 45% up to 47%. However, the mass pull to the gravity concentrate was significantly higher (3–4%) than that which would be achievable in a full-scale plant (typically 0.2 –0.5%) .

Gold extraction was excellent for all tests, with the worst performer still producing over 89% extraction.

There was no evidence of reactive sulphide minerals in the composites, as no drop in sulphur grades was observed in the leach tailing residues.

Carbon Composite #1 exhibited a small amount of preg-robbing, with gold extraction peaking after eight hours residence time then dropping away slightly over the final 40 hours.

Carbon composites #2, #4 and #5 appeared to contain minerals that inhibited gold dissolution kinetics, as leaching was still occurring after 48 hours residence time.

Cyanide and lime consumption was very low for all tests.

13.2.1.5   ALS Study A15850 – 2014

The Company’s geological personnel collected samples of various diamond cores representing potential mill feed. ALS Metallurgy carried out a program of metallurgical test work, including extraction and testing some extraction properties on a gold sample composite originating from the Cosmo Mine.

Head Assays

Analyte Unit Au1 g/t Au2 g/t As % CTotal % COrganic % Fe % STotal % SiO2 %
Composite # 1 2.98 3.01 0.754 2.07 1.44 14.2 4.38 48.6
Composite # 2 3.06 1.47 0.266 1.17 0.84 12.9 4.92 57.2
Composite # 3 3.66 4.58 0.33 1.44 1.26 10.7 5.48 58.2
Composite # 4 2.73 2.71 0.259 0.75 0.63 14.1 6.36 49.4
F10 Composite 0.08 0.11 0.023 2.67 2.40 6.77 0.90 63.6

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Analyte Unit Au1 g/t Au2 g/t As % CTotal % COrganic % Fe % STotal % SiO2 %
PM Composite 0.36 0.42 0.230 3.09 2.58 12.7 2.52 48.8
PCM Composite 0.11 0.07 0.002 7.71 7.86 9.23 5.62/6.26 51.0

TABLE 13-16 COSMO MINE HEAD ASSAYS COMPOSITES SUMMARY

Gold Extraction and Preg-robbing Test Work

Sample ID Preg
Robbing
Calc
Head
Residue % Au Extraction Consumption
(kg/t)
(%) Au (g/t) Au
(g/t)
As
(ppm)
S (%) Gravity Cyanidation Total NaCN Lime
Composite # 1 14.35 4.04 0.50 7420 3.58 29.65 56.98 87.63  0.29 0.95
Composite # 2 8.26 1.69 0.30 2650 4.36 36.25 46.02 82.27  0.24 1.11
Composite # 3 8.7 3.76 0.44 3150 4.82 24.10 64.21 88.31  0.29 0.82
Composite # 4 7.83 2.72 0.26 1920 4.32 25.56 64.88 90.44  0.36 1.00
F10 46.96                  
PM 72.00                  
PMC 53.48                  

TABLE 13-17 GOLD EXTRACTION AND PREG ROBBING RESULTS

The results indicate that almost all of the tested samples exhibited gold preg-robbing characteristics, ranging from 1.52% up to 72%. This finding was supported by the head analysis data, which showed organic/graphitic carbon in all of the samples.

Additional preg-robbing characterization test work was conducted on sub-samples of each of the carbonaceous composites (F10, PM, PMC) in order to investigate the validity of using kerosene to inhibit preg-robbing occurring during cyanidation leaching

The results indicate that the use of kerosene was able to inhibit the negative impact of the preg-robbing minerals in the ore. However, preg-robbing of the gold solution was still significant, ranging from 15–36% at the highest kerosene dosage (1,000g/t).

Gravity recovery of gold was moderate for each of the composites, ranging from 24% up to 36%. However, the mass pull to the gravity concentrate was significantly higher (3–4%) than that which would be achievable in a full-scale plant (typically 0.2 –0.5%) . Gold extraction was relatively high for all tests, ranging from 82 to 90% extraction. Further assays and mineralogical examination would be warranted to determine the nature of the minerals inhibiting overall extraction. There was some evidence of reactive sulfide minerals in the composites, as a drop in sulfur grades was observed in the leach tailing residues. Composite #1 exhibited a small amount of preg-robbing, with gold extraction peaking after 24 hours residence time and then dropping away slightly over the final 24 hours

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Oxygen Uptake Rate Determination


Time
(hours)
Oxygen Uptake Rate* (mg/L/min)
Composite #1 Composite #2 Composite #3 Composite #4
Ambient 450C Ambient 450C Ambient 450C Ambient 450C
0** -0.0380 -0.0016 0.0996 0.0291 -0.0006 0.1065 0.0288 -0.0003
1 -0.1449 -0.2930 -0.0982 -0.1677 -0.1197 -0.0770 -0.1174 -0.0960
2 -0.1080 -0.1433 -0.1045 -0.1550 -0.1102 -0.0599 -0.0957 -0.0901
3 -0.0669 -0.1458 -0.1023 -0.0952 -0.1013 -0.0998 -0.0733 -0.0906
4 -0.0836 -0.1968 -0.0625 -0.0898 -0.0937 -0.0435 -0.0732 -0.1165
5 -0.0513 -0.1919 -0.0591 -0.0582 -0.0705 -0.0603 -0.0641 -0.0932
6 -0.0464 -0.1029 -0.0582 -0.0681 -0.0821 -0.0703 -0.0355 -0.0840
24 -0.0518 -0.0741 -0.0467 -0.0461 -0.0682 -0.0358 -0.0647 -0.0675

TABLE 13- 18 COSMO MINE OXYGEN UPTAKE RESULTS SUMMARY

13.2.1.6   ALS Study A16598 – 2015

A total of five samples were submitted to ALS Metallurgy in Perth for a variety of tests. The samples were under the control of Mr. Earl Henriques, the Company’s metallurgist. A summary of test results follows:

Bond Abrasion Index (ai) Determination

Composite ID Bond Abrasion Index (Ai)
MET 650- 110-F10 4-05-15 0.2869
MET 650-110-f5 5-05-15 0.1856
MET 650-120-F17 24-05-15 0.1488
MET 650-120-F5 4-05-15 0.1329
MET 650-120-F10 12-05-15 0.1495

TABLE 13-19 COSMO MINE BOND ABRASION (AI) DETERMINATIONS SUMMARY

The Bond Abrasion Test determines the Abrasion Index, which is used to determine steel media and liner wear in crushers, rod mills, and ball mills. Bond developed this test, which is based on the wear rate in pounds of metal wear/kWh of energy used in the comminution process.

Bond Abrasion results are similar to those determined in the past for Cosmo type mineralization.

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Head Assays



Analyte


Unit
Composite ID (MET 650 – xxx)
110-F5
5-05-15
110-F10
4-05-15
120-F5
4-05-15
120-F10
12-05-15
120-F17
24-05-15
Au g/t 3.62 8.6 5.42 3.84 10.8
Au (dup) g/t 3.28 10.1 5.57 3.54 8.35
As ppm 2340 5900 6300 3660 840
CTotal % 1.05 1.05 0.84 1.26 1.29
COrganic % 0.72 0.87 0.75 0.72 0.72
Fe % 14.2 16.2 15.2 12.9 14.0
STotal % 1.74 3.88 3.84 3.92 4.96
SiO2 % 59.8 56.2 56.6 60.2 59.4

TABLE 13- 20 COSMO MINE 1 HEAD ASSAYS: SUMMARY

Variations in the duplicate gold assays indicate the samples are likely to contain some coarse gold. This is supported by the significant gravity gold recoveries achieved during gold extraction test work. Past test work also concluded that coarse gold was present.

It is known the samples contain arsenopyrite and thus the elevated arsenic and sulphide grades. Arsenopyrite and other sulphide minerals, such as pyrite, may contain refractory gold. Despite this, high overall gold extraction, ranging from 91.1% to 95.1% was achieved.

All samples contain high organic carbon levels. Organic carbon may contribute to preg-robbing during cyanide leaching. Despite this, there was no evidence of preg-robbing occurring during cyanide leaching for any of the samples.

Grind Establishment Test work

Composite Sample Mill Requisite Grind Time on 1 Kg
(min’sec”)
MET 650-110-F10 4- 05- 15 2D 21’27”
MET 650-110-f5 5-05-15 1B 16’30”
MET 650-120-F17 24-05-15 5A 19’02”
MET 650-120-F5 4-05-15 3C 17’44”
MET 650-120-F10 12-05-15 4A 17’38”

TABLE 13- 21 COSMO MINE GRIND ESTABLISHMENT TEST WORK SUMMARY

Representative sub-samples of each sample were submitted for grind establishment test work. The objective was to determine the grind time required to achieve a target grind size (P80 75μm) using a laboratory rod mill.

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Gravity/Direct Cyanidation Test work

Sample Test ID
(DM-)
Au Head
Grade g/t
% Gold Extraction @ hrs
Leach Tail
Gold Grade
(g/t)
Consumption
Kg/t
Assay Calc’d    Grav. 4 24 48 NACN Lime  
MET 650-110
F15
4-05-15
3588 3.62/
3.28
3.53 57.9 92.1 92.9 92.9 0.25  0.22 1.14
3593 4.63 - 91.4 96.1 94.6 0.25  0.22 1.12
MET 650- 110-
F10
4-05-15
3589 8.6/
10.1
7.07 32.8 91.1 91.1 91.1 0.63  0.29 1.01
3594 7.12 - 89.7 91.1 91.3 .062  0.41 0.91
MET 650- 120-
F5
4-05-15
3590 5.42/
5.57
5.61 28.5 91.6 92.3 93.6 .036  0.32 1.32
3595 5.41 - 91.3 94.5 93.0 0.38  0.36 1.44
MET 650- 120-
F10
4-05-15
3591 3.84/
3.54
3.02 38.2 92.8 94.7 94.7 0.29  0.29 1.11
3596 2.55 - 88.8 91.1 92.2 0.49  0.49 1.11
MET 650- 120-
F17
4-05-15
3592 10.8/
8.35
11.6 35.2 89.7 95.0 95.1 0.32  0.32 1.04
3597 8.54 - 92.1 92.5 93.9 0.41  0.41 1.04

TABLE 13-22 COSMO MINE SUMMARY OF GOLD EXTRACTION TEST WORK

All samples contain appreciable levels of gravity recoverable gold, with gold recovery by means of gravity separation and mercury amalgamation of the gravity concentrate ranging from 28.5% to 57.9% . Despite the high gravity recoverable gold, the inclusion of gravity gold recovery did not appear to have a significant impact on overall recovery or final leach residue grade.

Oxygen Uptake Rate Determination

Test Conditions Test #1 Test #2
% Solids (w/w) 40 40
pH (lime) 10.5 10.5
NaCN (%,w/w) 0.05 0.05
Temperature Ambient 450C

TABLE 13-23 COSMO MINE OXYGEN UPDATE RATE DETERMINATIONS, TEST CONDITIONS

Sub-samples of each sample were ground to P80 75μm and the freshly ground material submitted for oxygen uptake rate determination. For all samples, the baseline oxygen decay rate appears to be quite low. However, these values are somewhat misleading, as the low decay rate is actually due to low initial dissolved oxygen levels. For all samples, oxygen consumption was higher at 45ºC than at ambient temperature (<25ºC). In some instances, the reported oxygen uptake rate (in mg/L/min) for the two tests appear very similar, however, it was noted that in most cases the initial dissolved oxygen concentration (at the start of the 15 minute monitoring period) was lower for the test conducted at 45ºC.

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Time
(hours)
Oxygen Uptake Rate* (mg/L/min)
MET 650-110-F5
5-05-15
MET 650-110-F10
4-05-15
MET 650-120-F5
4-05-15
MET 650-120-F10
12-05-15
MET 650-120-F17
24-05-15
Ambient 450C Ambient 450C Ambient 450C Ambient 450C Ambient 450C
0**  -0.0618 -0.0094  -0.0295 -0.0070  -0.0026 -0.0062  -0.0892 -0.0289  -0.0407 -0.0202
1  -0.1544 -0.1427  -0.1161 -0.2826  -0.1511 -0.2096  -0.1528 -0.1782  -0.1140 -0.1961
2  -0.1201 -0.1200  -0.0908 -0.2810  -0.1280 -0.1096  -0.1301 -0.1803  -0.1233 -0.2056
3  -0.1152 -0.1131  -0.0878 -0.2588  -0.1163 -0.1538  -0.1154 -0.2092  -0.0734 -0.1286
4  -0.0964 -0.0910  -0.0859 -0.2364  -0.0531 -0.1623  -0.0996 -0.1709  -0.0686 -0.1782
5  -0.0237 -0.0914  -0.0329 -0.2634  -0.0506 -0.2323  -0.0939 -0.1617  -0.0599 -0.1581
6  -0.0797 -0.1434  -0.0739 -0.1794  -0.0455 -0.1253  -0.0539 -0.1795  -0.0524 -0.2013
24  -0.0291 -0.0373  -0.0390 -0.0864  -0.0556 -0.0851  -0.0492 -0.0910  -0.0310 -0.1017

TABLE 13-24 COSMO MINE SUMMARY OF OXYGEN UPTAKE RATE TEST WORK

* Ambient temperature
**baseline data prior to aeration

13.2.2      UNION REEFS METALLURGICAL TEST WORK

Recent metallurgical test work on the Union Reefs deposit has been limited over the past few years apart from testing done on several representative samples of oxide and sulphide mineralization from the Prospect deposit.

Historical test work is presented for the Esmeralda deposit as it has relevance to ongoing work in this area.

Date Report No. Description
2013 ALS A15107 Prospect deposit: 2 composite samples, head assays, gravity separation/cyanidation tine leach, grind times
1997 Metcon 97356 Esmeralda deposit: Abrasion index, Bond Rod Mill work index,
1996 Metcon 95218 Esmeralda deposit: Head assays, trail grinds, gravity leach, direct cyanide leach, size assay of leach residues

TABLE 13-25 SUMMARY OF REPORTS AVAILABLE FOR UNION REEFS DEPOSITS METALLURGICAL TEST WORK

13.2.2.1   ALS Study A15107 – Prospect Deposit - 2013

Two samples representing oxide mineralization (Lode 200/300) and sulphide mineralization (Lode 400) were submitted for metallurgical test work. Test results are as follows:

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Head Assays

Sample ID Au1
(g/t)
Au2
(g/t)
Ag
(g/t)
Cu
(ppm)
Fe
(%)
Lode
200/300
6.18 5.84 10 45 4.76
Lode 400 21.9 19.2 18 35 3.88

TABLE 13- 26 PROSPECT DEPOSIT HEAD ASSAYS COMPOSITES SUMMARY

Gravity Separation/ Cyanidation Time Leach Test Work

TABLE 13- 27 PROSPECT DEPOSIT SUMMARY OF GOLD EXTRACTION TEST W ORK

The gravity recoverable gold component of the 200/300 Lode was deemed to be high at 32.8% -51.9%, while the 400 Lode composite exhibited gravity recoveries of 15.2% .

Cyanide leaching was rapid with a significant percentage of gold extracted in the first eight hours. After 24 hours total extraction for the 200/300 Lode composite was 81% to 87%. For composite 400 Lode total extraction is reported as 57.1% . This is interpreted to be low and the metallurgical test results need to be reviewed.

Lime and cyanide consumption were seen to be very low.

It should be noted the calculated head for composites of 200/300 Lode were listed as 3.03g/t Au and 3.81g/t Au while head assays were 6.18 and 5.84g/t Au. The discrepancy between head grades and calculated grades is not explained. However, within internal reports relating to historic mining Makar noted that from mining activities at Prospect “Gold recoveries were in excess of 93% with nearly 50% recovered by gravity means recorded during milling of trial parcels of Prospect Claim ore (59% gravity in the first trial parcel, 38% in the second trial parcel which was of much lower grade)” (B. Makar 2005b). More work is recommended to determine the actual recovery of these higher grade lodes.

13.2.2.2  METCON Study 95218 – Esmeralda Prospect -1996

In 1996 Acacia Resources submitted 59 RC samples for metallurgical testing. These were composited into seven samples, three weathered, one transition and three fresh. The samples were described as coming from a series of steeply dipping mineralized lenses hosted by quartz/chert breccia, which has associated with an argillite. Four composites are from Zone A and three composites come from Zone B.

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Head Assays

Composite Type ~ Au g/t Ag g/t As ppm Fe % S% total
ESM-1 Zone A weathered 3.7 1.4 344 6.86 0.15
ESM-3 Zone A weathered 1.89 <0.5 234 2.9 <0.01
ESM-5 Zone B weathered 2.66 <0.5 1070 7.87 0.05
EsM-6 Zone B transition 2.16 <0.5 2110 6.88 2.4
ESM-2 Zone A fresh 2.28 0.7 154 6.50 4.6
ESM-4 Zone A fresh 6.71 0.8 30 2.88 0.22
ESM-7 Zone B fresh 1.26 <0.5 444 6.59 2.9

TABLE 13-28 ESMERALDA PROSPECT HEAD ASSAYS COMPOSITES SUMMARY

Acacia requested a gravity concentration at P80250 microns followed by cyanide leaching P8075 microns.

Grinds were based on 1kg portions of –2mm crushed material. The following table summarizes grind results.

Grind ESM-1 ESM-2 ESM-3 ESM-4 ESM-5 ESM-6 ESM-7
P80=250µm 7 12 1.5 8 4 11 12
P80=75µm 16.5 24.5 6.5 17 9.5 22.5 25

TABLE 13- 29 ESMERALDA PROSPECT SUMMARY OF GRIND RETENTION TIMES IN MINUTES

Grind times are variable but reflect the geological description of degree of weathering.

Gravity/Leach and Direct Cyanidation Leach Test Work

Sample Au Head Grade
g/t
% Gold Extraction
Leach Tail Consumption
Kg/t
Assay Calc’d Grav. 48 Hr
Leach
Total % Gold Grade
(g/t)
NACN Lime
ESM-1 3.70 2.63 12.0 80.2 92.2 0.19 .38 2.86
ESM-2 2.28 2.27 19.5 75.1 94.6 .012 1.02 9.25
ESM-3 1.89 1.56 7.60 87.7 95.3 0.08 0.44 7.46
ESM-4 6.71 5.47 18.2 75.4 94.6 0.25 0.54 5.12
ESM-5 2.66 1.00 17.1 78.7 95.8 0.06 .050 4.70
ESM-6 2.16 2.09 57.8 35.1 92.9 0.18 0.49 4.32
ESM-7 1.26 1.34 34.2 59.1 93.3 0.11 0.53 4.07

TABLE 13-30 ESMERALDA PROSPECT SUMMARY OF GOLD EXTRACTION TEST WORK

The results of gravity followed by a cyanide leach with those of a direct cyanide leach are remarkably similar.

Lime consumptions are relatively high likely reflecting relatively high sulphur content in some samples. Cyanide consumption was deemed to be moderate.

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It should be noted that free gold flakes up to 250 micron were observed. Nevertheless total gold extractions exceeded 90% for all samples.

Leach residues were sized down to 38 microns and assayed for gold to highlight any grind liberation effects. It was determined that an optimum grind of about 53 microns is indicated.

13.2.2.3   METCON Study 97356 – Esmeralda Prospect -1997

In 1997 Acacia submitted four samples of drill core to determine Abrasion index, Bond Rod Mill work index and Bond Ball Mill work index at 75 micron.

One weathered and one fresh sample were submitted from each of Zone A and Zone B of the Esmeralda deposit.

TABLE 13-31 ESMERALDA PROSPECT - SUMMARY OF ABRASION, ROD AND BALL MILL WORK INDEX RESULTS

13.2.3      PINE CREEK PROCESSING

The ore mined at Pine Creek between 1985 and 1994 was processed at the now defunct Pine Creek Mill. The mill was located on site at the southern end of the Enterprise deposit Pit and has since been removed. The processing plant employed carbon in pulp (CIP) technology to extract gold from both primary and oxidized ore. Small amounts of gold were also extracted using a heap leach method. This yielded just over 20kg of gold bullion between 1991 and 1993.

Gold recovery in Pine Creek was found to be higher from oxidized ore than it was from primary ore; this is due to the somewhat refractory nature of the gold. Over all recovery was found to be 79%, though mining reports showed that this fluctuated from year to year depending on the ratio of primary to oxidized ore the mill was processing. It was found that oxide ore recovery could be increased from 75% to 85% when the ore was finely ground so that ore that 80% of the particles were finer than 75 microns. The same increase was not seen when primary mineralization was finely ground.

 
Tonnes
Grade
g/t Au
Gold
(oz)
Recovery
%

Year
Pits Mined
1985/86 461,655 2.37 29,266 82.2 Enterprise
1986/87 1,010,076 2.27 61,697 85.4 Enterprise/Czarina
1987/88 1,322,339 2.31 80,030 82.5 Enterprise/Czarina
1988/89 1,434,717 2.48 83,933 72.9 Enterprise/Czarina
1989/90 1,133,976 2.66 74,075 77.2 Enterprise/Czarina
1990/91 890,021 3.48 77,667 77.9 Enterprise/Czarina/Monarch/Millwood

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Year

Tonnes
Grade
g/t Au
Gold
(oz)
Recovery
%

Pits Mined
1991/92 1,201,236 3.62 106,962 76.6 Enterprise/Czarina/Millwood
1992/93 1,435,318 2.24 80,313 78.0 Enterprise/Czarina/Millwood
1993/94 1,588,871 1.74 77,783 84.9 International/Gandy's
1994/95 1,170,000 1.90 57,923 78.1 International/Gandy's
1995/96 430,842 1.59 17,641 72.2 International/Gandy's
Total 10,016,554 2.53 642,459 79.4  

TABLE 13-32 PINE CREEK MILL PRODUCTION FIGURES FROM OPEN FILE REPORTS

Lime was used in the Pine Creek mill to produce the alkaline conditions required. It was found that more lime was required for oxidized ore than was required for primary ore. Both types consumed the same quantity of cyanide during processing.

Experimental work concluded that gravity separation of free gold could recover between 2.5% and 40% of the total gold, this averaged to about 15%.

In the years 1991-1993 the final product from the Pink Creek mill contained 16% silver.

Between 2011 and 2012 Crocodile Gold carried out a limited amount of metallurgical test work on mineralization from the International deposit at a number of different laboratories listed below.

13.2.4      INTERNATIONAL DEPOSIT

Date Report No. Description
February 2011 Ammtec A 13327 Grind recovery test work on red, blue & green composites
January 2011 NAL Bottle rolls on red, blue & green composites (as per Ammtec)
December 2012 Stawell Gold Mine Recovery Test work on prepared composites

TABLE 13- 33 METALURGICAL TEST-WORK FOR INTERNATIONAL DEPOSIT

13.2.4.14   AMMTEC Test-Work A13327

Crocodile Gold contracted Amtec for a defined program of metallurgical test work that was carried out on Red, Green and Blue composites representing different types of mineralization from the International deposit. Salient test data are summarized below:

Head Assays

Duplicate gold fire assay was performed on each composite

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Bond Ball Mill Work Index Determination

A sub-sample of the three composites was combined to produce an Overall Composite, which was then submitted for Bond Ball Mill Work Index test work

Gravity Separation/Cyanidation Time Leach Test Work

Gravity separation/direct cyanidation time leach tests were conducted on the three composite samples - Red, Green and Blue.

Comments

  (i)

The portion of gravity recoverable gold ranged from 63.86 to 36.82%.

  (ii)

Red Composite gave the best overall gold recovery at 95.35% and Blue, the lowest at 76.64%.

  (iii)

Sodium cyanide and lime consumption were relatively low for all samples, being <0.50 and <1.00kg/t, respectively.

13.2.4.2   NAL Test Work -2011

PF101122 231110 Au Au(R) Au Cu Pb Zn Ni Co Ag
Data Store Units ppm ppm ppm ppm ppm ppm ppm ppm ppm
LLD's in Store Units 0.01 0.01 0.01 1 5 2 2 2 1
Inter Met Green 1.03 1.04 - 69 568 1233 4 11 L
Inter Met Blue 0.78 0.8 - 288 2211 1203 36 84 1
Inter Met Red 1.23 1.33 - 172 509 572 9 26 L
Inter Met Green L/Residue 0.18 0.18 - - - - - - -
Inter Met Blue L/ Residue 0.24 0.22 - - - - - - -
Inter Met Red L/ Residue 0.09 0.08 - - - - - - -
Inter Met Green Soln 2 Hr - - 0.47 - - - - - -
Inter Met Green Soln 4 Hr - - 0.48 - - - - - -
Inter Met Green Soln 8 Hr - - 0.48 - - - - - -
Inter Met Green Soln 24 Hr - - 0.48 - - - - - -
Inter Met Blue Soln 2 Hr - - 0.32 - - - - - -

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PF101122 231110 Au Au(R) Au Cu Pb Zn Ni Co Ag
Data Store Units ppm ppm ppm ppm ppm ppm ppm ppm ppm
Inter Met Blue Soln 4 Hr - - 0.34 - - - - - -
Inter Met Blue Soln 8 Hr - - 0.33 - - - - - -
Inter Met Blue Soln 24 Hr - - 0.33 - - - - - -
Inter Met Red Soln 2 Hr - - 0.58 - - - - - -
Inter Met Red Soln 4 Hr - - 0.6 - - - - - -
Inter Met Red Soln 8 Hr - - 0.6 - - - - - -
Inter Met Red Soln 24 Hr - - 0.61 - - - - - -

TABLE 13-37 INTERNATIONAL DEPOSIT HEAD ASSAY RESULTS OF NAL TESWORK

13.2.4.3   Stawell Mine Test Work - 2012

Once assaying and logging of the Crocodile Gold diamond drill core was complete a series of metallurgical samples were composited using different lithological domains and grade ranges. Each interval was supplied with surrounding dilution material for the test work. This allowed the Company to assess the effect of dilution on the mineralized zones in terms of overall recovery.

Sample Number Description Grade Estimate Section Lithology Lode   Grade
Sample 1 Lode 500 Siltstone Section 12540 Estimated grade ~2g/t      12540 Silt 500 HG
Sample 2 Lode 100 Siltsone/Greywacke interbedded Section 12540 Estimated grade 0.79g/t      12540 Silt/Grey 100 LG
Sample 3 Lode 100 Greywacke ore Section 12540 Estimated grade +5.0g/t      12540 Grey 100 HG
Sample 4 Lode 200 Siltstone Ore Section 12540 Estimated grade 1.61g/t      12540 Silt 200 HG
Sample 5 Lode 200 Siltsone/Greywacke interbedded Section 12540 Estimated grade 0.85g/t      12540 Silt/Grey 200 LG
Sample 6 Lode 100 Greywacke Ore Section 12540 Estimated grade 1.58g/t      12540 Grey 100 HG
Sanple 7 Lode 100 quartz vein in greywacke ore Section 12540 Estimated grade 1.22g/t      12540 Grey 100 MG
Sample 8 Lode 200 wide zone Siltstone ore with internal dilution Section 12540 Estimated grade 1.90g/t      12540 Silt 200 HG
Sample 9 Lode 500 Siltsone Ore Section 13000 Estimated grade 1.13g/t      13000 Silt 500 MG
Sample 10 Lode 100 Greywacke and Qtz vein ore on Section 13000 Estmated grade 1.83g/t      13000 Grey 100 HG
Sample 11 Lode 100 interbedded Greywacke and Siltsone ore on Section 13000 Estimated grade 0.83g/t      13000 Silt/Grey 100 LG
Sample 12 Lode 200 low grade mineralisation in siltstone section 13000 Estimated grade 0.55g/t      13000 Silt 200 SG
Sample 13 Lode 300 mineralisation interbedded Greywack and Siltstone on section 13000 Estimated grade 0.77g/t      13000 Silt/Grey 300 LG
Sample 14 Lode 500 mineralisation interbedded Greywacke and Siltstone low grade Estimated grade 0.60g/t      13000 Silt/Grey 500 SG
Sample 15 Lode 100 interbedded Greywacke and Siltstone with Qtz veins, section 13000. Contains internal waste Estimated undiluted grade 1.30g/t 13000 Silt/Grey 100 MG
Sample 16 Lode 100 contact mineralisation section 13000 Estimated grade ~2g/t      13000 Silt/Grey 100 HG
Sample 17 Lode 200 high grade Greywacke Ore section 13000 Estimated Grade +5g/t      13000 Grey 200 HG
Sample 18 Lode 300 Siltstone Ore section 13000 Estimated grade 1.02g/t      13000 Silt 300 MG
Sample 19 Lode 300 Greywacke ore section 13000 Estimated grade 2.12g/t      13000 Grey 300 HG

TABLE 13- 38 METALLURGICAL SAMPLES FROM INTERNATIONAL DEPOSIT

The material was gathered and shipped to the Stawell Gold Mine, which has a simple metallurgical lab based on site. This lab allows for bottle roll and consumable consumption test-work to be completed. This work was planned to work in conjunction with the previous work completed by past companies. The results confirmed the overall gold recovery of 88%.

The results of this work can be seen below;

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TEST DETAILS    
MET SAMPLE
CONDITION
HEAD
RESIDUE
9 HR
RESIDUE
15 HR
RESIDUE
24 HR
REC %
9 HR
REC %
15 HR
REC %
24 HR
NT 15Nov2012-01 ORE 4.53    0.97 0.90 0.90 78.59% 80.13% 80.13%
  ORE & DILUTION 0.55    0.11 0.11 0.11 80.00% 80.00% 80.00%
NT 15Nov2012-02 ORE 1.02    0.11 0.06 0.06 89.22% 94.12% 94.12%
  ORE & DILUTION 0.71    0.04 0.04 0.04 94.37% 94.37% 94.37%
NT 15Nov2012-03 ORE 4.25    0.78 0.77 0.68 81.65% 81.88% 84.00%
  ORE & DILUTION 2.00    0.23 0.18 0.26 88.50% 91.00% 87.00%
NT 15Nov2012-04 ORE 1.64    0.16 0.17 0.29 90.24% 89.63% 82.32%
  ORE & DILUTION 0.82    0.07 0.11 0.12 91.46% 86.59% 85.37%
NT 15Nov2012-05 ORE 1.81    0.13 0.12 0.14 92.82% 93.37% 92.27%
  ORE & DILUTION 1.00    0.10 0.07 0.09 90.00% 93.00% 91.00%
NT 15Nov2012-06 ORE 2.08    0.26 0.22 0.24 87.50% 89.42% 88.46%
  ORE & DILUTION 1.45    0.17 0.18 0.20 88.28% 87.59% 86.21%
NT 15Nov2012-07 ORE 1.11    0.26 0.19 0.08 76.58% 82.88% 92.79%
  ORE & DILUTION 0.84    0.20 0.19 0.13 76.19% 77.38% 84.52%
NT 15Nov2012-08 ORE 2.05    0.22 0.26 0.21 89.27% 87.32% 89.76%
  ORE & DILUTION 1.36    0.19 0.18 0.22 86.03% 86.76% 83.82%
NT 15Nov2012-09 ORE 1.57    0.19 0.11 0.16 87.90% 92.99% 89.81%
  ORE & DILUTION 0.87    0.06 0.04 0.05 93.10% 95.40% 94.25%
NT 15Nov2012-10 ORE 1.14    0.18 0.14 0.14 84.21% 87.72% 87.72%
  ORE & DILUTION 0.52    0.08 0.05 0.07 84.62% 90.38% 86.54%
NT 15Nov2012-11 ORE 0.81    0.15 0.15 0.15 81.48% 81.48% 81.48%
  ORE & DILUTION 0.47    0.10 0.11 0.12 78.72% 76.60% 74.47%
NT 15Nov2012-12 ORE 1.01    0.17 0.16 0.19 83.17% 84.16% 81.19%
  ORE & DILUTION 0.36    0.11 0.11 0.14 69.44% 69.44% 61.11%
NT 15Nov2012-13 ORE 0.91    0.14 0.16 0.15 84.62% 82.42% 83.52%
  ORE & DILUTION 0.67    0.10 0.09 0.11 85.07% 86.57% 83.58%
NT 15Nov2012-14 ORE 1.15    0.15 0.10 0.11 86.96% 91.30% 90.43%
  ORE & DILUTION 0.42    0.07 0.06 0.05 83.33% 85.71% 88.10%
NT 15Nov2012-15 ORE 1.33    0.20 0.19 0.21 84.96% 85.71% 84.21%
  ORE & DILUTION 1.36    0.20 0.18 0.21 85.29% 86.76% 84.56%
NT 15Nov2012-16 ORE 0.97    0.17 0.15 0.16 82.47% 84.54% 83.51%
  ORE & DILUTION 0.89    0.14 0.12 0.13 84.27% 86.52% 85.39%
NT 15Nov2012-17 ORE 11.59    0.76 0.71 0.83 93.44% 93.87% 92.84%
  ORE & DILUTION 3.31    0.36 0.31 0.34 89.12% 90.63% 89.73%
NT 15Nov2012-18 ORE 1.28    0.03 0.06 0.06 97.66% 95.31% 95.31%
  ORE & DILUTION 0.69    0.02 0.01 0.02 97.10% 98.55% 97.10%
NT 15Nov2012-19 ORE 1.52    0.18 0.16 0.15 88.16% 89.47% 90.13%
  ORE & DILUTION 1.34    0.08 0.13 0.14 94.03% 90.30% 89.55%
                 
TOTAL ORE 2.20    0.27 0.25 0.26 87.53% 88.56% 88.25%
  ORE & DILUTION 1.03    0.13 0.12 0.13 87.62% 88.44% 87.01%

TABLE 13-39 RESULTS OF THE BOTTLE ROLL TEST-WORK ON INTERNATIONAL DEPOSIT MINERALIZATION

MET SAMPLE
CONDITION
HEAD
g/t
RESIDUE
9 HR
RESIDUE
15 HR
RESIDUE
24 HR
REC %
9 HR
REC %
15 HR
REC %
24 HR
CN
kg/t
LIME
kg/t
TOTAL ORE 2.20 0.27 0.25 0.26 87.5% 88.6% 88.2% 0.75 1.5
  ORE & DILUTION 1.03 0.13 0.12 0.13 87.6% 88.4% 87.0% 0.71 1.3

TABLE 13-40 CONSUMABLE REQUIREMENTS FOR INTERNATIONAL DEPOSIT MINERALIZATION WITH RESIDUE TIMINGS

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13.2.5      BURNSIDE METALLURGICAL TEST WORK

13.2.5.1  AMMTEC A13929 – Rising Tide Deposit

Date Report No. Description
February
2012
Ammtec
A13929
Fresh dolomite, carbonaceous shale & transitional mineralization. Leach/Gravity Recovery,
Preg-robbing characterization/recovery. Mineralogy.

TABLE 13-41 TEST WORK SUMMARY COMPLETED ON RISING TIDE DEPOSIT

Three (3) gold-bearing RC composites from the Rising Tide deposit were tested for quantitative automated mineralogical analysis. The samples were also separated into a Knelson gravity concentrate and tail fractions. The Knelson concentrate was further upgraded by hand panning and the rejects were recombined with the Knelson tails.

The sample details are listed in the table below:

 Samples Received, Mass Split of Gravity Separation and Selected Assay Results  
Sample Sample
Type
Mass %
Retained
(1.0 kg
Feed)
Au Grade
Au1/Au2

(ppm)
Ag
Grade
(ppm)
Fe
Grade
(%)
Al
Grade
(%)
Ca
Grade
(%)
S
TOTAL
Grade
C
TOTAL
Grade
(%)
Mineralogy Sample No.
Fresh
Dolomite Ore
Comp
Knelson
Pan Con
5.05 1.41/1.43 <0.3 19.3 2.16 8.00 8.24 0.09 MIN969A1A
Knelson
Pan Tail
94.95 MIN969A2A
Fresh Carbonaceous Shale Ore Comp Knelson
Pan Con
4.16 0.85/0.73 <0.3 14.8 4.08 6.90 6.04 0.21 MIN969B1A
Knelson
Pan Tail
95.84 MIN969B2A
Transitional Ore Comp Knelson
Pan Con
4.03 1.77/1.74 0.6 14.3 3.76 0.30 7.06 0.36 MIN956C1A
Knelson
Pan Tail
95.97 MIN956C1A

TABLE 13-42 SAMPLE DETAILS FOR RISING TIDE DEPOSIT

Quantative Mineralogy

The sample also comprises minor amounts of feldspars, fluorite, iron and titanium oxides, chlorite and pyrite. Arsenopyrite and other minerals occur as trace amounts. The Knelson concentration has been largely inefficient in separating the heavier sulphide phases from the common gangue silicates. Some large particles composed of dominant silicates especially amphiboles have preferentially deported to the concentrate. The upgrading of (liberated) pyrrhotite in the concentrate is about fourfold compared to the feed sample.

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The sample contains a considerable amount of iron and titanium oxides and fluorite (~3% each). Other minerals are negligible. Upgrading was relatively inefficient with largely liberated pyrite being weakly concentrated. The pyrite content in the concentrate is four times greater than in the feed sample.

Gold-Silver Mineralization

The number of detected gold grains of interest does not allow any characterization of gold mineralization from the analyzed samples. The panning Knelson concentration was largely inefficient in separating sulphides from the common gangue minerals with an exception of the Transitional Mineralization. Instrumental settings were able to pick up a fine grain of a silver telluride (hessite).

  Summary of Detected Gold-Silver Minerals 
Product No. of
Particles
Grain
Size
Dominant Mineral Phase/Assoc Dominant
Liberation Type
by Mass
Qualitative
sizing
Fresh
Carbonaceous
Shale Ore
Panned Knelson
Con
Con 1 1 Hessite/Pyrite Encapsulated Very Fine and Grainy
Transitional Ore
Panned Knelson
Con
Con 1 1 Electrum/Arsenop yrite Lollingite Encapsulated Fine and Grainy

TABLE 13-43 RISING TIDE DEPOSIT - SUMMARY OF DETECTED GOLD -SILVER MINERAL

Sulphides may host some proportion of very fine gold-silver mineral grains. The liberation reports of the major sulphides demonstrate that liberated sulphides account for the great majority of the sulphide mass. Therefore sulphides are amenable for further upgrading with a possible gold-rich concentrate for additional processing.

Additional work included the following:

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A defined program of metallurgical test work was carried out on composites created from RC chip samples representing the Rising Tide deposit.

The following composites were created for the test program:

Head Assays

Analyte Unit Fresh Dolomite Ore Fresh Carbonaceous
Shale Ore
Transitional Ore
Au1 g/t 1.41 0.85 1.26
Au2 g/t 1.43 0.73 1.77
Ag g/t <0.3 <0.3 0.6
CTOTAL % 0.09 0.21 0.36
Cu ppm 578 366 184
Hg ppm 0.3 0.3 <0.1
Pb ppm 10 <5 20
STOTAL % 8.24 6.04 7.06
Te ppm 4.4 2.4 1
Zn ppm 120 150 142

TABLE 13- 44 HEAD ASSAY FOR RISING TIDE SAMPLES

Gravity Separation/Cyanidation Test Work

Each composite was submitted for gravity separation with subsequent cyanidation test work at three grind sizes.


Composite
Identity

Test
No

Grind Size
(µm)
% Au Extraction @
Hours

Residue
(ppm)

Consumption (kg/t)
       
Gravity 36 Lime NaCN
Fresh
Dolomite
Ore
MA935 150 16.66 68.45 0.42 0.48 1.17
MA936 106 15.39 77.07 0.29 0.88 1.20
MA937 75 16.20 78.94 0.27 0.60 1.26
Fresh
Carbonaceous
Shale Ore
MA938 150 19.82 61.87 0.29 1.47 1.63
MA939 106 42.43 77.02 0.22 1.51 1.72
MA940 75 15.17 75.31 0.22 1.65 1.78

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Composite
Identity

Test
No

Grind Size
(µm)
% Au Extraction @
Hours

Residue
(ppm)
Consumption (kg/t)
   
Gravity 36 Lime NaCN
Transitional Ore MA941 150 75.56 93.02 0.15 2.58 0.74
MA942 106 74.57 96.09 0.09 3.06 0.78
MA943 75 72.21 97.48 0.05 2.99 0.87

TABLE 13-45 GRAVITY AND CYANIDE EXTRACTION RESULTS FOR RISING TIDE DEPOSIT

Preg-Robbing Characterization

Each composite was submitted for preg-robbing characterization at grind size P80 75μm.

Sample Identity Preg-Robbing
ppm %
Fresh Dolomite Ore -0.2 -1.08
Fresh Carbonaceous Shale Ore 0.8 4.3
Transitional Ore 17.3 92.85

TABLE 13-46 PREG-ROBBING CHARATERISTICS FOR RISING TIDE DEPOSIT SAMPLES

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14 MINERAL RESOURCE ESTIMATIONS

14.1 INTRODUCTION

The NT Operations have previously been individually identified but frequently referred to as the Cosmo Mine, the Burnside Gold & Base Metals Project, the Union Reefs Gold Project and the Pine Creek Gold Project. Within each of these project areas are located numerous gold deposits with estimated mineral resources and mineral reserves. The processing facility at Union Reefs is factored into the economic evaluation of all of the Company’s mineral resources and mineral reserves in the NT Operations and as a result of the shared infrastructure and close proximity of the various projects Newmarket Gold has determined it is prudent to prepare one report and treat the NT Operations as one project.

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14.2 COSMO MINE MINERAL RESOURCE

14.2.1      INTRODUCTION

Cosmo Mine Mineralized Domains (Au >= 2 g/t) 
Domain Tonnes Gold Grade g/t Oz Gold
Measured 1,650,000 3.63 192,500
Indicated 2,987,000 2.99 287,600
Total (Measured and Indicated only) 4,637,000 3.22 480,100
Inferred 678,000 2.76 60,200

TABLE 14-2 MINERAL RESOURCE ESTIMATIONS COSMO MINE PROJECT NORTHERN TERRITORY DEPLETED TO 31ST DECEMBER 2015

Notes on Table 14-2:

1.

Mineral resources are stated as of December 31, 2015.

   
2.

Mineral resources are inclusive of mineral reserves, which are set out below.

   
3.

Mineral resources are calculated using these parameters.

   
4.

Gold Price of $A1,500/oz, metallurgical recovery of 92.0%.

   
5.

Lower cut-off of 2.0g/t Au is used to calculate the mineral resources.

   
6.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces.

   
7.

The mineral resource estimate was prepared by Mark Edwards, B.SC. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold.

   
8.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.


14.2.2      COSMO MINE RECONCILIATION

The Cosmo Mine has been extracting ore since late 2011.

Table 14-3 below relates to the mining period from January 2015 to December 2015. The grade control model data in the table is taken from the series of grade control block models that have been generated for mine scheduling and planning at certain points in time. These block models change as new data is added and geological understanding increases. During 2015 significant changes have been made to the geological interpretation of the Cosmo mineralization. Changes and improvements have also been made to the kriging parameters used in the block model estimation.

Overall for the mining period, the Cosmo Underground Mine produced 5% more tonnes whilst the grade of this material was down 11% versus the block model designed tonnes and grade. This increased tonnage and reduced grade is mainly due to the dilution of the stopes with the failure of the F10 Fault. Minor development tonnage increases are being seen through stripping, higher backs and fillets taken on the turn-outs; these tonnes are not incorporated in the planned tonnes. The increased tonnage seen in the development is not classified as dilution, as the development is predominantly through the mineralization lodes.

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Month Grade Control Model
Grade
Reconciled Mined % Diff Reconciled Mined
to Grade Control Model
  Tonnes Grade Ounces Tonnes Grade Ounces Tonnes   Grade Ounces
Jan 65,552 4.41 9,297 71,333 3.81 8,740 9% -14% -6%
Feb 63,706 4.38 8,980 62,392 3.74 7,493 -2% -15% -17%
Mar 62,607 3.14 6,326 63,864 3.36 6,890 2% 7% 9%
Apr 49,685 3.50 5,587 53,552 3.57 6,145 8% 2% 10%
May 56,674 3.12 5,679 61,698 2.70 5,359 9% -13% -6%
Jun 64,417 2.98 6,171 68,056 2.76 6,032 6% -7% -2%
Jul 52,203 2.87 4,811 58,173 2.62 4,897 11% -9% 2%
Aug 46,709 3.23 4,850 50,560 2.78 4,519 8% -14% -7%
Sep 53,543 3.17 5,449 55,755 2.59 4,649 4% -18% -15%
Oct 50,413 2.99 4,841 52,095 2.74 4,586 3% -8% -5%
Nov 64,571 3.17 6,575 67,023 2.76 5,947 4% -13% -10%
Dec 61,793 3.02 6,000 62,136 2.37 4,734 1% -22% -21%
YTD 691,873 3.35 74,567 726,637 3.00 69,991 5% -11% -6%

TABLE 14-3 RECONCILIATION RESULTS FOR COSMO MINE JANUARY - DECEMBER 2015, AU G/T

14.2.3      GEOLOGICAL INTERPRETATION

Lithological/Structural interpretation.

Wireframes are created of the major lithological contacts. There are four such contacts:

Lithological contacts are digitized from points on diamond drill holes and surveyed contacts in underground (mostly) and open pit workings.

Similarly generated major faults include:

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Gold Lode Interpretation

The gold lodes in the footwall are remarkably planar. The hangingwall lodes are more complex due to parasitic folding of which many are isoclinal. Each lode is correlated by grade within its stratigraphic position in the mineralization bearing siltstones. To date seven lodes have been identified:

100 Lode – The best gold-endowed lode is constrained within the contact of the Graphitic Mudstone (Pmc unit) and the F10 Fault. In the hangingwall lodes, the F10 Fault deviates away from the lodes and mineralization appears to be related to parasitic folding against the Graphitic Mudstone. The 100 Lode contains, near its center, a thin internal Graphitic Mudstone unit (termed the 11 unit), which is often un-mineralized. Gold grades are easily correlated in plan and section.

200 Lode – The first mineralization that occurs west of the F10 Fault. Gold grades are usually more erratic and lower grade than in the 100 Lode, but is still clearly correlated. The economic portion of 200 Lode terminates to the north where the F10 Fault deviates northeast, becoming cross cutting. The 200 Lode essentially becomes the 600 Lode as it strikes towards and crosses the Cosmo Anticline fold hinge.

300 Lode – This is the next lode to the west of the 200 Lode being separated generally by 5-6m of less -altered and –sulfidic, barren siltstone or mudstone. This lode is of lower average gold grades with variable and indistinct grade contacts. At depth and in the southern extent of this lode the gold grades may improve.

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400 Lode – This is the innermost lode. It is located close to the Zamu Dolerite and is of consistently low gold grade with similar indistinct grade boundaries as the 300 Lode. This lode splays off the 300-Lode.

500 Lode – This is the continuation of the 100 Lode as it wraps around the Cosmo Anticline fold hinge and becomes part of the Western Lodes. Like the 100 Lode it occurs nearest to the Pmc Graphitic Mudstone. There is no F10-Fault in the Western Lodes

600 Lode – Similar to the 500 Lode this is the continuation of the 200 Lode as it wraps around the Cosmo Anticline fold hinge to become part of the Western Lodes.

101 Lode – Termed the Sliver Zone, this lode is a subsidiary fold that veers to the north. The lower extents and internal faulting of this mineralization are unclear and are the subject of ongoing mineral resource definition drilling and spatial mineralization studies.

Methodology

In a given stratigraphic position, all contiguous mineralization greater than 0.2g/t Au is coded as the relevant lode using wireframes created from geological data collected such as diamond drilling, face mapping, underground backs/wall mapping and survey picked up contacts. Some lodes, such as the 100-Lode, are strata-bound; in this case by the graphitic mudstone lithological contact and the F10 Fault. Other lodes such as the 300 Lode have diffuse, grade dependent, boundaries with lithological units including gold mineralization above 0.2g/t Au.

Lode wireframes are snapped to relevant contacts on drill holes wherever possible. Face map chip sample line information is also incorporated into lode and lithology modeled wireframes where relevant. Sludge sample gold assay results can also be used in wireframing modeling decisions to define the edges of the mineralization to within 0.9m but sample grades are not used for mineral resource grade estimation.

Upon completion of each wireframe contact, the wireframe is spatially closed and validated. Each wireframe must not overlap with any other of the lode wireframes or errors will occur during block modeling. Each closed wireframe is a separate file for estimation purposes.

Waste zones that occur between the lodes are wireframed from the lode contacts thereby preventing overlaps. These wireframes are closed, validated and once given waste codes saved as separate files.

14.2.4      BULK DENSITY

There have been several campaigns of bulk density test work on the Cosmo mineralization. The first was during the exploration phase. These densities have been used in previous reports. A density of 2.93t/m 3 was used in reports of model tonnages and grades for the 100 Lode, with 2.88t/m 3 used to report the 200 and 300 Lodes. These densities were derived from a total of 103 wax encapsulated mineralized samples taken from ½ HQ drill core. A total of 170 samples were also selected from various lithologies and lodes for air pycnometer testing to validate the wax encapsulated density values, with results indicating a close correlation. There are over 1,000 samples with associated bulk density analysis for the Cosmo deposit, with the majority of these located in the sulphide mineralization zone.

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Since the underground has become operational, onsite bulk density (BD) data is collected, to confirm the SG data values being used are correct.

The diamond drill core samples were selected as representative of each lithology present at Cosmo. Samples were chosen by geologists (core pieces around 20cm in length), approximately one every second core tray, and the intervals and lithology recorded in the log.

In situ bulk density (BD) determinations were estimated using the water displacement method on drill core. BD data derived from Fresh and Transitional material had been considered reliable, as rock competency and core recovery improved with depth, reducing the variability of results. In oxide zones, the same technique of water displacement was used, on handpicked sticks of competent core. This may have imposed a bias towards overestimation. Bulk density measurements estimated using the water displacement method was calculated from the following formula:

BD = WAD

WAS – WWS

Where WAD = Weight of dry sample in air
  WAS = Weight of saturated sample in air
  WWS = Weight of saturated sample immersed in water

Method

Determining WAD: For each meter of core requiring measurement, a 0.2m piece of core was removed from the tray (prior to core being cut). Core affected by grease or drilling fluid was not chosen. Each piece of core was placed on the scales and its weight recorded.

Determining WAS: The sample was re-weighed in air allowing minimal drying.

Determining WWS: A container of water with sufficient volume to immerse a 0.2m length of core was placed on the scales. The pieces of core (in a wire basket) were lowered into the water and when the weight steadies (after the sample has become saturated) the weight was recorded.

All measurements were recorded and BD’s calculated using the above formula.

The results of all the operational onsite bulk density data are separated into geological domains in Surpac Mine software.Table 14-4 summarizes the results.

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Domain bulk density g/cm3
Dolerite pdz 2.93
All meta seds 2.88
100, 110, 120, 130, 150, 101, 500, 550 Lode 2.93
200, 210, 220, 230, 250, 300, 310, 320, 330, 350, 400, 420, 430, 450,
600, 650 lode

2.88
Pca (dolomite) 2.88
Pmc 2.88
Unclassified metaseds 2.88

TABLE 14-4 BULK DENSITY FOR LODES AT COSMO MINE

14.2.5      DATA TYPES

The estimation of contained gold has been based on assays sourced from drilling and face sample data, detailed in Section 10, above. The data available as at December 2015 consisted of diamond core samples derived from historic exploration and mining definition campaigns as well as face chip samples derived during mine grade control. Sludge drilling results were included in the database but excluded from compositing and subsequent estimation. Also excluded were 28 drill holes and one face sample (875_ACC_LW) that had failed the data validation process.

All data is provided in local grid co-ordinates.

Drilling provides data to depths up to 1,000m below surface. The total database consisted of 1,305 (for 221,389m) diamond drill holes and 4,653 face sample lines (for 29,099m). The drill core and face sampling were sampled and assayed mostly at 1m intervals, although the database contains intervals at varying lengths within mineralized lodes as summarized in Table 14-5. The higher sample lengths in most holes are due to core loss in the sample interval.

Mineralised
Domain
Minimum Length
(m)
Maximum Length
(m)
Mean Length
(m)
100 0.1 5.0 0.976
110 0.2 1.75 0.961
120 0.1 2.4 0.966
130 0.2 2.2 0.951
200 0.15 2.3 0.925
210 0.1 1.9 0.914
220 0.1 3.0 0.970
230 0.25 3.68 0.967
300 0.2 5.0 1.001
310 0.3 2.0 0.911
320 0.2 2.7 1.032
330 0.15 2.9 0.988
400 0.1 2.2 0.936
420 0.4 2.6 0.959
430 0.2 2.0 0.948
500 0.3 1.75 0.859

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Mineralised
Domain
Minimum Length
(m)
Maximum Length
(m)
Mean Length
(m)
600 0.2 1.9 0.990
101 0.1 5.0 0.978
150 0.1 4.88 1.006
250 0.1 3.05 1.001
350 0.1 4.6 1.044
450 0.2 2.1 1.025
550 0.16 5.5 0.961
650 0.1 4.57 0.984

TABLE 14-5 SUMMARY OF SAMPLE LENGTHS BY MINERALIZED DOMAIN

14.2.6      GEOLOGICAL INTERPRETATION

The Cosmo deposit is interpreted to be a series of steep northwest plunging gold mineralized vein, breccia and shears in the Howley Anticline and hosted within inter-bedded siltstones, mudstones, banded iron, phyllites, dolerite sills and greywacke (Upper to Middle Koolpin Formation).

Mineralization is generally strongest adjacent to dilational structures within the sedimentary host package and elevated sulphides (pyrite and arsenopyrite) within greywacke units.

For the purposes of estimation the interpreted mineralized zones were domained into Hangingwall and Footwall based on their proximity to the F1 Fault structure.

Waste domains were interpreted between each of the mineralized domains to enable estimation of background gold grades into areas surrounding the main mineralized lodes.

14.2.7      MINERAL RESOURCE INTERPRETATION

Interpretation of mineralized domains has been informed by geological stratigraphic units and a relative gold cut-off grade based on continuity, with a lower limit of ~0.2g/t Au to an upper limit of +100g/t Au used as the basis for defining mineralized material.

The mineral resource domain interpretations were wireframed and numbered Table 14-6according to mineralized/waste, Hangingwall and Footwall mineralized domains as outlined in .

Domain Type Domain Number   Domain Type Domain Number
Footwall 100   Footwall 400
Footwall 110   Footwall 420
Footwall 120   Footwall 430
Footwall 130   Footwall 500
Footwall 200   Footwall 600
Footwall 210   Hanging wall 101
Footwall 220   Hanging wall 150
Footwall 230   Hanging wall 250
Footwall 300   Hanging wall 350

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Domain Type Domain Number   Domain Type Domain Number
Footwall 310   Hanging wall 450
Footwall 320   Hanging wall 550
Footwall 330   Hanging wall 650

TABLE 14-6 MINERALIZED DOMAIN NOMENCLATURE

The mineral resource wireframes were used to code the drill intercepts contained within them by flagging into a new table in the database, the “intercepts” table. This flagging allows the selection of data within domains by codes for the purposes of sample analysis and compositing.

All mineral resource interpretation wireframes have been used as hard boundaries for this estimate.

14.2.8      COMPOSITING AND STATISTICS

Compositing of the raw drilling sample data is necessary to establish a single support for the data (length) and to avoid bias when calculating statistics and undertaking any estimation of the data into three dimensional volumes. A number of items are considered when selecting an appropriate composite length; they include the original support of the raw sample data, the assumed selectivity (and therefore the block size) of the model and the imposed spatial dimensions of the mineralized domains.

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An examination of sample statistics by domain (Table 14-7) and combined (Table 14-8 ), for mineralized intercepts reveals that the majority of sampling is on 1m downhole support, although sample lengths vary from a minimum of 0.10m to a maximum of 5.50m downhole. The number of instances of samples less than 1m is 30.0% . Similarly the number of instances of samples greater than 1m is 25.3% . The higher sample lengths usually tend to focus on areas where there was lost core.

Domain All Mineralised Samples     Percentiles
Number 63,783   10 0.62
Minimum 0.10   20 0.80
Maximum 5.50   30 0.98
Mean 0.97   40 1.00
      50 100
      60 1.00
      70 1.00
      80 1.10
      90 1.25
      95 1.30
      97.5 1.40

TABLE 14-7 SATISTICAL SUMMARY, SAMPLE LENGTH ALL MINERALISZED DOMAINS (FOOTWALL AND HANGING WALL)

Within the mineralized domains the drill samples were composited to 1m downhole to provide equal support data for statistical evaluation and estimation.

The waste domains were also composited to 1m downhole to provide equal support data for statistical evaluation and estimation.

The effect of a small number of outlier composite grades or spatially isolated composites may have an undue effect on the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the composite data in 3D visualization for both mineralized and waste domains.

A number of high cuts were identified as necessary within both mineralized and waste domains. A statistical summary of the mineralized/waste domains is detailed in Table 14-8 and Table 14-9.

Domain Number of
Composites
Minimum Gold
Grade g/t Au
Maximum Gold
Grade g/t
Au
Mean Gold
Grade g/t Au
Co-efficient
of
Variation
100 8,949 0.01 72.22 3.11 1.35
110 2,674 0.01 38.00 2.20 1.71
120 2,888 0.01 127.31 4.33 1.36
130 1,595 0.01 59.90 3.31 1.13
200 4,861 0.01 35.58 1.89 1.02

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Domain Number of
Composites
Minimum
Gold
Grade g/t
Au
Maximum
Gold Grade g/t
Au
Mean Gold
Grade g/t Au
Co-efficient
of
Variation
210 2,057 0.01 28.00 0.82 1.98
220 985 0.01 44.96 2.22 1.15
230 973 0.01 62.00 1.84 1.48
300 5,705 0.00 127.00 2.46 1.40
310 1,017 0.01 24.65 1.17 1.46
320 829 0.01 32.00 2.95 1.12
330 1,686 0.01 117.00 2.17 1.63
400 1,807 0.01 37.00 1.49 1.25
420 239 0.01 50.01 1.68 2.09
430 965 0.01 49.60 1.69 1.63
500 83 0.01 6.32 1.39 1.08
600 165 0.01 36.00 1.67 1.88
5 (Waste) 113,597 0.00 990.00 0.41 9.21
10 (Waste) 3,767 0.01 15.38 0.25 2.46
11 (Waste) 2,121 0.01 8.52 0.14 3.29
12 (Waste) 861 0.01 20.10 0.52 2.57
13 (Waste) 1460 0.01 20.30 0.34 2.85
20 (Waste) 4,603 0.01 20.60 0.29 1.79
21 (Waste) 1,900 0.01 9.57 0.28 2.21
23 (Waste) 1,213 0.01 9.87 0.27 1.99
30 (Waste) 2,134 0.01 15.90 0.39 1.91
33 (Waste) 1,345 0.01 6.34 0.35 1.37
50 (Waste) 101 0.02 8.61 0.71 1.97

TABLE 14-8 STATISTICAL SUMMARY, GOLD PPM - FOOTWALL DOMAINS

Domain Number of
Composites
Minimum
Gold Grade g/t
Au
Maximum
Gold Grade g/t
Au
Mean Gold
Grade g/t Au
Co-efficient
of
Variation
101 9,968 0.01 66.90 2.17 1.55
150 5,522 0.00 62.60 3.17 1.33
250 1,097 0.01 77.20 2.50 1.78
350 1,052 0.01 52.00 2.20 1.66
450 210 0.00 13.30 1.76 1.23
550 5,545 0.00 161.20 2.63 2.45
650 2,949 0.00 65.56 2.69 1.62
15 (Waste) 1,570 0.01 13.61 0.46 1.97
25 (Waste) 1,074 0.01 7.70 0.37 1.65
35 (Waste) 464 0.00 7.81 0.37 2.20
55 (Waste) 3,489 0.00 39.79 0.51 3.83

TABLE 14-9 STATISTICAL SUMMARY, GOLD PPM – HANGINGWALL DOMAINS

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High-grade gold cuts were applied to the raw assay data within table ‘assay’ of the database ‘cosmo.mdb’ and then domain composites were generated from these high cut assay data columns. Summary statistics for cut composites are detailed in Table 14-10 and Table 14-11.

Domain Number of
Composites
Applied
High Cut g/t Au
Cut Mean
g/t Au
Cut Standard
Deviation
Cut Co-efficient of
Variation
100 8,979 25 3.05 3.82 1.25
110 2,674 25 2.18 3.64 1.67
120 2,888 25 4.17 4.64 1.11
130 1,595 25 3.26 3.33 1.02
200 4,861 20 1.88 1.83 0.97
210 2,057 20 0.80 1.39 1.74
220 985 20 2.18 2.14 0.99
230 973 20 1.77 1.85 1.04
300 5,705 20 2.39 2.60 1.09
310 1,017 20 1.15 1.48 1.30
320 829 20 2.92 3.10 1.06
330 1,686 20 2.11 2.15 1.02
400 1,807 20 1.48 1.70 1.15
420 239 20 1.54 1.89 1.22
430 965 20 1.63 1.87 1.15
500 83 20 1.39 1.51 1.08
600 165 15 1.54 1.94 1.26
5 (Waste) 113,597 5 0.32 0.70 2.19
10 (Waste) 3,767 5 0.24 0.51 2.09
11 (Waste) 2,121 5 0.13 0.37 2.83
12 (Waste) 861 5 0.46 0.85 1.86
13 (Waste) 1,460 5 0.31 0.68 2.15
20 (Waste) 4,603 5 0.29 0.42 1.47
21 (Waste) 1,900 5 0.27 0.52 1.92
23 (Waste) 1,213 5 0.26 0.40 1.57
30 (Waste) 2,134 5 0.37 0.49 1.33
33 (Waste) 1,345 5 0.35 0.46 1.33
50 (Waste) 101 5 0.60 0.90 1.50

TABLE 14-10 STATISTICAL SUMMARY FOR HIGH GRADE CUT COMPOSITES, GOLD G/T - FOOTWALL DOMAINS

Domain Number of
Composites
Applied
High Cut g/t Au
Cut Mean
g/t Au
Cut Standard
Deviation
Cut Co-efficient of
Variation
101 9,968 20 2.11 2.97 1.41
150 5,522 25 3.10 3.75 1.21
250 1,097 20 3.33 2.93 1.26
350 1,052 20 2.10 2.85 1.36
450 210 15 1.76 2.15 1.23
550 5,545 20 2.30 3.54 1.53
650 2,949 15 2.45 3.14 1.28
15 (Waste) 1,570 5 0.43 0.63 1.47
25 (Waste) 1,074 5 0.37 .55 1.51

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Domain Number of
Composites
Applied
High Cut g/t Au
Cut Mean
g/t Au
Cut Standard
Deviation
Cut Co-efficient of
Variation
35 (Waste) 464 5 0.34 0.58 1.70
55 (Waste) 3,489 5 0.38 0.78 2.04

TABLE 14-11 STATISTICAL SUMMARY FOR HIGH GRADE CUT COMPOSITES, GOLD G/T – HANGINGWALL DOMAINS

The data populations within the majority of mineralized domains are positively skewed with moderate variability. The variability is reduced somewhat by high cutting of gold grades in those domains with relatively high coefficients of variation.

Within waste domains the high cut was applied with the aim of reducing the influence of singular ‘outlier’ high grades whilst allowing any genuine anomalous areas to be represented within the estimation.

14.2.9      VARIOGRAPHY

Variography was used to characterize the spatial behavior of the composite data for establishing estimation parameters. Variogram stability and quality is dependent on the statistical properties of defined domains and the amount of data available within domains. After an initial investigation of the gold data, isotropic variogram models were defined individually for mineralized and waste domains. The omnidirectional variogram models (relative sills) are detailed in Table 14-12 and Table 14-13.

Domain Nugget Struct Sill Major (m) Major/ Semi Ma  jor/ Minor
100 0.212 St1 0.456 16 1 2.0
St2 0.332 190 1 10.6
110 0.273 St1 0.440 18 1 1.8
St2 0.288 140 1 10.0
120   No Variogram Defined – Used Domain 100 Model
130   No Variogram Defined – Used Domain 100 Model
200 0.290 St1 0.543 15 1 3.0
St2 0.167 100 1 6.7
210 0.288 St1 0.515 22 1 3.1
St2 0.197 180 1 12.9
220   No Variogram Defined – Used Domain 200 Model
230   No Variogram Defined – Used Domain 200 Model
300 0.288 St1 0.530 14 1 2.8
St2 0.183 50 1 3.3
310 0.240 St1 0.491 15 1 1.9
St2 0.258 125 1 8.3
320   No Variogram Defined – Used Domain 300 Model
330   No Variogram Defined – Used Domain 300 Model
400 0.227 St1 0.458 15 1 1.9
St2 0.314 58 1 4.1
420   No Variogram Defined – Used Domain 400 Model
430   No Variogram Defined – Used Domain 400 Model
500   No Variogram Defined – Used Domain 600 Model

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Domain Nugget Struct  Sill  Major (m) Major/ Semi Major/ Minor
600 0.245  St1 0.366 6 1 1.0
 St2 0.389 18 1 1.0
5 (Waste) 0.367  St1 0.452 12 1 1.0
 St2 0.181 150 1 1.0
10 (Waste) 0.334  St1 0.506 17 1 3.4
 St2 0.159 78 1 7.8
11 (Waste) 0.351  St1 0.530 13 1 2.6
 St2 0.119 75 1 7.5
12 (Waste)   No Variogram Defined – Used Domain 10 (Waste) Model
13 (Waste)   No Variogram Defined – Used Domain 10 (Waste) Model
20 (Waste) 0.359  St1 0.467 15 1 3.0
 St2 0.174 150 1 15.0
21 (Waste)   No Variogram Defined – Used Domain 11 (Waste) Model
23 (Waste)   No Variogram Defined – Used Domain 20 (Waste) Model
30 (Waste) 0.313  St1 0.415 15 1 3.8
 St2 0.272 88 1 11.0
33 (Waste)   No Variogram Defined – Used Domain 30 (Waste) Model
50 (Waste) 0.393  St1 0.389 14 1 1.0
 St2 0.218 45 1 1.0

TABLE 14-12 ISOTROPIC VARIOGRAM MODELS FOR GOLD – FOOTWALL

Domain Nugget Struct Sill Major (m) Major/ Semi  Major/ Minor
101   No Variogram Defined – Used Domain 150 Model
150 0.323 St1 0.449 15 1 3.0
St2 0.228 70 1 5.8
250 0.308 St1 0.420 18 1 2.3
St2 0.272 65 1 4.3
350 0.316 St1 0.314 15 1 3.0
St2 0.370 45 1 3.8
450   No Variogram Defined – Used Domain 350 Model
550   No Variogram Defined – Used Domain 150 Model
650   No Variogram Defined – Used Domain 250 Model
15 (Waste) 0.395 St1 0.425 18 1 2.6
St2 0.180 70 1 4.7
25 (Waste) 0.347 St1 0.459 14 1 3.5
St2 0.193 65 1 8.1
35 (Waste)   No Variogram Defined – Used Domain 25 (Waste) Model
55 (Waste) 0.393 St1 0.389 14 1 1.0
St2 0.218 45 1 1.0

TABLE 14- 13 ISOTROPIC VARIOGRAM MODELS FOR GOLD – HANGINGWALL

The modeled variograms resulted in a moderate relative nugget ranging from 21% to 39%. Two spherical structures were used throughout with a moderate amount of variability demonstrated over a short range by the first structure (6-22m) and a longer range of within 18-190m (Footwall) and 45-70m (Hangingwall) for the second structure.

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A robust variogram could not be modeled for all domains so variography from comparable domains was substituted on the basis of geological similarity as indicated in Table 14-12 and Table 14-13.

For the ‘dynamic Kriging’ (refer Section 14.2.7) estimation of gold in the block model the rotation of the variogram models was adjusted to follow the orientation of the search ellipsoid and better fit the orientation of each individual mineralized domain. Search ellipsoids and variogram orientations were individually adjusted to fit within these estimation sub-domains. Anisotropy in the minor direction was introduced in the variogram models at a ratio of 1:4 (for bearing, plunge, dip orientations refer Figure 14-2 & Figure 14-3).

14.2.10      GRADE INTERPOLATION METHODOLOGY

Both the Footwall and Hangingwall Lodes of Cosmo were estimated using a ‘dynamic Kriging’ technique developed by Newmarket Gold and modified by Cube Consulting. It consists of an Ordinary Kriged (OK) estimate within each lode that has been divided into angular sectors. This process utilizes an orientation set for the search and variogram while applying a sector (angular corridor) when constraining the block used in the estimate. Each block is pre-populated with dip and dip direction values drawn from the triangle centroids of the constraining wireframe. To perform an estimate, space (spherical/polar co-ordinate system) is divided into two groups of sectors (10° increments) of both dip direction and dip. An estimate is then performed for each sector.

For blocks with a dip between 40° and 50° and with a dip direction between 60° and 70° an estimate is performed with orientations for dip and dip direction of 45° and 65° respectively. The constraint is simply the upper and lower limits of the sector as described above.

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The estimation methodology used for the Cosmo mineralization style is considered appropriate by Newmarket Gold, based on experience with similar deposit types. In previous estimations it had produced reasonable, unbiased reproductions of the drilling data where areas of adequate sampling were present. Outside areas of adequate sampling the mineral resource classification applied reflects the uncertainty of the estimate. Validation of the model also confirmed the estimation approach for Cosmo was reasonable and appropriate. As an additional validation measure an omnidirectional check estimate (not subject to dynamic Kriging sectors) was undertaken in order to compare global gold results against the dynamic Kriging.

A three dimensional, one pass Ordinary Kriging estimate was run using the ‘dynamic Kriging’ method discussed above to estimate the high cut gold grade 1.0m downhole composite data within each mineralized and waste domain. Table 14-14 and Table 14-15 summarize the estimation search parameters by domain.

A three dimensional, one pass inverse distance weighted ‘check’ estimate was also run to estimate the high cut gold grade 1.0m down-hole composite data within each mineralized and waste domain. Table 14-16 and Table 14-17 summarize the estimation search parameters by domain.

Minimum composites used throughout the mineral resource model were, eight composites in the hangingwall mineralization lodes, 6 composites in the footwall mineralization lodes, 10 composites in the hangingwall waste lodes and eight composites in the footwall waste lodes. Maximum composites used were, 20 composites in the hangingwall mineralization lodes, either 20 or 18 composites in the footwall mineralization lodes, 22 composites in the hangingwall waste lodes and 20 composites in the footwall waste lodes. A block discretization of 1 in X, 2.5 in Y and 2.5 in Z were used throughout.

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Mineralised
Domain
Search Radius (m) Bearing
Sector
Width°
Plunge
Sector
Width°
Major/Semi Major/Minor
100 50 10 10 1 4
110 50 10 10 1 4
120 50 10 10 1 4
130 50 10 10 1 4
200 50 10 10 1 4
210 50 10 10 1 4
220 50 10 10 1 4
230 50 10 10 1 4
300 50 10 10 1 4
310 50 10 10 1 4
320 50 10 10 1 4
330 50 10 10 1 4
400 50 10 10 1 4
420 50 10 10 1 4
430 50 10 10 1 4
500 50 10 10 1 4
600 50 10 10 1 4
5 50 10 10 1 4
10 50 10 10 1 4
11 50 10 10 1 4
12 50 10 10 1 4
13 50 10 10 1 4
20 50 10 10 1 4
21 50 10 10 1 4
23 50 10 10 1 4
30 50 10 10 1 4
33 50 10 10 1 4
50 50 10 10 1 4

TABLE 14- 14 DYNAMIC KRIGING SEARCH PARAMETERS FOR GOLD – FOOTWALL DOMAINS – MINERALIZED AND WASTE

Mineralised
Domain
Search Radius
(m)
Bearing
Sector Width°
Plunge Sector
Width°
Major/Semi Major/Minor
101 50 10 10 1 4
150 50 10 10 1 4
250 50 10 10 1 4
350 50 10 10 1 4
450 50 10 10 1 4
550 50 10 10 1 4
650 50 10 10 1 4
15 50 10 10 1 4
25 50 10 10 1 4
35 50 10 10 1 4
55 50 10 10 1 4

TABLE 14-15 DYNAMIC KRIGING SEARCH PARAMETERS FOR GOLD – HANGINGWALL DOMAINS – MINERALIZED AND WASTE

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Mineralised Domain   Search Radius (m)  Bearing Plunge Dip Major/Semi Major/Minor
100 50 0 0 0 1 4
110 50 0 0 0 1 4
120 50 0 0 0 1 4
130 50 0 0 0 1 4
200 50 0 0 0 1 4
210 50 0 0 0 1 4
220 50 0 0 0 1 4
230 50 0 0 0 1 4
300 50 0 0 0 1 4
310 50 0 0 0 1 4
320 50 0 0 0 1 4
330 50 0 0 0 1 4
400 50 0 0 0 1 4
420 50 0 0 0 1 4
430 50 0 0 0 1 4
500 50 0 0 0 1 4
600 50 0 0 0 1 4
5 50 0 0 0 1 4
10 50 0 0 0 1 4
11 50 0 0 0 1 4
12 50 0 0 0 1 4
13 50 0 0 0 1 4
20 50 0 0 0 1 4
21 50 0 0 0 1 4
23 50 0 0 0 1 4
30 50 0 0 0 1 4
33 50 0 0 0 1 4
50 50 0 0 0 1 4

TABLE 14- 16 INVERSE DISTANCE WEIGHTED SEARCH PARAMETERS FOR GOLD – FOOTWALL DOMAINS – MINERALIZED AND WASTE

Mineralised Domain Search Radius (m) Bearing Plunge Dip Major/Semi Major/Minor
101 50 0 0 0 1 4
150 50 0 0 0 1 4
250 50 0 0 0 1 4
350 50 0 0 0 1 4
450 50 0 0 0 1 4
550 50 0 0 0 1 4
650 50 0 0 0 1 4
15 50 0 0 0 1 4
25 50 0 0 0 1 4
35 50 0 0 0 1 4
55 50 0 0 0 1 4

TABLE 14-17 INVERSE DISTANCE WEIGHTED SEARCH PARAMETERS FOR GOLD – HANGINGWALL DOMAINS – MINERALIZED AND WASTE

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14.2.11      BLOCK MODEL DEFINITION

The primary consideration of the 3D model was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralized zone in the local grid co-ordinate system. The model was rotated to bearing 330°.

All mineralized and waste domains were coded and estimated into a single block model, cosmo_underground_ni43101_eoy2015_depleted.mdl. Table 14-18 presents the 3D block model definition and extents.

  Northing Easting RL
Minimum 1100 4900 270
Maximum 2400 5600 1200
Block Size 5 2 5
Sub-block 2.5 1 2.5

TABLE 14-18 COSMO_UNDERGROUND_NI43101_EOY2015_DEPLETED.MDL BLOCK MODEL DEFINITION

The chosen block size represents approximately half the best data spacing in the Northing direction and a choice in the Vertical and Easting dimension controlled by the need to appropriately represent the volume of the wireframes.

A summary of relevant field names and descriptions is presented in Table 14-19.

Attribute Type Default Description
au_id2 real -99 Au Grade (ppm) from Inverse Distance
au_krig real -99 Au Grade (ppm) from Kriging
density real 2.88 Specific Gravity
krig_var real -99 Au Kriging Variance
lodecode Integer -99 Lode Code
mined Integer -99 Mined Code used for mining depletion
no_samp Integer -99 No of Samples used for interpolation
resclass Integer -99 mineral resource Classification Code

TABLE 14-19 3D BLOCK MODEL ATTRIBUTES

14.2.12      MODEL VALIDATION

Model validations were undertaken on all Footwall and Hangingwall domains. The validations include both mineralized and waste domains and an inspection of the audit documentation of the individual estimation runs, visual inspection of the block outcomes and input data and statistical comparisons of input data and block outcomes. Grade Tonnage curves were used as a means of validating the dynamic Kriged estimate against an inverse distance weighted check estimate.

Statistical comparisons of input data and block model outcomes for the mineralized domains are shown in Table 14-20.

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Domain Cut Composite
Average Grade
g/t Au
Block Model
Average
Grade g/t Au
Variance
%
Un-depleted
Percentage of
Total Gold Oz
Indicated / Measured
Model Average Grade

g/t Au
Variance
%
  Footwall Lodes - mineralization Domains  
100 3.05 2.35 77.0 12.2% 2.40 78.7
110 2.18 1.15 52.8 4.6% 1.53 70.2
120 4.17 3.69 88.5 4.4% 3.70 88.7
130 3.26 2.79 85.6 1.7% 2.97 91.1
200 1.88 1.76 93.6 6.9% 1.78 94.7
210 0.80 0.83 103.8 2.2% 0.97 121.3
220 2.18 2.12 97.2 1.2% 2.12 97.2
230 1.77 1.79 101.1 1.5% 1.85 104.5
300 2.39 1.99 83.3 5.5% 2.04 85.4
310 1.15 1.27 110.4 3.4% 1.21 105.2
320 2.92 2.80 95.9 1.0% 2.83 96.9
330 2.11 2.04 96.7 2.2% 2.03 96.2
400 1.48 1.51 102.0 2.1% 1.54 104.1
420 1.54 1.53 99.4 0.2% 1.55 100.6
430 1.63 1.62 99.4 1.3% 1.75 107.4
500 1.39 1.80 129.5 0.4% - -
600 1.54 1.86 120.8 0.5% - -
  Footwall Lodes - Waste Domains  
10 0.24 0.17 70.8 0.7% 0.21 87.5
11 0.13 0.09 69.2 0.5% 0.12 92.3
12 0.46 0.39 84.8 0.3% 0.42 91.3
13 0.31 0.25 80.6 0.3% 0.29 93.5
20 0.29 0.24 82.8 1.0% 0.27 93.1
21 0.27 0.19 70.4 0.7% 0.24 88.9
23 0.26 0.22 84.6 0.3% 0.24 92.3
30 0.37 0.29 78.4 0.6% 0.36 97.3
33 0.35 0.31 88.6 0.5% 0.35 100.0
50 0.60 0.50 83.3 0.2% - -
  Hangingwall Lodes - mineralization Domains  
101 2.11 1.66 78.7 9.8% 1.81 85.8
150 3.10 3.23 104.2 8.9% 3.23 104.2
250 2.33 2.38 102.1 2.3% 1.76 75.5
350 2.10 2.58 122.9 2.7% 1.74 82.9
450 1.76 1.72 97.7 0.5% - 102.6
550 2.30 1.90 82.6 9.1% 2.36 97.6
650 2.45 2.43 99.2 7.2% 2.39 -
  Hangingwall Lodes - Waste Domains  
15 0.43 0.48 111.6 0.8% 0.39 90.7
25 0.37 0.39 105.4 0.5% 0.41 110.8
35 0.34 0.33 97.1 0.3% - -
55 0.38 0.35 92.1 1.4% 0.40 105.3

TABLE 14-20 MINERALIZED DOMAIN AVERAGE GOLD GRADE (G/T) COMPARISONS

The mineralized domain comparisons display some variation between input and outcome average grades when the total domain is reported. The larger variations in grade appear to occur in areas that generally contain weaker drilling density and in portions of domains that represent the margins of the modeling area. This can be demonstrated with consideration to the 11 Lode with the Measured/Indicated grade variance good at 92.3%, while grade variance for the total domain is reduced 69.2% .

Comparison of the Measured and Indicated portions of the mineralized domains in Table 14-20 above show that for the most significant domains by contained ounces (100, 200, 300, 101, 150 & 550) the comparison to the combined average composite grade agrees within an 11% tolerance.

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Grade Tonnage curves were generated for the combined Footwall Mineralized Lodes and the combined Hangingwall Mineralized Lodes as well as the larger Lodes (100, 150, 101 and 550) in Figure 14-4 to Figure 14-9.

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The Grade Tonnage curves suggest a good replication in results between the two estimate types, with no variation in tonnages and only slight variations between grades. This variation generally increases with respect to the increase in cut-off grade due to the reduced tonnage being reported. This variation in grades is due in part to the directional nature of the dynamic Kriged estimate enabling it to place greater emphasis on high-grade or conversely low grade samples along strike than the inverse distance estimate which will smooth the result with respect to the across strike samples.

14.2.13      MINERAL RESOURCE CLASSIFICATION

The classification of the Cosmo mineral resources was based on information provided by Newmarket Gold and outcomes of the estimation process review undertaken by Cube Consulting. The mineral resource has been classified in accordance with CIM guidelines. Assessment criteria include data integrity, drillhole spacing, sample locations, sampling density, and lode geometry, geological confidence and grade continuity. Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data.

The mineral resource has been classified as Measured, Indicated and Inferred categories. Additionally, exploration targets have also been identified and recorded with the mineral resource estimate for future follow up work, these figures have not been reported in this document and are used for internal purposes only.

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14.2.14      LOGGING

The drilling data provided for this resource estimate contains descriptive explanations of the geology from RC and diamond drill core observations. The logging information was considered of sufficient detail and quality to be used in this estimation at the current level of confidence.

14.2.15      DATA SPACING AND DISTRIBUTION

The Cosmo mineral resource model was subject to varying drill hole density and sample locations in relation to the lode geometry. In most domains the drilling was of regular spacing and sufficient density within the upper/central parts of lodes but subject to decreasing densities and irregular spacing at depth. The block model outcomes at depth in most lodes were considered to be higher risk and are classified with less confidence than the shallower parts. For classification purposes each mineralized domain was considered individually and where sufficient data density was present a classification solid was extruded.

14.2.16      ORIENTATION OF DATA IN RELATION TO GEOLOGICAL STRUCTURE

The orientation of the deposit is interpreted to be close to vertical and the drilling is considered to be appropriately targeted for this geological orientation.

14.2.17      GEOLOGICAL INTERPRETATION

The geological interpretation of the Cosmo deposit was undertaken by Newmarket Gold geologists.

14.2.18      DEPOSIT DIMENSIONS

The mineralized portion of the Cosmo deposit extends within drill testing from 1130 to 2360 meters in Northing, within the Easting plane the dimensions of the mineralization are tightly constrained by drilling extending from 4670 to 5220 meters and vertically the deposit extends within drilling from surface (at approximately 1170mRL) to 290mRL. The dimensions of the mineralization are adequately defined by the available drilling with acceptable extensions beyond data.

14.2.19      ESTIMATION AND MODELLING TECHNIQUES

Refer to Section 14.2.10 above for grade interpolation methodology.

14.2.20      MOISTURE

The estimate has been made on the basis of dry tonnes.

14.2.21      METALLURGICAL FACTORS OR ASSUMPTIONS

No metallurgical factors or explicit assumptions have been used in this mineral resource estimate, except that the estimated gold content is in some proportion able to be liberated from the gangue material. The

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nature of the deposit and its history of mining would indicate that this is a reasonable assumption. Testing of material sourced from depth is required to confirm this.

14.2.22      SELECTIVE ASSUMPTION

The mineral resource estimate contains implicit assumptions of mining selectivity represented by the block size of 5m x 2m x 5m (Y x Z).

14.2.23      MINERAL RESOURCE AND AUDITS AND REVIEWS

No mineral resource audits or reviews have been undertaken on the current mineral resource. However, over previous years several audits have been completed on the methodology and approach for mineral resource Estimation at Cosmo. Any comments or recommendations have been reviewed and implemented as required.

14.2.24      DISCUSSION OF RELATIVE ACCURACY/CONFIDENCE

At this stage no quantitative testing of the accuracy of the estimate or establishment of confidence limits has been undertaken. However, the continual reconciliation of the mineral resource model through the mining and milling of gold material, confidence can be obtained through the accuracy of the results observed. See Section 14.2.3.

14.2.25      MINERAL RESOURCE STATEMENT

The mineral resource statement contains a depleted mineral resource for both the Hangingwall and Footwall Lodes.

The depletion was carried out using underground development and stoping solids as well as an existing pit surface. The “as-mined” solids were taken up to December 31, 2015.

The Cosmo classified mineral resource statements for combined Hangingwall and Footwall Lode Models are tabulated below in Table 14-21. The Table reports depleted resources and with a lower cut-off grade of 2.0g/t Au within the mineralized wireframe interpretations and model.

Mineralized Domains (Au >= 2g/t) 
Domain Tonnes Gold Grade g/t Oz Gold
Measured 1,650,000 3.63 192,500
Indicated 2,987,000 2.99 287,600
Total (Measured and Indicated only) 4,637,000 3.22 480,100
Inferred 678,000 2.76 60,200

TABLE 14-21 MINERAL RESOURCE STATEMENT FOR COSMO MINE COMBINED HANGINGWALL AND
FOOTWALL LODES AT 2.0 G/T GOLD CUT OFF, EFFECTIVE DEC 31 2015

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1.

Mineral resources are stated as of December 31, 2015.

   
2.

Mineral resources are inclusive of mineral reserves, which are set out below.

   
3.

Mineral resources are calculated using these parameters:


  d.

Gold price of $A1,500/oz, metallurgical recovery of 92.0%; and

  e.

Lower cut-off of 2.0 g/t Au is used to calculate the mineral resources.


4.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces.

   
5.

Mineral resource estimate was prepared by Mark Edwards, B.SC. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold.

   
6.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.

The mineral resource for the Cosmo Project has been depleted to December 31, 2015. The mineral reserves as stated in Section 15 have also been depleted to December 31, 2015.

14.2.26      RECOMMENDATIONS

In order to improve the quality of the estimated mineral resource the following actions are recommended:

There are no known situations where the mineral resources outlined above could be materially affected by environmental, permitting, legal, title, infrastructure, metallurgical treatment, socio-economic or political issues, other than as outlined elsewhere in this technical report. There is, however, some risk, as with any gold mineral resource where the gold price achieved may affect the overall economic viability of a mining operation.

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14.3   UNION REEFS AREA

    Indicate mineral resource Inferred mineral resource

Project

Deposit
Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
Union Reefs Prospect Claim 2.0 450,000 5.07 73,200 2.0 380,000 7.23 88,400
Crosscourse E-Lens 1.0 2,301,000 1.85 136,900 1.0 479,000 1.96 30,200
Crosscourse Western Lode 2.0 191,000 3.67 22,500 2.0 96,000 4.05 12,500
Low-Grade Stockpiles         N/A 260,000 0.75 6,300
Esmeralda 0.5/2.0 558,000 2.08 37,300 0.5/2.0 142,000 2.60 11,800
Lady Alice 0.5 68,000 1.88 4,100
Millars/Big Tree/PingQue 0.5 523,000 1.79 30,100
Orinoco 0.5 80,000 1.32 3,400 0.5 17,000 2.42 1,300
Union North   0.5 559,000 1.52 27,300
Union South/ Temple   0.5 818,000 1.33 35,000
Sub-total   3,579,000 2.38 273,300   3,342,000 2.30 246,900

TABLE 14-22 MINERAL RESOURCE ESTIMATIONS NEWMARKET GOLD DEPOSITS UNION REEFS AREA

Notes for Table 14-22:

1.

Mineral resources are stated as of December 31, 2015.

   
2.

Mineral resources are inclusive of mineral reserves.

   
3.

Mineral resources are calculated using these parameters.


  f.

Gold Price of $A1,500/oz, metallurgical recovery of 90.0% depending on mineral resource.

  g.

Lower cut-off of 2.0g/t Au is used to calculate the mineral resources for Underground deposit and 0.5g/t Au for open pit mineral resources. A lower cut of 1.0 g/t Au for underground mineral resources at Crosscourse due to size of potential deposit.

  h.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces.

  i.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.


4.

The mineral resource estimates were prepared by Mark Edwards, B.Sc. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold who has over 18 years of relevant experience and is a qualified person for mineral resources as per the NI43- 101.


14.3.1      INTRODUCTION

At this point in time there are no known events or situations, which would materially affect the mineral resource as stated for Union Reefs deposits, these include metallurgical, social, permitting, political, legal or environmental impacts.

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During 2015 all resource estimations were reviewed. Further work will be required on some of the mineral resources but all have been reviewed as being suitable for reporting. Newmarket Gold has decided to keep all mineral resources at Union Reefs (excluding the Esmeralda deposit) the same as the 2013 statement if no additional on ground work (drilling/sampling/mining) has been completed, this means that only the mineral resources to change are for the Esmeralda deposit. The Author believes that any change in optimization results from 2013 would be minimal and therefore not material.

The optimization process for the Esmeralda open pit mineral resources was as follows;

  (i)

Model imported into MineMap™ software for processing:


  a.

Model was optimized using the Lerch-Grossman Pit Optimizer. This optimizer uses several inputs, which are detailed as below.

  b.

Average density (SG) was set as the oxide density value on unpopulated blocks. Assigned density values in the models were used for populated blocks.

  c.

Gold price used was $A1,500 per ounce.

  d.

Process recovery was set at 90% for all oxide and 85% for all fresh material; this reflects the average recovery seen through the process plant.

  e.

Fixed processing cost was set at $A22.50 per tonne, this reflects the costs during 2012 when open pit mining was taking place.

  f.

Mining costs were taken from the current contract estimates of $4.50 per tonne of oxide material and $5.00 per tonne of fresh material.

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  g.

The pit wall angle was also utilized; this was set to 40o for material mined within oxide and 50o for material mined below the oxide zone. These figures are generally what are used in Newmarket Gold’s current mining areas but will need more detailed review before mining can commence in new mining areas.

  h.

When all these parameters are added to the optimizing process, an LG pit shell is generated for reporting.

  i.

For all optimizations of 2012 mineral resources, this LG1450 pit shell was then exported out of MineMap™ and imported into Micromine™ software. At this point the mineral resource estimation was then coded for material above the pit shell and below the current surface (which could be the mined surface).

  j.

This coded data was then exported into Microsoft™ Excel for processing. The block size and cut-off grade was used to determine the tonnes and grade of material within the shell. All these model calculation spreadsheets have been saved for future use and review.

  k.

The optimized numbers were then entered into the current mineral resource statement above but only for mineral resources that have had new models completed in the past two years.


Project Deposit Mineral
resources
Type New
Model
QA/QC
2011/12
SG
2011/12
Twinned
Holes
Model
Constructed by
Year of
Model
Union Reefs Prospect Ind & Inf UG N Y Y Y Cube 2012
Low-Grade Stockpiles Inf OP N N N NA URGM -
Esmeralda Ind & Inf OP Y Y Y Y Newmarket Gold 2015
Lady Alice Inf OP N Y N Y Makar 2003
Millars/Big Tree/Ping Que Inf OP N Y N Y Makar 2003
Orinoco Ind & Inf OP N Y Y Y Cube 2012
Crosscourse Ind & Inf UG N Y Y Y Cube 2013
Union North Inf OP N Y N Y Makar 2003
Union South/ Temple Inf OP N Y N Y Makar 2002

TABLE 14-23 MODEL SUMMARY FOR UNION REEFS DEPOSITS

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Project mineral resource Method Grade cap
Au g/t
Block size
E x N x RL
(meters)
Union Reefs Prospect OK 30g/t (All) 1 x 20 x 20 (Vein)
2 x 10 x 5 (Stockwork)
Low-Grade Stockpiles Mining - -
Esmeralda OK 10- 13g/t 2.5 x 10 x 5
Lady Alice ID 25g/t 2.5 x 10 x 2.5
Millars/Big Tree/Ping Que ID 20g/t 2.5 x 10 x 2.5
Orinoco OK 5, 8 & 10g/t depending on lode 2.5 x 10 x 2.5
Crosscourse 2D OK (West)
MIK E-Lens
10g/t (West) NA
(E-Lens)
2 x 10 x 5 (West)
5 x 25 x 25 (E-Lens)
Union North ID 20g/t 2.5 x 10 x 2.5
Union South/ Temple ID 20g/t 2.5 x 10 x 2.5

TABLE 14- 24 UNION REEFS DEPOSITS MODEL SUMMARY OF MODEL INPUTS

14.3.2      PROSPECT DEPOSIT

14.3.2.1   Introduction

During September 2012, Cube Consulting Pty Ltd was requested by Crocodile Gold to undertake a mineral resource estimation update of the Prospect deposit. The estimation incorporated a number of recently drilled infill holes as detailed in Section 10. Figure 14-11 shows a plan view of the drilling on the Prospect used for this mineral resource estimation. The majority of grade control drilling during past mining was RC, shown in green. All new drilling was diamond core shown in red. This modeling has been reviewed by the Author and is defined as approapraite for reporting.

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Figure 14-12 shows a typical cross section through the deposit at Section 7325mN in the local grid system, looking north. It shows the two most significant mineral resource interpretations that of stockwork domain 400 and the internal vein domain 40. Also shown are the RC grade control, RC exploration and diamond core exploration hole traces utilized in defining the mineral resource.

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Indicated mineral resource Inferred mineral resource
 Deposit Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
100 2 3,035 2.38 231 2      
200 2 4,002 2.58 332 2 131,859 3.38 14,345
30 2       2 54,688 19.61 34,513
300 2 537 2.11 36 2      
31 2       2 3,598 12.17 1,408
310 2 1,515 2.25 109 2 10,046 2.10 679
40 2 99,360 12.44 39,645 2 94,658 10.10 30,764
400 2 201,606 2.71 17,512 2 81,952 2.21 5,815
41 2 4,246 11.59 1,579 2 1,168 18.90 710
410 2 76,508 3.36 8,248 2      
500 2 58,170 2.91 5,436 2      
600 2 1,022 2.17 71 2 2,031 2.53 165

TABLE 14- 25 PROSPECT DEPOSIT LOAD SUMMARY

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14.3.2.2   Data Types

The estimation of contained gold has been based on assays sourced from drilling data, detailed in Section 10, above. The data available as at August 12, 2012 consisted of RC grade control, RC exploration and diamond core samples.

All data is in local grid co-ordinates.

The RC grade control drilling provides close spaced data within the existing pit boundary and the RC exploration drilling provides data in close proximity to the pit area to moderate depths (up to 150m below surface). The diamond core drilling provides definition of the mineralization at greater depths up to 500m below surface. Due to the difference in spatial coverage of the data types, all data types have been used in this estimate.

The total database supplied consisted of 2,450 drill holes, including 2,088 (for 18,456.48m) RC grade control holes, 270 for 23,431m RC exploration holes and 92 for 19,446m and 14 diamond core hole.

Within the mineralized domains the drill data consisted of 1,062 RC intercepts for 4,344.24 downhole meters and 255 diamond core intercepts for 649.47 downhole meters.

The drill cuttings and core are sampled and assayed on varying lengths as summarized within mineralized lodes in Table 14-26.

Mineralized Domain Minimum Length (m) Maximum Length (m)
100 0.27 5.0
200 0.3 2.0
300 0.2 2.0
310 0.31 1.51
400 0.21 3.0
410 1 2.7
500 0.3 2.0
600 0.33 1.0
20 0.26 1.01
30 0.24 1
31 0.3 1
40 0.26 1.5
41 0.34 1

TABLE 14-26 PROSPECT DEPOSIT SUMMARY OF SAMPLE LENGTHS BY MINERALIZED DOMAIN

Geological Interpretation

The Prospect deposit is interpreted to be a steeply dipping semi-continuous gold and silver mineralized quartz stockwork domain, containing, at times, a centralized core of elevated gold mineralization associated with steep dipping quartz veining. The steep dipping quartz veining domain is often associated with visible gold occurrences.

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The host lithologies of the Prospect deposit have been logged systematically for varying states, two is considered oxidized; three and four transitional and five un-oxidized. Surfaces interpreted from this geological logging have been used to flag oxidation state into the 3 dimensional block model.

An examination of the possible effects the state of oxidation has on the grade tenor has been undertaken. Figure 14-13 shows the comparison of gold sample grades by oxidation state for gold plus 0.4g/t Au. The following inferences can be drawn from the figure:

An examination of the interpreted surfaces with close inspection of the boundary behavior of grades was undertaken to test support for the application of the oxidation boundaries during estimation. The interpreted boundaries show a degree of variability in position from hole to hole, related to the intensity of shearing and degree of alteration this means the position of the boundary is not exact. Additionally, the grade transitions across the boundaries are shown to be graduation.

As a consequence, the Company has not used the interpreted oxidization surfaces as hard boundaries during the estimation of gold, despite the statistical differences.

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14.3.2.4   MINERAL RESOURCE INTERPRETATION

Based on the interpreted geology, the mineral resource interpretation incorporated two distinct domain types, stockwork and vein. Two basic assumptions underlie this interpretation; firstly, that the vein domain would be mined completely from hangingwall to footwall incorporating dilution where required to form a minimum mining width from the surrounding stockwork mineralization; secondly, that the vein material would be readily identified visually at the mining face.

Interpretation of mineralized domains has been informed by gold cut-off grade, with a lower limit of 0.4g/t used as the basis for defining mineralized stockwork material and a lower limit of 5g/t gold used to define the vein domains within the stockwork material. No minimum length criteria have been applied during the interpretations.

The mineral resource stockwork interpretations were wireframed and nominated 100, 200, 300, 310, 400, 410, 500 and 600. Figure 14-14, below shows the stockwork wireframes and drilling traces in an oblique view to the northwest. Lodes 500 and 600 are small lodes located on the northwestern edge of, and obscured by Lode 400.

Within these stockwork domains a number of mineral resource vein interpretations were wireframed and nominated 20 (inside 200), 30 (inside 300), 31 (inside 310), 40 (inside 400) and 41 (inside 410). Figure 14-15 below shows the vein wireframes in an oblique view to the northwest.

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The mineral resource wireframes were used to code the drill intercepts contained within them by flagging into a new table in the database, the zonecode table. This flagging allows the selection of data within domains by codes for the purposes of sample analysis and compositing.

All mineral resource interpretation wireframes have been used as hard boundaries for this estimate.

The estimation of the two mineral resource domain types was undertaken using different methodologies. Within the stockwork domains Ordinary Kriging into a 3D block model has been used to interpolate grades, while within the vein domains a 2D accumulation method has been used (see Section 14.4.2.7) .

14.3.2.5   COMPOSITING AND STASTICS

Compositing of the raw drilling sample data is necessary to establish a single support for the data to avoid bias when calculating statistics and undertaking any estimation of the data into three dimensional volumes. A number of items are considered when selecting an appropriate composite length; they included the original support of the raw sample data, the assumed selectivity (and therefore the block sizes) of the model and the imposed spatial dimensions of the mineralized domains.

An examination of sample statistics for mineral resource domains reveals that the majority of sampling of the mineralization is on 1m downhole support. Within the stockwork domains sample lengths vary from a minimum of 0.2 to a maximum of 5.0m downhole. Within the vein domains (coded 20 to 41) downhole sample lengths vary from 0.24 to 1.5m.

The number of instances of samples over 1m are small representing 3% of the data in each domain type. In the vein mineralized domains 9 of 308 samples are greater in length than 1m and within the stockwork domains there are 181 samples greater than 1.0m from a total of 5,269 samples.

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Within the stockwork domains (200 to 410) the drill samples were composited to 1m downhole using a best fit algorithm to provide equal support data for estimation. The best fit compositing method was used with a tolerance of 50% set, yielding composites of between 0.5 and 1.5m in down hole length.

Within the vein domains (20 to 41) the drill samples were interval composited down the entire coded zone. This compositing method yields a single composite of varying length for each coded interval. The interval composites extracted varied in length from 0.48 to 5m. Further processing is required to produce an additive variable of equal support. The process used was to calculate a horizontal width of the vein domain at each interval composite centroid and multiply the grade by the calculated width. The horizontal width used for each intercept grade has been calculated directly from the width of the interpreted wireframe at the midpoint of the downhole coded intercept. Horizontal widths calculated in this way vary from 0.3 to 2.22m and are detailed by domain.

The resulting variable is an accumulation variable of equal support (assuming a constant density) at each sampled point within the vein domains. This accumulation variable is suitable for the calculation of statistics and for use in estimation.

The effect of a small number of outlier composite grades or spatially isolated composites may have an undue effect on the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the composite data in 3D visualization.

A statistical summary of the stockwork domains is shown in Table 14-28, below. It should be noted that statistical summaries and charts of the gold grade within the vein domains are not strictly representative as the gold intercept grades are not on equal support. Gold intercept composite statistics have been used in this instance to identify the need for high cuts and demonstrate the overall effect of cuts in the vein domains.

A number of high cuts or limits were identified as necessary within domains as detailed in Table 14-27, below. High-grade gold cuts were applied to the 1m composites within the stockwork domains and the cut composites used in the estimation. The high-grade cuts found to be required in the vein domains were applied to the gold intercept composites prior to the multiplication by horizontal width.

No horizontal width outliers were identified or cut.

Domain Min Gold
Grade
g/t
Maximum
Gold Grade
g/t
Mean Gold
Grade
g/t
High Grade Cut
g/t Au
Mean Cut Gold
Grade g/t
20 0.99 15.20 7.28 NA 7.28
30 1.02 47.72 18.60 30 15.83
31 8.91 14.51 12.56 NA 12.56
40 0.07 451.10 17.89 30 11.604
41 0.29 79.80 20.67 30 10.99
100 0.005 17.81 0.59 NA 0.59
200 0.005 38.07 1.16 30 1.12
300 0.005 12.42 0.80 NA 0.80
310 0.015 12.06 1.96 NA 1.96

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Domain Min Gold
Grade
g/t
Maximum
Gold Grade
g/t
Mean Gold
Grade
g/t
High Grade
Cut

g/t Au
Mean Cut
Gold
Grade
g/t
400 0.005 104.63 1.68 30 1.57
410 0.005 70.50 1.96 30 1.73
500 0.005 116.00 2.09 30 1.66
600 0.005 11.80 0.92 NA 0.92

TABLE 14-27 PROSPECT DEPOSIT HIGH-GRADE COMPOSITE CUTS BY DOMAIN

Domain Number Cut Mean
g/t Au
Cut Median
g/t Au
Standard
Deviation
Co-efficient of
Variation
100 930 0.59 0.14 1.69 2.84
200 235 1.13 0.45 2.57 2.28
300 212 0.80 0.42 1.39 1.72
310 34 1.96 1.47 2.31 1.18
400 1875 1.57 0.66 3.25 2.07
410 598 1.73 0.41 4.52 2.61
500 507 1.65 0.59 3.94 2.38
600 184 0.92 0.41 1.57 1.72

TABLE 14-28 PROSPECT DEPOSIT STATISTICAL SUMMARY, STOCKWORK DOMAINS

The general statistics of gold composites within all domains can be described as positively skewed with moderate to high variability as is the case with most gold occurrence. The high variability is reduced somewhat by high cutting of gold grades in those domains most affected.

Within the stockwork domains a measure of variability the Co-efficient of Variation (CV) remain at two or more, indicating significant variability remains within the domains after high grade cutting.

Within the vein domains the use of the high-grade cutting and the accumulation variable considerably reduces the variability of the raw gold data for all domains, reducing the CV to less than one in all domains.

14.3.2.6  Variography

Variography was used to characterize the spatial behavior of the composite data primarily, as an aid to establishing estimation parameters. Variogram stability and quality is dependent on the statistical properties of defined domains and the amount of data available within domains. After an initial investigation, two models were established, one for the stockwork and one for the vein domains. The final model established for the vein domains was based on a modeling of data from domain 40 the most populous with 129 accumulation composite data, all other vein domains contain insufficient numbers of composite data. The stockwork domains were grouped together for final modeling of the variography. This was considered the most robust solution to very noisy models observed within individual stockwork domains as a result of moderate to high domain variability.

The estimation of the vein domains requires an additional variogram model for the horizontal width variable as detailed in Section 14.3.2.7, below. In this instance the model used for the estimation of horizontal width has been set to the accumulation variable variogram. The modeled horizontal width

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variogram was modeled and found to be proportional to that of the accumulation. This finding and the advantage of internal consistency between accumulation and horizontal width support this approach.

The final variogram models are detailed in Table 14-29, below.

Domain Nugget Structure Sill Range
m
Azimuth Plunge Dip Major/Semi Major/Minor
Vein (40) 218 1 1872 53 0 20 0 2.0 2.0
Stockwork 16.5 1 5.6 15 0 0 -90 1.0 1.0
Combined   2 1.5 130 0 0 -90 1 2.77

TABLE 14- 29 PROSPECT DEPOSIT FINAL VARIGRAM MODELS BY DOMAIN

The variogram model for vein domains was modeled as a single spherical structure with a 10% nugget. The axis of greatest continuity was observed to plunge to the south at 20°. The variogram model for stockwork domains was modeled with two spherical structures and a 70% nugget. The first structure includes 24% of the sill within 15m and the second the remaining 6% of the sill within a range of 130m. The variogram models are quite different and appropriately reflect the differences seen in the summary statistics of the composite and accumulation data.

14.3.2.7  Grade Interpolation Methodology

There are several key physical features of the Prospect vein domains that need to be considered and accommodated by the selected mineral resource modeling technique. These features are:

A method that allows estimation of metal content on a projected plane is considered to be an appropriate approach to addressing the features outlined above. This mineral resource modeling approach is best achieved using geological intercept composites and accumulation estimation. Vein domain gold grades are composited across the entire coded interval resulting in a single geological intercept composite at each intercept location. The geological composites are projected onto a vertical 2D plane approximately parallel with the vein structure. The mid-point of each geological composite is assigned the horizontal width of the vein structure and used to compute a ‘metal accumulation’ variable.

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Geological intercept composites are not of equal support as the lode thickness varies. Given the variable thickness, an additive variable must be created as the product of grade and thickness or ‘metal accumulation’ variable. The accumulation a(x) is defined as the product of thickness t(x) and grade z(x) assuming a constant density:

a(x) = t(x) . z(x)

This variable is estimated along with the horizontal width t(x) into each block and a final grade back calculated by dividing the estimated accumulation by the estimated horizontal width at each block centroid.

The project plane model is defined as 1 block thick, and post estimation was re-located into real world co-ordinates and imported into the 3D block representation of the domain. This process assigns a single grade for each estimated block from hangingwall to footwall of the vein domain wireframed blocks.

Estimation of the two variables accumulation and horizontal width has used a two pass Ordinary Block Kriging. The same parameters have been used for both attributes in each domain. The variogram models are detailed in Table 14-30 and the estimation parameters are detailed in Table 14-31. A constant minimum of 4 and maximum of 8 (pass 2 minimum 2) data have been set and a discretization of 1 in X, 5 in Y and 5 in Z has been used throughout. The optimal first pass search radii have been increased in all vein domains to fill the interpreted wireframe volumes.

Domain Nugget Structure Sill Range Azimuth Plunge Dip Major/Semi Major/Minor
20 218 1 1872 53 0 0 0 1.0 1.0
30 218 1 1872 53 0 0 0 1.0 1.0
31 218 1 1872 53 0 0 0 1.0 1.0
40 218 1 1872 53 0 20 0 2.0 2.0
41 218 1 1872 53 0 0 0 1.0 1.0

TABLE 14-30 PROSPECT DEPOSIT VEIN DOMAIN ESTIMATION VARIOGRAM MODELS

Domain Search
Radius m
Pass2
Radius m
Azimuth Plunge Dip Major/Semi Major/Minor
20 60 180 0 0 0 1.0 1.0
30 60 150 0 0 0 1.0 1.0
31 60 180 0 0 0 1.0 1.0
40 60 180 0 20 0 2.0 2.0
41 60 180 0 0 0 1.0 1.0

TABLE 14-31 PROSPECT DEPOSIT VEIN DOMAIN ESTIMATION PARAMETERS

Stockwork Domains

A standard three dimensional single or two pass Ordinary Kriging methodology has been used for the estimation of the cut gold 1 meter down hole composite data within each stockwork domain. Table 14-32 details the variogram models used and Table 14-33 summarizes the estimation parameters by domain. A constant minimum of 4 and maximum of 25 (pass 2 minimum 4) data have been set and a discretization of 5 in X, 5 in Y and 1 in Z has been used throughout. The optimal first pass search radii have been increased in selected domains to fill the interpreted wireframe volumes.

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Domain Nugget Structure Sill Range m Azimuth Plunge Dip Major/
Semi
Major/
Minor
100 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77
200 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77
300 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77
310 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77
400 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77
410 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77
500 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77
600 16.5 1 5.6 15 0 0 -90 1.0 1.0
    2 1.5 130 0 0 -90 1.0 2.77

TABLE 14-32 PROSPECT DEPOSIT STOCKWORK DOMAIN ESTIMATION VARIOGRAM MODELS

Domain Search
Radius m
Pass2
Radius
m
Azimuth Plunge Dip Major/Semi Major/Minor
100 130 NA 0 0 -90 1.0 2.7
200 130 195 0 0 -90 1.0 2.7
300 130 195 0 0 -90 1.0 2.7
310 130 NA 0 0 -90 1.0 2.7
400 130 195 0 0 -90 1.0 2.7
410 130 NA 0 0 -90 1.0 2.7
500 130 NA 0 0 -90 1.0 2.7
600 130 NA 0 0 -90 1.0 2.7

TABLE 14- 33 PROSPECT DEPOSIT STOCKWORK DOMAIN ESTIMATION PARAMETERS

14.3.2.8  Block Model Definition

Vein Domains

The primary consideration of the 2D model was to provide an appropriate block size for the interval composite data spacing in long section view. Figure 14-16 is a long section view of the domain 40 interpreted wireframe and composite data. The 20x20m grid demonstrates that within the best sampled parts of the domain, above RL 980 a choice of 20x20m block size ensures the majority of blocks contain at least one composite data.

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Table 14-34 summarizes the base definition used for the 2D block models. A separate block model was created for each vein domain. All individual vein domain models were estimated, re-located into real world co-ordinates and ultimately combined into the 3D block model defined below in the local grid coordinate system.

  Northing Easting RL
Minimum 7000 0 700
Maximum 7800 1 1300
Block Size m 20 1 20
Sub-block m NA NA NA

TABLE 14-34 PROSPECT DEPOSIT VEIN DOMAIN PROJECTION BLOCK MODEL DEFINITION

A standard list of field names and descriptions used in the block model are shown in Table 14-35, below.

Attribute Type Default                                        Description
Au Float -1 Back Calculated Gold ppm
H width Float -1 estimated Horizontal Width
Accum._au Float   estimated Accumulation AuxHW
Ads Float 0 Average distance to composite data

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Attribute Type Default Description
Dns Float 0 Distance to nearest composite data
Kv Float -1 Kriging Variance
Ns Integer 0 Number of composite data
Pass Integer -1 Estimation Pass Number

TABLE 14- 35 PROSPECT DEPOSIT VEIN DOMAIN PROJECTION BLOCK MODEL ATTRIBUTES

Stockwork Domains

The primary consideration of the 3D model was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralized zone. All individual mineralized zones were ultimately combined to create a single block model in the local grid coordinate system. Table 14-36 summarizes the 3D block model “Prospect_Sep2012.mdl” definition.

  Northing Easting RL
Minimum 7000 4700 700
Maximum 7800 4950 1300
Block Size m 10 2 5
Sub-block m 5 0.5 2.5

TABLE 14-36 PROSPECT DEPOSIT STOCKWORK AND FINAL 3D BLOCK MODEL DEFINITION

The chosen block size represents approximately half the best data spacing in the Northing direction and a choice in the vertical and easting dimension controlled by the need to appropriately represent the volume of the wireframes.

A standard list of field names and descriptions used in the block model are shown in Table 14-37.

Attribute Type Default                                        Description
Au Float 0.01 estimated Gold ppm
H width Float -1 estimated Horizontal Width
Density Float 2.72 Density
Zonecode Char BKGR Zonecode
Rescat Integer 4 Meas =1; Ind = 2; Inf = 3, Waste = 4
Depletion Integer 1 Insitu = 1; pillar = 2; Mined = 0
Oxidation Integer 0 1 = Fresh; 2 = Trans; 3 = oxidized
Ads Float 0 Average distance to composite data
Dns Float 0 Distance to nearest composite data
Kv Float -1 Kriging Variance
Ns Integer 0 Number of composite data
Pass Integer -1 Estimation Pass Number

TABLE 14-37 PROSPECT DEPOSIT STOCKWORK AND FINAL 3D BLOCK MODEL ATTRIBUTES

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Table 14-38 below confirms the close agreement of the 3D block model volumes and the original interpreted wireframe volumes, supporting the 3D model block size choice as appropriate.

Domain Wireframe
Volume
Block Model
Constraint
Volume
Difference
%
20 10701 10744 0%
30 18747 19019 1%
31 1286 1256 -2%
40 69291 69281 0%
41 1934 1913 -1%
100 115033 114919 0%
200 185898 186638 0%
300 253097 253406 0%
310 13090 13044 0%
400 803236 804132 0%
410 101884 102006 0%
500 111921 111575 0%
600 27629 27538 0%

TABLE 14-38 PROSPECT DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK

14.3.2.9  Specific Gravity / Bulk Density Assignment

The specific gravity of the waste and mineralized rock of the final 3D block model has been assigned according to oxidation state, using interpreted surfaces described in Section 14.4.2.3 to control the blocks assigned. The Company has determined the specific gravity for a total of 1,141 diamond core intervals during the latest drilling program. With outliers removed a total of 1,117 of these measurements were used to determine appropriate density average values by oxidation state. Table 14-39 summarizes the oxidation state specific gravity data statistics.

Oxidation State Number Min SG
gm/cm3
Maximum SG
gm/cm3
Mean SG
gm/cm3
Median
SG
gm/cm3
STD
1 to 2 Oxidized 204 2.03 2.95 2.58 2.61 0.434
3 – 4 Transitional 241 2.29 3.00 2.68 2.70 0.109
5 Fresh 672 2.31 4.51 2.76 2.75 0.102

TABLE 14-39 PROSPECT DEPPOSIT SPECIFIC GRAVITY DATA STATISTICS BY OXIDATION STATE

Within the final 3D block model blocks coded fresh were assigned a density of 2.76g/cm 3; those coded transitional, a density of 2.68g/cm 3 and those blocks below the topographical surface and coded oxidized a density of 2.58g/cm 3. Blocks located above the topographical surface were assigned a zero density.

14.3.2.10  Model Validation

Model validation has been undertaken to ensure no material error has been made in the estimation of Prospect Claim. The validations include inspection of the audit documentation of the individual estimation

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runs; visual inspection of the block outcomes and input data; statistical comparisons of input data and block outcomes, and swath plots of the most significant domains.

Statistical comparisons of input data and block model outcomes for the total stockwork domains are shown in Table 14-40.

Domain Composite
Average
Grade g/t
Au
Block Model
Average
Grade
g/t Au
Variance
%
Percentage of
Total Gold
Stockwork
Ounces
Indicated Block
Model Average
Grade g/t Au
Variance%
100 0.60 0.79 133 4 0.73 122
200 1.13 1.58 140 15 1.04 92
300 0.8 0.69 86 9 0.95 118
310 1.96 1.78 91 1 1.66 85
400 1.57 1.38 88 52 1.54 98
410 1.73 1.92 111 8 1.97 114
500 1.66 1.47 88 9 1.56 94
600 0.92 1.03 112 1 1.03 113

TABLE 14- 40 PROSPECT DEPOSIT STOCKWORK DOMAIN AVERAGE GOLD GRADE COMPARISONS

The stockwork domain comparisons display a moderate to large variation between input and outcome average grades when the total domain is reported. As can be confirmed in the visual inspection and swath plot investigations the comparisons include significant volumes at depth in each domain containing a lower density of sample data. This results in extrapolation of the sample data into these volumes and while it is considered a reasonable estimate of the grades in these volumes a simple statistical comparison of total volumes will not result in close comparisons. Figure 14-17 below demonstrates this situation within the stockwork domain 300. Two views of the block model domain 300 are shown side by side, the first with composite data and the second with blocks colored by mineral resource classification, blue for Indicated and green for Inferred classification. Composite data is colored by gold grade distribution as shown in the legend. The contrasting data densities above and below an RL of 1050 are evident, supporting the assumption that relying only on raw composite to block grade comparisons can be misleading.

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Comparison of the Indicated portions of the stockwork domains in Table 14 18 above show that for the most significant domains by contained ounces (400, 200, and 500) the comparisons to average composite grades agree within a 10% tolerance.

Data is analyzed by northing and by elevation for each domain. Reproduced below are four example swath plots one pair for domain 400, Figure 14-18 and one for domain 300, Figure 14-19.

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The swath plots demonstrate that where there is regularly spaced data the block model reflects those data. The plots also highlight that a paucity of drilling data (in particular below an approximate RL of 1050m) results in parts of the model that rely on only a few measured points and are therefore less likely to match local composite data and are of reduced certainty and increased risk.

Statistical comparisons of input data and block model outcomes for the total vein domains are shown in Table 14-41.

Domain Composite
Average
Grade g/t
Au
Block Model
Average
Grade
g/t Au
Variance %
20 7.51 7.54 100
30 13.96 19.62 140
31 12.13 12.16 100
40 13.28 11.12 84
41 10.52 12.92 123

TABLE 14- 41 PROSPECT DEPOSIT VEIN DOMAIN BACK CALCULATED AVERAGE GOLD GRADE COMPARISONS

Again these comparisons show some anomalies; Domains 30, 40 and 41 in particular. Swath plots of Domains 40 (in Figure 14-21) and Domain 30 (in Figure 14-20) show that they are subject to a variable data density spatially, illustrated in Figure 14-8 above. Domain 40 contains 61% of the total contained vein domain ounces of gold while domain 30 contains 30% of the total ounces. Domain 30 is characterized by very few data, and as a consequence is classified as Inferred.

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Domain 40 demonstrates clearly that the estimation represents a close reproduction of the input data where it is consistently and regularly sampled, above 1,050m RL. Below this RL the model is based on few widely spaced data and the estimate deviates from these local data.

14.3.2.11  Mineral resource Classification

The classification of the Prospect deposit mineral resources was based on information provided by Newmarket Gold and outcomes of the estimation processes. The mineral resource has been classified in accordance with the NI43-101 guidelines. Assessment criteria include data integrity, drill hole spacing, sample locations, sampling density, and lode geometry, geological confidence and grade continuity. Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data.

The mineral resource has been classified as Indicated and Inferred. No Measured Resource has been identified.

14.3.2.12  Recommendation

The following points summarize the most relevant recommendations:

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14.3.3      CROSSCOURSE

14.3.3.1  Introduction

During December 2012, Cube Consulting Pty Ltd was requested by Crocodile Gold to undertake a mineral resource estimation of the remaining Crosscourse deposit, consisting of a number of relatively high-grade mineralization shoots below and to the north of the existing Crosscourse Pit at Union Reefs Gold Mine (URGM). A substantial volume of the mineralization was extracted during bulk mining of the Crosscourse Pit, but exploration and grade-control drilling indicates that the mineralization shoots continue down-plunge below the lowest mined level of the pit, and remain open at depth. Figure 14-22 shows a plan view of the drilling coverage and hole types of the Crosscourse area used for this mineral resource estimation

For modeling and estimation purposes, the Crosscourse area has been divided into the Union Reefs West (Lode domain 1001) and E-Lens (Lode domain 100 and 200) areas outlined in Figure 14-23.

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Figure 14-23, Figure 14-24 and Figure 14-25 show typical cross sections through the deposit at Sections 6360mN, 6790mN, and 6960mN looking north in the local grid system. They show the location of the mineralized domains in relation to the open pit. Also shown are the percussion, RC and diamond core exploration hole traces on which the lode interpretation and mineral resource is based.

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14.3.3.2  Data Types

The estimation of contained gold has been based on assays sourced from drilling data, detailed in Section 10 above. The data available consisted of RC drilling, diamond core and undefined hole type samples from historic exploration and mining definition campaigns.

All data is in local grid co-ordinates.

The older drilling (RC and unknown hole types) provides close spaced data from surface prior to mining excavations and the newer RC drilling from pit benches provides data infilling gaps from the previous drilling to the pit area to moderate depths (up to 180m below surface). The diamond holes covering the lodes provide core data for definition of the mineralization at greater depths up to 500m below surface and pit excavations. The database includes 273 RC grade control holes drilled from the lower flitches of the Crosscourse open pit on a 10mN x 5mE grid. No other grade control sampling data was included in the Crosscourse databases. The total database supplied is summarized in Table 14-20 below;

Hole Type # of
Holes
Meters
Drilled
Average
Depth m
RC 395 15,809.0 40.0
DDH 172 31,627.7 183.9
Unknown 425 40,920.9 96.3
TOTAL 992 88,357.6 106.7

TABLE 14- 42 SUMMARY OF DRILLING STATISTICS FOR THE D ATA SET COVERING THE CROSSCOURSE DEPOSIT LODES.

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Within the Union Reefs West deposit (URW) and Crosscourse deposit E-Lens mineralized domains the drill data consisted of 286 drill intercepts for 7,499.8 downhole meters for all drill types. A visual comparison in section between the unknown holes types, RC and diamond holes was completed to test if any material difference was observed between data types. There are several examples where variability in grade continuity in between data types was present; however, it was not possible to determine whether the difference was due to the difference in data type or the inherent variability of the deposit. In general, for the case of the intersections within the URW mineralized domain a gradational drop off in grade down the hole is observed, regardless of hole type. On the other hand, within the E-Lens lodes mineralized intersections show an irregular variation of grades down the hole regardless of hole type.

Hole
Type
# of
Intersections
Minimum
Length (m)
Maximum
Length (m)
RC 131 1 48
DDH 84 2.94 102.5
Unknown 71 3 92
TOTAL 286 1 102.5

TABLE 14-43 SUMMARY OF SAMPLE LENGTHS BY HOLE TYPE FOR CROSSCOURSE DEPOSIT MINERALIZED DOMAINS

The Company has decided to include all data types in this estimate to improve data density. The majority of historical drilling of unknown type is located within the mined out top of lodes. The most recent drilling RC and diamond core informs the remaining parts of lodes. The level of confidence in the data is reflected in the mineral resource classification in Section 14.4.3.13.

14.3.3.3  mineral resource Interpretation

Interpretation of the mineralized domain has been informed by gold cut-off grade, with a lower limit of 0.7g/t Au in the case of the URW and 0.4g/t Au in the case of the E-Lens used as the basis for defining mineralized veining material. No minimum length criterion has been applied during the interpretations.

The mineral resource domain interpretations were wireframed and nominated domain 1001 for the URW Lode and domains 100 and 200 for the E-Lens Lodes.

Figure 14-26 below shows the Crosscourse Pit outline, Union Reefs West and E-Lens lode wireframes and drilling traces in an oblique view looking to the northeast.

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The mineral resource wireframes were used to code the drill intercepts contained within them by flagging into a new table in the database called the zonecode table. This flagging allows the selection of data within the lode domain by codes for the purposes of sample analysis and compositing.

The mineral resource interpretation wireframes have been used as a hard boundary for this estimate.

The estimation of the two mineral resource Lodes was undertaken using two different methodologies. Within the Union Reefs West Lode a 2D accumulation method has been used with one geological composite interval. On the other hand, within the E-Lens Lodes Multiple Indicator Kriging using 5.0m downhole composites has been used to estimate grades.

14.3.3.4  Compositing and Statistics

Compositing of the raw drilling sample data is necessary to establish a single support for the data to avoid bias when calculating statistics and undertaking any estimation of the data into three dimensional volumes. A number of items are considered when selecting an appropriate composite length; they include the original support of the raw sample data, the assumed selectivity (and therefore the block sizes) of the model and the imposed spatial dimensions of the mineralized domains.

An examination of sample statistics for mineral resource domains reveals that the majority of sampling of the mineralization is on 1m downhole. Within the URW Lode (coded 1001) sample lengths vary from a minimum of 0.3 to a maximum of 1.03m downhole. Within the E-Lens Lodes (coded 100 and 200) downhole sample lengths vary from 0.3 to 3m. The number of instances of samples over 1m are small representing less than 1% of the data in each domain type.

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Within the E-Lens Lodes the drill samples were composited to 5m downhole using a best fit algorithm to provide equal support data for estimation. The best fit compositing method was used with a tolerance of 50% set, yielding composites of between 2.5 and 5 meters in down hole length.

Within the URW Lode the drill samples were interval composited across the entire coded zone. This compositing method yields a single composite of varying length for each coded interval. The interval composites extracted varied in length from 2.94 to 34m in length. Further processing is required to produce an additive variable of equal support. The process used was to calculate a horizontal width of the Lode at each interval composite centroid and multiply the grade by the calculated width. The horizontal width used for each intercept grade has been calculated directly from the width of the interpreted wireframe at the midpoint of the down hole coded intercept. Horizontal widths calculated in this way vary from 1.44 to 9.43m.

The resulting variable is an accumulation variable of equal support (assuming a constant density) at each sampled point within the lode. This accumulation variable is suitable for the calculation of statistics and for use in estimation.

The effect of a small number of outlier composite grades or spatially isolated composites may have an undue effect of the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the composite data in 3D visualization.

A statistical summary of the E-Lens domains is shown in Table 14-44, below. It should be noted that statistical summaries and charts of the gold grade within the URW domain are not strictly representative as the gold intercept grades are not on equal support. Gold intercept composite statistics have been used in this instance to identify the need for high cuts and demonstrate the overall effect of cuts in the vein domains.

A high cut or limit was identified as necessary within URW domain as detailed in Table 14-45 below. The high-grade gold cut was applied to the gold intercept composites prior to the multiplication by horizontal width. No high-grade cuts were required for the E-Lens domains given that the MIK estimation methodology used can handle outliers in a more appropriate manner. No horizontal width outliers were identified or cut.

Area Domain Number Cut
Mean
g/t Au
Cut
Median
g/t Au
Standard
Deviation
Co-efficient
of
Variation
URW 1001 27 3.19 1.55 3.20 1.00
E-Lens 100 1333 2.30 1.52 3.22 1.40
E-Lens 200 109 1.33 0.72 2.31 1.74

TABLE 14-44 CROSSCOURSE AND UNION REEFS WEST DEPOSITS STATISTICAL SUMMARY MINERALIZED DOMAINS

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Area

Domain
Minimum
Gold Grade
g/t Au
Maximum
Gold Grade
g/t Au
Mean Gold
Grade g/t
Au
High
Grade Cut
g/t Au
Mean Cut
Gold Grade
g/t Au
URW 1001 0.059 11.42 3.27 10.00 3.19
E-Lens 100 0.005 59.50 2.30 NA 2.30
E-Lens 200 0.037 16.95 1.33 NA 1.33

TABLE 14-45 CROSSCOURSE AND UNION REEFS WEST HIGH GRADE COMPOSITE STATISTICS BY MINERALIZED DOMAIN

The general statistics of gold composites within all domains can be described as positively skewed with moderate to high variability as is the case with most gold occurrence. The high variability is reduced somewhat by high cutting of gold grades in the domains affected.

Within the E-Lens domains a measure of variability the Co-efficient of Variation (CV) remains at 1.4 or more, indicating significant variability within the domains.

Within the URW domains the use of the high-grade cutting and the accumulation variable considerably reduces the variability of the raw gold data, reducing the CV to 1.

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14.3.3.5  Variography

Variogram models were used to characterize the spatial behavior of the composite data, primarily as an aid to establishing estimation parameters. Variogram stability and quality is dependent upon the statistical properties of defined domains and the amount of data available within domains. The variogram model established for the URW lode was based on the modeling of composite data from Domain 1001 containing 27 accumulation composite data. The E-Lens domains were grouped together for final modeling of the variography. This was considered the most robust solution to very noisy models observed within individual stockwork domains as a result of moderate to high domain variability. Several variograms were modeled for different cut-off grades as required by the MIK estimation methodology.

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The estimation of the URW Lode requires an additional variogram model for the horizontal width variable as detailed in Section 14.4.3.6 below. In this instance the model used for the estimation of horizontal width has been set to the accumulation variable variogram. This approach ensures internal consistency between the accumulation and horizontal width variables.

The final variogram models are detailed in Table 14-46 below.

Area Indicator Nugget Structure Sill Range
m
Azimuth Plunge Dip Major/
Semi
Major/
Minor
URW NA 0.46 1 0.54 25 0 0 0 1 1
E-
Lens
Au >= 0.498 0.394 1 0.39 8 0 0 0 1 1
2 0.21 150 0 0 0 1 1
Au >= 0.902 0.526 1 0.32 8 0 0 0 1 1
2 0.16 125 0 0 0 1 1
Au >= 1.304 0.567 1 0.34 8 0 0 0 1 1
2 0.10 110 0 0 0 1 1
Au >= 1.815 0.653 1 0.22 8 0 0 0 1 1
2 0.12 50 0 0 0 1 1
Au >= 2.602 0.675 1 0.23 8 0 0 0 1 1
2 0.10 50 0 0 0 1 1
Au >= 2.797 0.691 1 0.20 8 0 0 0 1 1
2 0.11 50 0 0 0 1 1
Au >= 4.284 0.714 1 0.18 8 0 0 0 1 1
2 0.10 50 0 0 0 1 1
Au >= 5.8 0.738 1 0.11 8 0 0 0 1 1
2 0.15 50 0 0 0 1 1
Au >= 7.52 0.763 1 0.08 8 0 0 0 1 1
2 0.16 50 0 0 0 1 1
Au >= 9.82 0.833 1 0.06 8 0 0 0 1 1
2 0.10 20 0 0 0 1 1
Au >= 16.329 0.851 1 0.02 8 0 0 0 1 1
2 0.13 15 0 0 0 1 1

TABLE 14-46 CROSSCOURSE AND UNION REEFS WEST DEPOSITS FINAL VARIOGRAM MODELS FOR LODE DOMAIN

The variogram model for the URW domain was modeled as a single, omnidirectional spherical structure with a 46% nugget and a range of 25m. The variogram models for each cut-off of the E-Lens domain were modeled with two spherical structures, 40% to 85% nugget, and ranges from 15m to 150m.

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14.3.3.6  Block Modeling and Estimation

2D ORDINARY KRINKING MODELLING TECHNIQUE – UNION REEFS WEST

When assessing the modeling technique for the narrow Union Reefs West Lode, several key physical features were considered and accommodated by the selected mineral resource modeling technique. These features are:

A method that allows estimation of metal content on a projected plane is considered to be an appropriate approach to addressing the features outlined above. This resource modeling approach is best achieved using geological intercept composites and accumulation estimation. Lode gold grades are composited across the entire coded interval resulting in a single geological intercept composite at each intercept location. The geological composites are projected onto a vertical 2D plane approximately parallel with the vein structure. The mid-point of each geological composite is assigned the horizontal width of the vein structure and used to compute a ‘metal accumulation’ variable.

Geological intercept composites are not of equal support as the lode thickness varies. Given the variable thickness, an additive variable must be created as the product of grade and thickness or ‘metal accumulation’ variable. The accumulation a(x) is defined as the product of thickness t(x) and grade z(x) assuming a constant density:

a(x) = t(x) . z(x)

This variable is estimated along with the horizontal width t(x) into each block and a final grade back calculated by dividing the estimated accumulation by the estimated horizontal width at each block centroid.

The project plane model is defined as 1 block thick, and post estimation was re-located into real world co-ordinates and imported into the 3D block representation of the domain. This process assigns a single grade for each estimated block from hangingwall to footwall of the vein domain wireframed blocks.

Cube utilized 2D projection Ordinary Block Kriging of intercept composites for the grade estimation of the Union Reefs West Lode. Some of the advantages of using this method over 3D block modeling for this domain are:

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A 2D block model consisting of 25m N x 1m E x 25m RL parent cells (longitudinal grid) was created with a single cell 1m thick in the longitudinal plane. Data spacing, geometry of the mineralized zone, and volume fill were the primary considerations taken into account when selecting an appropriate estimation block size. Block discretization points were set to 5(Y) x 1(X) x 5(Z) points. The Isatis’ interpolation module was used to for the grade interpolation process.

Estimation of the two variables accumulation and horizontal width has used a single pass Ordinary Block Kriging. The same estimation parameters have been used for both attributes. A “unique” neighborhood search was used, in which all the samples are used for the estimation of each block.

14.3.3.7  MIK Modeling Technique – E-Lens

Multiple Indicator Kriging (MIK) was selected as the estimation methodology for the E-Lens mineral resource area as this method is known to deal with highly skewed distributions and erratic spatial variability more appropriately than OK. MIK involves the individual Kriging of a set of increasing grade indicators to yield a suite of probability estimates above a range of grade cut-offs. These probability estimates can be used to calculate grade class probabilities, and ultimately a block grade estimate (called the e-type estimate). Prior to calculating grade bin probabilities, any order relations problems must first be rectified (i.e. the estimated probabilities above successively higher cut-offs should always be decreasing). The probability of being within a particular grade bin is then weighted by a "mean" grade for that bin, which is usually calculated from the grade sample data, or by discretization of a function defined using the results of the Kriging. The Isatis software package was used to calculate the MIK estimates.

The indicator grade cut-offs were selected according to the following schema:

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Bin grade statistics were then calculated for each of the intervals between cut-offs, using the sample data. The mean, declustered mean, median and declustered median values were calculated for each bin. The purpose of this exercise was to provide the means to decide on which bin cut-offs to use for eventual calculation of the e-type estimate. Table 14-47 summarizes the indicator bin statistics for gold and Table 14-48 summarizes the MIK search strategies.

Bin g/t Au No Samples Mean
g/t Au
Median
g/t Au
St. Dev. Cum. Grade
%
Cum. Rank %
0 - 0.498 215 0.253 0.238 0.140 1.9 14.9
0.498 - 0.902 217 0.688 0.671 0.123 8.5 30.0
0.902 - 1.304 216 1.088 1.088 0.124 15.5 44.9
1.304 - 1.815 217 1.539 1.544 0.152 26.4 60.0
1.815 - 2.602 216 2.205 2.176 0.226 37.8 75.0
2.602 - 2.797 38 2.692 2.689 0.052 39.7 77.6
2.797 - 4.284 179 3.398 3.367 0.390 54.9 90.0
4.284 - 5.800 59 4.896 4.696 0.428 62.5 94.1
5.800- 7.520 35 6.661 6.502 0.413 69.6 96.5
7.520 - 9.820 18 8.665 8.783 0.918 75.5 97.8
9.820 - 16.329 18 12.059 11.579 1.960 84.9 99.0
>= 16.329 14 20.597 16.932 9.564 100.0 100.0

TABLE 14- 47 STATISTICS FOR CROSSCOURSE DEPOSIT E-LENS DOMAINS GOLD INDICATORS

Ordinary Kriging was used to estimate each of the grade indicators. The indicator estimates were postprocessed to ensure decreasing probabilities with increasing grade and to deal with any indicator estimates above one or below zero. The final e-type estimates were calculated by discretization of the probability distribution resulting from the Kriging of the indicators.

MIK was run with the same set of search parameters across all indicators, in order to avoid serious order relations issues. The estimation search parameters are detailed in Table 14-48. A constant minimum of four and maximum of 28 data points were used in each case.

Area Domain Search
Radius m
Azimuth Plunge Dip Major/Semi Major/Minor
E-Lens 100 100 340 70 0 1 3
200 120 340 70 0 1 3

TABLE 14-48 CROSSCOUSE DEPOSIT E-LENS DOMAINS MIK ESTIMATION SEARCH PARAMETERS

14.3.3.8  3D Ordinary Kriging Modeling Technique

In addition to the modeling techniques mentioned above, 3D Ordinary Block Kriging of equal length downhole composites in a traditional 3D block model was also completed on both mineral resource areas as a validation method.

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The 3D block estimate was based on interpolation into 25mN x 5mE x 25mRL parent cells, with sub celling to 6.25mN x 1.25mE x 6.25mRL to control volume.

A standard 3 dimensional, single pass Ordinary Kriging methodology has been used for the estimation of gold grade using the down hole composite data within each mineralized domain. Table 14-49 summarizes the estimation parameters by domain. A constant minimum of three and maximum of 32 data points were used for the Union Reefs West while a minimum of four and maximum of 28 data points were set for the E-Lens.

Area Domain Search
Radius m
Azimuth Dip Plunge Major/Semi Major/Minor
URW 1001 100 10 90 0 1 4
E-Lens 100 120 340 70 0 1 3
200 165 340 70 0 1 3

TABLE 14-49 CROSSCOURSE DOMAINS OK ESTIMATION PARAMETERS

14.3.3.9  Block Model Definition

A 2D block model was created for the URW lode grade estimation (domains 1001). Table 14-50 presents the 2D model definition parameters whilst a list of field names and descriptions included in the models is shown in Table 14-51.

  Minimum Maximum Model Extent
Easting 4,750 4,751 1
Northing 6,700 7,200 500
RL 600 1250 650
Parent Cell X m 1 Min Sub- Cell X m 1
Parent Cell Y m 25 Min Sub-Cell Y m 25
Parent Cell Z m 25 Min Sub-Cell Z m 25

TABLE 14-50 URW DOMAIN 2D PROJECTION BLOCK MODEL DEFINITION

Field Name Description
x X Block Centroid
y Y Block Centroid
z Z Block Centroid
au_ok_2d Back calculated Gold Grade
au_ok Au OK – downhole composites (comparison)
hw Ordinary kriged estimate of horizontal width
accum Ordinary kriged estimate of gold accumulation
ns Number of samples used for au_hw estimate (au_ok
sd Standard Deviation (au_ok estimation)
sr Slope of Regression (au_ok estimation)
density Assigned In Situ Bulk Density
zonecode Mineralized domain

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Field Name Description
rescat mineral resource classification
depletion Depletion flag
wx_code Oxidation type
geo Rock Type

TABLE 14-51 URW DOMAIN PROJECTION BLOCK MODEL ATTRIBUTES

The primary consideration of the 2D model was to provide an appropriate block size for the interval composite data spacing in long section view. Figure 14-30 is a long section view of the URW domain 1001 interpreted wireframe and composite data. The 25x25m grid demonstrates that within the best sampled parts of the domain, above RL 1050 a choice of 25x25m block size ensures the majority of blocks contain at least one composite data.

The 2D domain model was estimated, re-located into real world co-ordinates and imported into the 3D block model defined in Figure 14-30, below in the local grid coordinate system.

A 3D block model was created to represent the final grade and volume model for reporting of the URW lode. The grade estimate for the E-Lens domains (100 and 200) was interpolated directly into this model and the grade estimate for the URW lode was imported from the 2D model described above. The model definition is shown in Table 14-52 and a list of field names and descriptions for the model are presented in Table 14-53.

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  Minimum Maximum Model Extent
Easting 4,600 5,200 600
Northing 6,000 7,200 1,200
RL 600 1250 650
Parent Cell X m 5 Min Sub- Cell X m 1.25
Parent Cell Y m 25 Min Sub-Cell Y m 6.25
Parent Cell Z m 25 Min Sub-Cell Z m 6.25

TABLE 14- 52 CROSSCOURSE DEPOSIT E-LENS DOMAINS AND FINAL 3D BLOCK MODEL DEFINITION

Field Name Description
x X Block Centroid
y Y Block Centroid
z Z Block Centroid
au_final Final reportable Au grade
au_ok_2d 2d OK Au estimate
au_mik_etype MIK Method Au estimate
au_ok OK Au estimate
density Assigned In Situ Bulk Density
zonecode Mineralized domain
rescat mineral resource Classification
depletion Depletion flag
wx_code Oxidation type
ads Average distance to samples for au estimate
dns Distance of nearest sample for au estimate
ns Number of samples used for au estimate
sd Standard Deviation
sr Slope of Regression

TABLE 14-53 CROSSCOURSE DEPOSIT E-LENS DOMAINS AND FINAL 3D BLOCK MODEL ATTRIBUTES

The primary consideration of the 3D model design was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralized zone.

The chosen block size represents approximately half the best data spacing in the Northing direction and a choice in the vertical and easting dimension controlled by the need to appropriately represent the volume of the wireframes.

Table 14-54 below confirms the close agreement of the 3D block model volumes and the original interpreted wireframe volumes, supporting the 3D model block size choice.

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Domain Domain Wireframe Volume Block Model
Constraint
Volume
Difference
%
URW 1001 261,924 261,963 0.0%
E-Lens 100 1,746,303 1,743,701 -0.1%
200 779,030 819,434 5.2%

TABLE 14-54 CROSSCOURSE DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK

14.3.3.10  Bulk Density Assignment

The bulk density of the waste and mineralized rock of the final 3D block model has been assigned according to oxidation state, using the interpreted surface described in Section 14.3.3 to control the blocks assigned. Crocodile Gold determined the specific gravity based on historic reports of determinations made on drill samples. Table 14-55 summarizes the oxidation state specific gravity assignations.

Oxidation State SG
gm/cm3
1 Oxide 2.5
2 Transitional 2.6
3 Fresh (sulphide) 2.7

TABLE 14-55 CROSSCOURSE DEPOSIT BULK DENSITY DATA STATISTICS BY OXIDATION STATE

Within the final 3D block model blocks coded fresh below the top of fresh surface (tofr) were assigned a density of 2.7g/cm 3; those coded transitional between the base of complete oxidation surface (box) and the top of fresh surface, a density of 2.6g/cm 3; and those coded oxidized below the topographical surface, a density of 2.5g/cm 3.

Blocks located above the topographical surface were assigned a zero density. Where fill areas exist on the surface, such as waste dumps, dams, back-filled excavations and other surface excavations, these have been assigned a density of 1.8g/cm 3.

14.3.3.11  Model Depletion

Mining depletion as a result of open pit mining and other surface excavations has been coded into the block model. Two DXF files supplied were imported into Surpac, from which 2 DTM surfaces were created:

Within the final 3D block model blocks in air and mined out areas above the topography with open pit excavations were assigned a code of zero. Those areas between the fill areas (waste dumps, dams, and other surface excavations) and the original surface topography were coded as 99. Blocks locatedbelow the mined out open pit and original topographical surface were assigned a code of “1” (Table 14-56).

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Depletion Code Description
0 Above topographic surface & open pits
1 Below topographic surface & open pits
99 Surface fill areas (waste dumps, dams, other fill)

TABLE 14-56 DEPLETION CODES ASSIGNED ABOVE AND BELOW TOPOGRAPHIC SURFACES

14.3.3.12  Model Validation

Model validation has been undertaken to ensure no material error has been made in the estimation of the Crosscourse Lodes. The validations include inspection of the audit documentation of the individual estimation runs; visual inspection of the block outcomes and input data; statistical comparisons of input data and block outcomes, comparison of different estimation methods and swath plots. Swath plots show the estimated tonnes, estimated grade, number of composites and mean uncut composite grade (tabulated by northing and elevation), and were created for all the interpolated mineralization domains.

Statistical comparisons of input data and block model outcomes for the total URW Domain is shown in Table 14-57.

Area Domain Composite
Average Grade
g/t Au
Block Model
Average Grade
g/t Au
Variance %
URW 1001 3.19 3.75 118

TABLE 14-57 URW DOMAIN BACK CALCULATED AVERAGE GOLD GRADE COMPARISONS

The data comparison for URW displays a reasonable variation between input and outcome average grades when the total domain is reported. As can be confirmed in the visual inspection and swath plot investigations the comparison includes significant volumes at depth containing a lower density of sample data. This results in extrapolation of the sample data into these volumes and while it is considered a reasonable estimate of the grades in these volumes a simple statistical comparison of total volumes will not necessarily result in close comparisons. Swath plots of Domain 1001 in Figure 14-31 show that the domain is subject to a variable data density spatially.

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For that portion of the URW lode that contained sufficient sample density a comparison between the 2D accumulation gold estimate and an Ordinary Kriged 3D method estimate is shown in Figure 14-33.

The two estimation methods result in grades and tonnages that compare closely at the zero cut off. The 2D accumulation estimate clearly is less smoothed than the Ordinary Block Kriging, with differences in both tonnes and grade at all cut offs. For the purpose of identifying higher grade areas of the lode, the 2D accumulation model is preferred.

Statistical comparisons of input data and block model outcomes for E-Lens domains are shown in Table 14-58.

Area Domain Composite
Average
Grade g/t Au
Block Model
Average
Grade g/t
Au
Variance %
E-Lens 100 2.3 2.45 107
200 1.33 1.27 95

TABLE 14- 58 E-LENS DOMAINS AVERAGE GOLD GRADE COMPARISONS

The E-Lens lode comparisons display reasonable variation between composite and block model average grades for the total domains. For domain 100, 3D visual inspection and the swath plots display good comparisons between composite data and the block model grades where data density is greater. The high number of composites around 475m RL relate to close spaced RC drilling near the base of the Crosscourse open pit as noted in Section 14.4.3.2.

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For domain 200, the swath plots display moderate variation between composite and block model grades for the total domain reported. As with lode Domain 1001, a visual inspection and swath plot investigations show significant volumes at depth containing a lower density of sample data This results in extrapolation of the sample data into these volumes and while it is considered a reasonable estimate of the grades in these volumes a simple statistical comparison of total volumes will not result in close comparisons.

Swath plot validations for the E-Lens domains are presented below in Figure 14-33 and Figure 14-34.

The swath plots demonstrate that where there is regularly spaced data the block model reflects those data. The plots also highlight that a paucity of drilling data (in particular below an approximate RL of 1050m) results in parts of the model relying on only a few sampled data and are therefore subject to reduced certainty and increased risk.

Figure 14-35 and Figure 14-36 below compare the outcomes of Ordinary Block Kriging and MIK estimation methodology in the two E Lens domains. Both estimation methods predict the same grade, tonnes and metal above the 0.0 g/t Au cut off.

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Comparison within Domain 100 shows the two methods are very similar with the preferred MIK method resulting in marginally lower - low grade tonnes to a cut off of 2.5g/t Au and then slightly more tonnes above cut off above 2.5g/t Au. The resulting metal above curves is materially the same.

The comparison within Domain 200 is materially the same, with the MIK method resulting in slightly more tonnes at a marginally higher grade estimated above the 1.3g/t Au cut off. The resulting contained metal is marginally higher across most cut-offs for the MIK estimate.

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14.3.3.13  Mineral Resource Classification

The mineral resource has been classified in accordance with the NI43-101 guidelines. Assessment criteria include data integrity, drillhole spacing, sample locations, sampling density, and lode geometry, geological confidence and grade continuity. Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data.

The mineral resource has been classified as Indicated and Inferred; no Measured resource has been identified.

14.3.3.14  Data Spacing and Distribution

The Crosscourse model has been shown in validation to be subject to varying drill hole density and sample location in relation to the lodes geometry. In all lodes the drilling is regular and of sufficient density within the upper parts of lodes but subject to decreasing densities and irregular spacing at depth. The block model outcomes at depth in all lodes are considered to be higher risk and are classified with less confidence than the shallower parts. Each lode was considered individually and for lodes with sufficient data density a depth limit digitized for the base of Indicated and Inferred boundaries.

14.3.3.15  Orientation of Data in Relation to Geological Sructures

The orientation of the mineralized lodes is interpreted to be close to vertical and the drilling is considered to be appropriately targeted for this geological orientation.

14.3.6.16  Geological Interpretaiton

The geological interpretation of the Crosscourse deposit was undertaken by Crocodile Gold geologists.

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14.3.3.17  Deposit Dimensions

The mineralized portion of the Crosscourse deposit extends within drill testing from 6,425m to 7,075m in Northing with extension beyond drilling of up to 10m; Within the Easting plane the dimensions of the mineralization are tightly constrained by drilling as detailed in Section 14.4.2.4; in the vertical the deposit extends within drilling from surface (at approximately 1225mRL) to 700mRL, extension beyond drilling was up to 50m from the last data point. The dimensions of the mineralization are adequately defined by the available drilling with limited and acceptable extensions beyond data.

14.3.3.18  ESTIMATION AND MODELING TECHNIQUES

The estimation methodologies used for the two mineralization styles are considered appropriate based on experience with similar deposit types. Both are shown to represent reasonable unbiased reproductions of the input data in areas of adequate sampling. Outside areas of adequate sampling the mineral resource classification is such as to reflect the uncertainty of the estimate.

14.3.3.19  Moisture

The estimate has been made on the basis of dry tonnes.

14.3.3.20  Bulk Density

As detailed in Section 14.4.2.3 and Section 14.4.3.10 the bulk density factors used in this estimate are derived from historic records from previous mining work. No distinction has been made between mineralized material and waste rock. When sufficient sample data are available this distinction could be made to optimize material movement, however, in the context of a small tonnage rate underground exploitation method the assigned bulk density methodology is considered sufficient for classification. Variations due to new bulk density data are not expected to be material to the estimate.

14.3.3.21  Classification

All material within the mineral resource interpretation has been classified to represent the Author’s opinion of the risk in the mineral resource estimated. This respects the assumption that within the lode domains no selectivity will be able to be applied and so the total lode domain will be mined. Within the E-Lens domains that have been defined on a plus 0.4g/t cut off, it is assumed that this edge material as well as some internal material will form dilution to the mining of the mineralized material. Similarly, within the Union Reefs West domain that has been defined on a plus 0.7g/t Au cut-off, it is assumed that this edge will form dilution to the mining of the vein material.

Mineral resource blocks have been classified as Indicated or Inferred on the basis of drillhole spacing, sample locations, sampling density, wireframe geometry, geological confidence and grade continuity. A boundary string was generated to separate Indicated from Inferred blocks, on the basis of the above criteria.

For the E-Lens mineralized domains, the portion of the mineral resource classified as Indicated is defined by a substantial number of drill holes, in most areas at a spacing of 25m to 50m or less. The Inferred portion of the mineral resource largely represents the poorly drilled, strike extents and down-plunge extents of the domains.

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For the Union Reefs West lode, due to the variable nature of the drilling spacing there are blocks that are not directly informed by a composite resulting in a lower slope of regression. Indicated material is characterized by a high estimation quality and delineated by a slope of regression (true to estimated blocks) typically greater than 0.7. The Inferred limit of the mineral resource largely represents the poorly drilled areas of the interpreted lode, to drill spacing limits between 50m and 75m. Beyond the Inferred boundary limit for Union Reefs West, blocks within the mineralized domain have been coded as Unclassified. Figure 14-37 illustrates the classification limits of the model in relation to the data spacing within the interpreted lode. The Unclassified material is the down-dip and down plunge extension of the lode and including one diamond hole intersection approximately 450m below the surface.

The classification of the Crosscourse Lodes into Indicated and Inferred, as set out below reflects the Author’s view of this deposit, as it is currently defined.

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14.3.3.22  Selectivity Assumptions

Within the URW lode, the selectivity assumption is based on a two dimensional 25mx25m grid, and by lode width. This assumption requires the addition of edge dilution when a mining study is undertaken. The amount of dilution estimated will be dependent on the type of mining proposed.

For the E-Lens lodes, the selectivity assumption is based on the parent cell size (25mN x 5mE x 25mRL), as no change of support has been undertaken on this model.

14.3.3.23  Grade Tonnage Curves

Figure 14-38, Figure 14-39 and Figure 14-40 below detail the grade and tonnage curves for the URW Lode (Domain 1001), E Lens Lode (Domain 100 and 200) reported mineral resources respectively.

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14.3.3.24  Recommendations

The following points summarize the most relevant recommendations for the Crosscourse lodes:

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14.3.4      ORINICO DEPOSIT

14.3.4.1  Introduction

During December 2012, Cube Consulting Pty Ltd was requested by Crocodile Gold to undertake a mineral resource estimation update of the Orinoco deposit. The estimation incorporated a number of recently drilled infill holes as detailed in Section 10.

14.3.4.2  Data Type

The estimation of contained gold has been based on assays sourced from drilling data, detailed in Section 10, above. The data available consisted of, RC exploration and diamond core samples.

All data is in local grid co-ordinates.

Due to the difference in spatial coverage of the data types, all data types have been used in this estimate.

The total data base supplied consisted of 120 drill holes for a total of 10,203m, including 23 AO series RC grade control holes (for 1,332m), 14 URNRC series RC holes (for 1,003 meters) two diamond core holes ORD92006 & URD91008 (for 323m), 11 OR series holes (unknown drill type for 876m), 9 ORP series holes (unknown drill type for 996m), 3 PZ series holes (unknown drill type for 120m), 35 URP series holes (unknown drill type for 3,615m) and 23 WB & WNP series holes (unknown drill type for 1,938m).

Within the mineralized domains the drill data consisted of 121RC intercepts for 854 downhole meters. The drill cuttings and core have all been sampled and assayed on 1m lengths.

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14.3.4.3  Mineral resource Interpretation

Mineralization is associated with quartz-sulphide veining, compromising 1mm to 2m thick lode-style veins in sheared pelites, stockwork veins in greywacke and sheeted vein systems in thinly inter-bedded pelites and psammites.

Interpretation of mineralized domains have been informed by a gold cut-off grade, with a lower limit of approximately 0.4g/t used as the basis for defining mineralized material. A minimum down hole length of 3.0m (corresponding to approximately 2.0m horizontal width) has been applied during the interpretations.

The mineral resource mineralization interpretations were wireframed and nominated Domains 100, 101, 102, 201, 202, 301, 302, 401, 501, 601, 602, 701, 702, 801 and 802. Figure 14-41 and Figure 14-42, below shows the mineralization wireframes and drilling traces in plan and oblique views to the northwest.

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The mineral resource wireframes were used to code the drill intercepts contained within them by flagging into a new table in the database, the zonecode table. This flagging allows the selection of data within domains by codes for the purposes of sample analysis and compositing.

All mineral resource interpretation wireframes have been used as hard boundaries for this estimate.

The host lithologies of the Orinoco deposit have been logged systematically for varying states of weathering/oxidation from one totally oxidized to five un-oxidized. Logging of one and two is considered oxidized; three and four transitional and five un-oxidized. Base of complete oxidation (BOCO) and top of fresh rock (TOFR) surfaces interpreted from this geological logging have been used to flag oxidation state into the three dimensional block model. The Company has not used the interpreted oxidization surfaces as hard boundaries during the estimation of gold as the number of samples in the oxidized and transitional zones is too low for a meaningful comparison.

The estimation of the mineral resource domains was undertaken using Ordinary Kriging to interpolate grades into a 3D block model.

14.3.4.4  Compositing and Statistics

An examination of sample statistics for mineral resource domains reveals that all the sampling of the mineralization is on 1m sample lengths, consequently compositing the data for equal support is not necessary. A statistical summary of the mineralization domains is shown in Table 14-59, below.

The effect of a small number of outlier sample grades or spatially isolated samples may have an undue effect of the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the sample data in 3D visualization. A number of high cuts or limits were identified and applied as necessary within the domains detailed in Table 14-59 below.

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Domain No. of
samples
Min Gold
Grade
g/t Au
Max Gold
Grade
g/t Au
Mean Gold
Grade
g/t Au
High
Grade Cut
g/t Au
Mean Cut
Gold
Grade
g/t Au
Cut CV
100 49 0.04 10.8 1.02 5 0.90 1.1
101 50 0.001 35.2 2.04 8 1.18 1.8
102 25 0.001 19.6 1.31 5 0.73 1.5
201 91 0.005 11.6 1.01 8 0.97 1.2
202 312 0.005 10.3 0.88 10 0.88 1.2
301 148 0.005 22.1 1.00 8 0.90 1.5
302 48 0.02 4 0.70 - 0.70 1.1
401 28 0.005 9.76 1.18 5 1.01 1.1
501 18 0.005 13.3 2.06 8 1.74 1.5
601 16 0.005 36.5 4.41 5 1.32 1.3
602 6 0.48 4.02 1.28 - 1.28 1.0
701 29 0.001 92.8 7.37 10 2.42 1.5
702 4 1.56 4.01 2.60 - 2.60 0.4
801 17 0.03 4.09 1.09 - 1.09 1.1
802 13 0.01 25.7 2.84 5 1.25 1.1

TABLE 14- 59 ORINOCO DEPOSIT - HIGH GRADE SAMPLE CUTS BY DOMAIN

The general statistics of gold samples within all domains can be described as positively skewed with moderate to high variability as is the case with most gold occurrence. The high variability is reduced somewhat by high cutting of gold grades in those domains most affected.

14.3.4.5  Variography

Variography was used to characterize the spatial behavior of the sample data primarily, as an aid to establishing estimation parameters. Variogram stability and quality is dependent on the statistical properties of defined domains and the amount of data available within domains. After an initial investigation, one model was established, for the most informed domain, 202. The other mineralization domains contained either insufficient numbers of sample data or poorly structured variograms. The variogram parameters established for domain 202 were adopted for the remaining mineralization domains, with some adjustments to the search orientation.

The final variogram model parameters are detailed in Table 14-60, below.

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Domain Nugget Sill
1
Range
1
Sill
2
Range
2
Azi Plunge Dip Major
/Semi
Major
/Minor
100 0.22 0.39 30m 0.39 120m 030 0 85 1 5
101 0.22 0.39 30m 0.39 120m 030 0 85 1 5
102 0.22 0.39 30m 0.39 120m 010 0 80 1 5
201 0.22 0.39 30m 0.39 120m 010 0 80 1 5
202 0.22 0.39 30m 0.39 120m 030 0 85 1 5
301 0.22 0.39 30m 0.39 120m 030 0 85 1 5
302 0.22 0.39 30m 0.39 120m 010 0 80 1 5
401 0.22 0.39 30m 0.39 120m 030 0 85 1 5
501 0.22 0.39 30m 0.39 120m 030 0 85 1 5
601 0.22 0.39 30m 0.39 120m 020 0 80 1 5
602 0.22 0.39 30m 0.39 120m 020 0 80 1 5
701 0.22 0.39 30m 0.39 120m 020 0 80 1 5
702 0.22 0.39 30m 0.39 120m 020 0 80 1 5
801 0.22 0.39 30m 0.39 120m 020 0 80 1 5
802 0.22 0.39 30m 0.39 120m 020 0 80 1 5

TABLE 14- 60 ORINOCO DEPOSIT FINAL VARIOGRAM MODELS BY DOMAIN

14.3.4.6  Grade Interpolation Methodology

A standard three dimensional single or two pass Ordinary Kriging methodology has been used for the estimation of the cut gold 1m down hole sample data within each mineralization domain. Table 14-61 summarizes the estimation parameters by domain. A constant minimum of three and maximum of 10 data have been set and a discretization of 1 in X 5 in Y and 1 in Z has been used throughout.

Domain Search
Radius
m
Pass2
Radius
m
Azimuth Plunge Dip Major/Semi Major/Minor
100 30 120 030 0 85 1 2
101 30 120 030 0 85 1 2
102 30 120 010 0 80 1 2
201 30 120 010 0 80 1 2
202 30 120 030 0 85 1 2
301 30 120 030 0 85 1 2
302 30 120 010 0 80 1 2
401 30 120 030 0 85 1 2
501 30 120 030 0 85 1 2
601 30 120 020 0 80 1 2
602 30 120 020 0 80 1 2
701 30 120 020 0 80 1 2
702 30 120 020 0 80 1 2
801 30 120 020 0 80 1 2
802 30 120 020 0 80 1 2

TABLE 14-61 ORINOCO DEPOSIT MINERALIZATION DOMAIN ESTIMATION PARAMETERS

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14.3.4.7  Block Model Definition

The primary consideration of the 3D model was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralized zone. Table 14-62 summarizes the 3D block model “orinoco_dec_2012.mdl” definition.

  Northing Easting RL
Minimum 8900 4700 950
Maximum 9800 5200 1200
Block Size m 10 2.5 2.5
Sub-block m 2.5 0.625 0.625

TABLE 14-62 ORINOCO DEPOSIT FINAL 3D BLOCK MODEL DEFINITION

The chosen block size represents approximately half to one quarter of the best data spacing in the Northing direction and a choice in the vertical and easting dimension controlled by the need to appropriately represent the volume of the wireframes.

A standard list of field names and descriptions used in the block model are shown in Table 14-63.

Attribute Type Default Description
au_ok_final Float -99 estimated gold by Ordinary Kriging (ppm)
au_ok_test1 Float -99 test estimation; 1 pass, min 3, max 10 (ppm)
au_ok_test2 Float -99 test estimation; 1 pass, min 3, max 30 (ppm)
au_id Float -99 estimated gold by inverse dist. 2 (ppm)
domain Integer -99 Mineralization code
oxcode Integer 0 0 = air; 1 = Fresh; 2 = Trans; 3 = oxidized
density Float 0 Bulk Density
classification integer -1 Ind = 2; Inf = 3, unclass = 4

TABLE 14-63 ORINOCO DEPOSIT FINAL 3D BLOCK MODEL ATTRIBUTES

Table 14-64 below confirms the close agreement of the 3D block model volumes and the original interpreted wireframe volumes, supporting the 3D model block size choice as appropriate.

Domain Wireframe
Volume
Block Model
Constraint
Volume
Difference
%
100 19,806 19,831 0%
101 38,190 38,349 0%
102 14,327 14,223 1%
201 43,408 43,361 0%
202 145,149 145,218 0%
301 64,338 64,321 0%
302 17,138 17,056 0%
401 16,335 16,412 0%
501 11,542 11,533 0%

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Domain Wireframe
Volume
Block Model
Constraint
Volume
Difference
%
601 11,437 11,359 1%
602 4,570 4,267 7%
701 13,482 13,518 0%
702 1,989 1,990 0%
801 9,951 9,800 2%
802 5,413 5,428 0%
TOTAL 417,075 416,666 0%

TABLE 14-64 ORINOCO DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK

14.3.4.8  Oxidation and Bulk Density Assignment

The bulk density of the waste and mineralized rock of the 3D block model has been assigned according to oxidation state, using interpreted surfaces described in Section 14.4.3. A total of 89 bulk density measurements were available in the supplied database, the majority of which were from the fresh oxidation zone. There is some risk of over stating the bulk density values (particularly with oxide and transitional zones) as only intact rock is generally used for the determinations by water immersion. The available data was insufficient to determine appropriate values for the oxide and transitional oxidation zones, so nominal values were assigned. Table 14-65 summarizes the oxidation state bulk density data statistics and assignment into the block model.

Oxidation
State
Number Min BD
gm/cm3
Max BD
gm/cm3
Mean BD
gm/cm3
Median BD
gm/cm3
STD Assigned BD value
in Model gm/cm3
Oxide 19 2.46 2.78 2.70 2.72 0.072 2.0
Transitional 17 2.52 2.78 2.70 2.72 0.069 2.5
Fresh 53 2.53 2.86 2.74 2.75 0.073 2.7

TABLE 14-65 ORINOCO DEPOSIT BULK DENSITY (BD) DATA STATISTICS BY OXIDATION STATE

Blocks located above the topographical surface were assigned a zero density.

14.3.4.9  Model Validation

Model validation has been undertaken to ensure no material error has been made in the estimation of the Orinoco deposit. The validations include inspection of the audit documentation of the individual estimation runs; visual inspection of the block outcomes and input data; statistical comparisons of input data and block outcomes, and swath plots of the most significant domains.

Statistical comparisons of input data and block model outcomes for each mineralized domain are shown in Table 14-66. Although these two items (Kriged values and mean values) are not strictly comparable due to data clustering and volume influences they provide a useful validation tool in detecting any major biases. Overall the grade estimates compare well with the sample means, with the larger deviations being the result of data cluster.

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Domain Number of
Samples
Sample
Average
Grade
g/t Au
Block Model
Average Grade
g/t Au
Variance
%
100 49 0.90 0.79 114
101 50 1.18 1.48 80
102 25 0.73 0.87 84
201 91 0.97 0.98 99
202 312 0.88 0.87 101
301 148 0.90 0.98 92
302 48 0.70 0.66 107
401 28 1.01 0.99 102
501 18 1.74 1.65 106
601 16 1.32 1.09 121
602 6 1.28 1.21 106
701 29 2.42 1.92 126
702 4 2.60 2.43 107
801 17 1.09 1.23 88
802 13 1.25 1.23 101
TOTAL 854 1.00 1.03 97

TABLE 14- 66 ORINOCO DEPOSIT MINERALIZED DOMAIN AVERAGE GOLD GRADE COMPARISONS

As can be confirmed in the visual inspection and swath plot investigations the comparisons include significant volumes at depth in each domain containing a lower density of sample data. This results in extrapolation of the sample data into these volumes and while it is considered a reasonable estimate of the grades in these volumes a simple statistical comparison of total volumes will not result in close comparisons.

Swath plot validations for the best informed domains (101, 201, 202, & 301) are presented in Figure 14-43 to Figure 14-46 below. Data is analyzed by northing and by elevation for each domain.

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The swath plots demonstrate that where there is regularly spaced data the block model reflects those data. The plots also highlight that a paucity of drilling data (in particular below an approximate RL of 1080m) results in parts of the model that rely on only a few measured points and are therefore less likely to match local composite data, and are of reduced certainty and increased risk.

Another validation step included estimating the gold grade by different methods to examine the potential sensitivity of the data set to differing input parameters and estimation technique. Variations in estimation included the following;

A summary of the results for the different estimations are illustrated by tonnage and grade curves in Figure 14-47 below. The parameters adopted for the December 2012 Orinoco mineral resource estimation are shown in green as "Final Grade" and "Final Tonnes".

14.3.4.10  Mineral Resource Classification

The classification of the Orinoco mineral resources was based on information provided by Newmarket Gold and outcomes of the estimation processes undertaken by Cube. The mineral resource has been classified in accordance with the NI43-101 guidelines. Assessment criteria include data integrity, drillhole spacing, sample locations, sampling density, and lode geometry, geological confidence and grade continuity. Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data.

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The mineral resource has been classified as Indicated and Inferred; no Measured mineral resource has been identified.

14.3.4.11  Data Spacing and Distribution

The Orinoco model has been shown in validation to be subject to varying drillhole density and sample location in relation to the lode geometry. In most lodes the drilling is regular and of sufficient density within the upper parts of lodes but subject to decreasing densities and irregular spacing at depth. The block model outcomes at depth in most lodes are considered to be higher risk and are classified with less confidence than the shallower parts. Each lode was considered individually and for lodes with sufficient data density, the following criteria of less than 40x20m to 40x40m data spacing where continuity is demonstrated over several sections, was utilized for an Indicated mineral resource classification.

14.3.4.12  Orientation of Data in Relation to Geological Structure

The orientation of the deposit is interpreted to be close to vertical and the drilling is considered to be appropriately targeted for this geological orientation.

14.3.4.13  Estimation and Modeling Techniques

The estimation methodology used is considered appropriate by the CP based on experience with similar deposit types. The methodology is shown to represent reasonable unbiased reproduction of the input data in areas of adequate sampling. Outside areas of adequate sampling the mineral resource classification is such as to reflect the uncertainty of the estimate.

14.3.4.14  Moisture

The estimate has been made on the basis of dry tonnes.

14.3.4.15  Classification

All material within the mineral resource interpretation has been classified to represent the Authors opinion of the risk in the mineral resource estimated. For reporting purposes the mineralized material has been reported with a lower cut-off of 0.5g/t gold within the interpreted wireframes. The classification of the Orinoco deposit into Indicated and Inferred as set out below reflects the interpreted view of this deposit as it is currently defined.

14.3.4.16  Recommendations

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14.3.5      ESMERALDA DEPOSIT

The Esmeralda deposit has been calculated into two mineable domains, the first is the open pitable mineral resource, which is 28,300 Indicated mineral resource ounces. The second is the underground potential below these optimized shells, which adds a further 9,000 Indicated mineral resource ounces of gold. Table 14-67 and Table 14-68.

Esmeralda deposit Mineralized Domains (Au>0.5 g/t)
Domain Tonnes Gold Grade g/t Oz Gold
Indicated 461,000 1.91 28,300
Inferred 53,000 2.08 3,500

TABLE 14-67 ESMERALDA OPEN PIT MINERAL RESOURCE ESTIMATION

Notes on Table 14-67:

  1.

Mineral resources are stated as of the December 31, 2015.

  2.

Mineral resources are inclusive of mineral reserves, which are set out below.

  3.

Mineral resources are calculated using these parameters;

  a.

Gold price of $A1,500/oz, metallurgical recovery of 90.0%

  b.

Lower cut-off of 0.5g/t Au is used to calculate the mineral resources

  c.

mineral resources table above have been optimized using Minemap™ software using parameters as set out earlier

  4.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces.

  5.

The mineral resource estimate was prepared by Mark Edwards, B.SC. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold .

  6.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.


Esmeralda deposit Mineralized Domains (Au>2.0 g/t)
Domain Tonnes Gold Grade g/t Oz Gold
Indicated 97,000 2.88 9,000
Inferred 89,000 2.91 8,300

TABLE 14-68 ESMERALDA UNDERGROUND MINERAL RESOURCE ESTIMATION

Notes on Table 14-68:

  1.

Mineral resources are stated as of the December 31, 2015.

  2.

Mineral resources are inclusive of mineral reserves, which are set out below.

  3.

Mineral resources are calculated using these parameters;

  a.

Gold price of $A1,500/oz, metallurgical recovery of 90.0%

  b.

Lower cut-off of 2.0g/t Au is used to calculate the mineral resources

  c.

mineral resources outlined in tables above are located directly below the optimized shell open pitable mineral resources outlined above in Table 14-67

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  4.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces

  5.

The mineral resource estimate was prepared by Mark Edwards, B.SC. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold .

  6.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.

14.3.5.1  Introduction

During January 2016 Newmarket Gold undertook a mineral resource estimation update of the Esmeralda deposit. The mineral resource update included the majority of historic RC holes and incorporated the 2015 drilling reported in Section 10. The majority of the 2015 drilling was RC, however, 8 diamond hole were also drilled primarily for geotechnical purposes.

The Esmeralda deposit consist of two zones of mineralization referred to as Zone A and Zone B. Zone A is the southern larger deposit (7500mN-9500mN) with Zone B a smaller northern area (9000mN – 10000mN)

Figure 14-48 shows a plan view of the drilling used in the Esmeralda model update, with historic RC drilling shown in green, 2015 RC drilling in pink and 2015 diamond drilling show in in blue.

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Esmeralda deposit Mineralized Domains Combined
Domain Tonnes Gold Grade g/t Oz Gold
Indicated 495,000 2.03 32,300
Inferred 80,000 2.75 7,000

TABLE 14-69 COMBINED MINERAL RESOURCE FOR ESMERALDA PROSPECT (OPEN PIT AND UNDERGROUND)

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14.3.5.2  Data Types

The estimation of contained gold has been based on assays sourced from drilling data. The data available mainly consists of RC chip assay results and a small number of diamond core sample assay results.

All data is in local grid coordinates.

Drilling provides data along the strike of both zones of mineralization to a depth of approximately 160m below surface. The resource update included 234 holes totaling 17,214.6m of drilling summarized in Table 14-70.

Hole Type # of Holes Meters Drilled Average Depth
RC 222 811.3 206.3
DDH 12 16403.3 213.7

TABLE 14- 70 ESMERALDA PROSPECT DRILLHOLE SUMMARY BY TYPE

Within the Esmeralda deposit there were 174 mineralized intersections totaling 896m downhole meters for all drill types summarized in Table 14-71. There are multiple instances of grade variability in all drill types, the nature of the variability appears to be related to the deposit and is not specific to a particular drill type.

Hole
Type
# of
Intersections
Minimum
Length
(m)
Maximum
Length
(m)
RC 163 1 20
DDH 11 0.62 14
TOTAL 174 0.62 20

TABLE 14-71 ESMERALDA PROSPECT SUMMARY OF SAMPLE LENGTHS BY HOLE TYPE FOR ESMERALDA MINERALIZED DOMAINS

All data types were included in the mineral resource evaluation to increase data density. The diamond drill holes have only a limited effect on defining the mineralization within the wireframes, with the seven recent diamond holes targeting specific regions for geotechnical purposes rather than defining mineralization. The level of confidence in the data is reflected in the mineral resource classification in Section 14.3.5.10.

14.3.5.3  mineral resource Interpretation

Mineralized domain interpretation was informed by gold grades with a lower grade cut-off of 0.5g/t Au for both Zone A and Zone B utilized for near surface mineralization. A lower cut-off grade of 2.0g/t Au was utilized for deeper mineralization. No minimum length criterion was applied. The mineralized domains were wireframed and nominated domain codes of 1-6 for Zone A and 11-13 for Zone B. The mineralized domains were used as hard boundaries for the estimation of grade within both zones.

The mineral resource wireframes were used to code the drill intercepts contained within each domain, flagging the intervals in the database. The flagging allowed the selection of data within the load domain for compositing and sample analyses.

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14.3.5.4  Compositing and Statistics

Within the Esmeralda mineral resource area, 16,888 samples were utilized in defining the mineral resource domains, with the average length of the samples being 1.01m.

Sample statistics for the Esmeralda deposit mineral resource domains reveal that the majority of the sampling contained within the domains is on 1m downhole lengths summarized in Table 14-72.

Domain # of
samples
Minimum
Length
(m)
Maximum
Length
(m)
Average
Length
(m)
1 305 0.7 1.4 1
2 113 1 2 1.01
3 29 1 1 1
4 75 1 1 1
5 49 1 1 1
6 4 1 2 1.25
11 202 0.62 1.1 1
12 89 1 1 1
13 30 1 1 1
Overall 896 0.62 2 1.00

TABLE 14-72 ESMERALDA PROSPECT MINERALIZED DOMAIN SAMPLE LENGTH STATISTICS

The raw assays were composited to 1m intervals, with intervals less than 0.5m combined in to the previous composite. Table 14-73 summarizes the composite statistics for the Esmeralda deposit

Domain Number Minimum
g/t Au
Maximum
g/t Au
Mean
g/t Au
Standard
Deviation
Co-efficient of
Variation
1 305 0.00 52.70 2.06 3.80 1.84
2 114 0.00 53.90 2.72 5.80 2.14
3 29 0.01 7.30 1.43 1.88 1.31
4 75 0.00 9.65 2.41 2.30 0.95
5 49 0.00 9.80 1.84 2.18 1.18
6 5 0.56 9.94 4.89 4.07 0.83
11 197 0.00 11.40 1.29 1.23 0.96
12 83 0.11 16.70 1.87 2.06 1.11
13 30 0.00 6.70 1.33 1.76 1.33

TABLE 14- 73 ESMERALDA PROSPECT COMPOSITE STATISTICS BY DOMAIN

Statistical evaluation of the composite distribution of gold grades within the mineralized domains highlighted of the presence of outlier gold grades. The upper limit of the idealised normal distribution of assays was determined by visual assessment of a histogram plot of gold grades such as Figure 14-50 below, which shows the outlier gold grades to the right of the histogram. Table 14-74 shows the resulting statistics the outlier restrictions on the co-efficient of variation and the percentage of metal removed.

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Domain Upper limit
Value
(g/t
Au)
Number of
Sample

Cut
Cut Mean
(g/t Au)
Cut Co-
efficient of
Variation
%Metal Cut %Data Cut
1 13 6 1.87 1.07 14.75% 2.11%
2 10 7 2.11 0.92 38.66% 7.29%
11 10 1 1.29 0.77 4.53% 0.54%
12 10 1 1.72 0.73 10.83% 2.56%

TABLE 14-74 ESMERALDA PROSPECT HIGH GRADE RESTIRCTIONS BY DOMAIN

The outlier results were not removed from the composites used in the model estimation, but their range of influence was restricted to the immediate vicinity (10m), before being capped to the upper limit value.

Within the domains there is a relatively consistent co-efficient of variation across all domains, particularly after the outlier restriction. This is a measure of the consistency of the variability across both zones and all domains, and highlights the similar population of gold grades across all domains.

14.3.5.5  Variography

Variography was used to characterize the spatial behavior of the composite data as an aid to establishing estimation parameters. Each domain was assessed with six domains (1, 2, 4, 5, 11 and 12) showing statistical valid variogram models. For each domain, the variogram search plane was set along the general strike of the domain, with analyses completed in the primary (along strike), secondary (down dip) and tertiary (across strike) directions. The three directions were then combined to create a 3D variogram model for each individual domain that was utilized in the model estimation. The tertiary search direction across the domain representing the closest spaced data, representative of downhole direction, was used to assess the nugget of each domain. Each domain was modeled as a single spherical structure with varying nugget values of typically around 50%. The variogram information is summarized in Table 14-75.

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Domain Azimuth Plunge Dip Nugget Sill Range
Primary
(m)
Range
Secondary
(m)
Range
Tertiary
(m)
1 351 0 70 1.98 2.2 40 40 2.5
2 190 0 86 1.34 2.46 75 75 3
4 8 0 64 1.1 4.7 30 30 2.4
5 3 0 83 1.98 2.2 45 30 2.5
11 10 0 72 0.55 0.45 40 40 3
12 11 0 71 0.2 1.35 45 17 1.9

TABLE 14-75 ESMERALDA PROSPECT VARIOGRAM MODELS BY DOMAIN

The variogram models are similar with slight variations in sills driven by the statistical grade variance in each domain. The relatively high nugget is observable in the drilling, with high variation in grades downhole. Each domain has a roughly isotropic search ellipse with no observable plunge in the grade distribution.

For the remaining domains (3, 6 and 13) no significant statistical analyses could be completed due to a lack of data. The average grade of the composites contained in these domains was utilized to populate mineralized blocks intersected by these domains. These domains are classified as Inferred mineral resources as discussed in Section 14.3.5.10.

14.3.5.6  Grade Interpolation Methodology

3D Ordinary Block Kriging of equal length downhole composites in a traditional 3D block model was completed on mineral resource domains. The 3D block estimate was based on interpolation into 10mN x 2.5mE x 5mRL parent cells, with sub celling to 2mN x 0.5mE x 1mRL to control volume.

A standard 3 dimensional, single pass Ordinary Kriging methodology has been used for the estimation of gold grade using the downhole composite data within each mineralized domain. Table 14-76 summarizes the estimation parameters by domain. A constant minimum of one and maximum of 30 data points were used for the majority of the Esmeralda deposit Domains except domain 11 where a more localized estimation was desired due to local grade variability.

   Search Distance Rotation Sample Selection number
definitions

Domain Primary Secondary Tertiary Major Minor Vertical Min Max Max per
hole
1 135 90 7.5 351 0 70  1 30 5
2 150 150 6 12 0 82  1 30 5
4 60 60 5 8 0 64  1 30 5
5 90 60 10 3 0 83  1 30 5
11 135 90 20 7.5 0 72  1 15 5
12 200 100 20 11 0 71  1 30 5

TABLE 14-76 ESMERALDA DEPOST ORDINARY KRIGING ESTIMATION PARAMETERS

A large search parameter at the same anisotropic ratio as the varigraphy ranges was utilized for estimation. This was intended to allow variogram models to select and weight samples in order to estimate block grades as a true statistical representation.

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14.3.5.7  Block Model Definition

  Minimum Maximum Model Extent
Easting 4,500 6,000 2.5
Northing 7,900 10,000 10
RL 50 300 5
Parent Cell
X m
2.5 Min Sub- Cell
X m
0.5
Parent Cell
Y m
10 Min Sub- Cell
Y m
2
Parent Cell
Z m
5 Min Sub- Cell
Z m
1

TABLE 14-77 BLOCK SIZE AND MODEL DIMENSIONS

Field
Name
Type Default Description
MODEL_X Float   Block Co-ordinate Easting
MODEL_Y Float   Block Co-Ordinate Northing
MODEL_Z Float   Block Co-ordinate RL
DX Float   Block Dimension East
DY Float   Block Dimension North
DZ Float   Block Dimension RL
AU Float -1 Ordinary Kriging Gold Grade g/t Au
BULKD Float -1 Assigned In Situ Bulk Density
TOPO integer -1 Topography
-1=Air 1=Below Surface
MATIL integer -1 Material Type
1=Fresh 2=Transitional 3=Oxide
RSCAT integer -1 mineral resource Classification
1=Measured 2= Indicated 3= Inferred

TABLE 14-78 ESMERALDA MODEL CODING

The primary consideration of the 3D model design was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralized zone.

The chosen block size represents approximately a third of the best data spacing in the Northing direction. In the vertical and Easting, dimensions size was controlled by geostatistical variance considerations and the need to appropriately represent the volume of the wireframes.

Table 14-79 below confirms the close agreement of the 3D block model volumes and the original interpreted wireframe volumes, supporting the 3D model block size choice.

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  Volume (m3)
Lode Wireframe Wireframe
cut to
surface
Mineral
resource
% Change
between cut
wireframe &
Mineral
resource
1 214,337 198,745 198,430 0%
2 76,427 71,575 69,640 -3%
3 34,801 30,244 30,016 -1%
4 41,262 38,745 38,370 -1%
5 33,831 30,765 29,597 -4%
6 1,545 1,545 1,433 -8%
11 190,672 181,714 181,367 0%
12 46,953 42,807 42,469 -1%
13 29,758 28,614 27,627 -4%

TABLE 14- 79 ESMERALDA PROSPECT 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK

14.3.5.8  Oxidation and Bulk Density Assignment

Depth of oxidation at the Esmeralda deposit varies across the region. The deposit has a slight increase in the depth of oxidation levels to the north, with local variation driven by lithology and topographical changes.

The host lithologies of the Esmeralda deposit have been logged systematically for varying states of weathering/oxidation from one being totally oxidized to five being un-oxidized. Logging of one and two is considered oxidized; three and four transitional and five un-oxidized. Solids interpreted from this geological logging have been used to flag oxidation state into the three dimensional block model.

The bulk density of the waste and mineralized rock of the final 3D block model has been assigned according to oxidation state, using the interpreted solids described above to control the blocks assigned. Bulk Density factors used in this estimate are derived from historic records from previous mineral resource estimations, these have been reviewed by the Author and are deemed suitable for use in this estimate. No distinction has been made between mineralized material and waste rock. Table 14-80 summarizes the oxidation state bulk density assignations.

Oxidation
State
Bulk Density
g/cm3
1 Oxide 2.32
2 Transitional 2.56
3 Fresh 2.69

TABLE 14-80 ESMERALDA PROSPECT BULK DENSITY DATA STATISTICS BY OXIDISATION STATE

Within the final 3D block model blocks coded fresh below the top of fresh surface were assigned a density of 2.69g/cm 3; those coded transitional between the base of complete oxidation surface (box) and the top of fresh surface, a density of 2.56g/cm 3; and those coded oxidized below the topographical surface, a density of 2.32g/cm 3.

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14.3.5.9  Model Validation

Model validation has been undertaken to ensure no material error has been made in the estimation of Esmeralda deposit. The validations include inspection of singular block estimation parameters; visual inspection of the block outcomes and input data; statistical comparisons of input data and block outcomes.

Singular block Kriging estimations were completed for a number of blocks within each domain. This allowed the assessment of composites selected and weights applied to all samples influencing the grade estimation of the block, as well as a visual representation of the spatial location of samples being incorporated in the estimation.

Statistical comparisons of input data and block model outcomes for the all domains are shown in Table 14-81.


Domain
Composite
Average Grade
g/t Au
Block Model
Average Grade
g/t Au

Variance %
1 2.06 2.00 3%
2 2.72 1.89 30%
3 1.43 1.43 0%
4 2.41 2.07 14%
5 1.84 1.89 -2%
6 4.89 4.89 0%
11 1.29 1.34 -4%
12 1.87 1.76 6%
13 1.33 1.33 0%

TABLE 14-81 ESMERALDA PROSPECT DOMAIN AVERAGE GOLD GRADE COMPARISON

The domain comparison shows minor variation across most domains. The significant variance in Domain 2 can be attributed to the range limiting of the outlier data as discussed in Section 14.3.5.4. Several very high-grade results fall within Domain 2, which is enough to elevate the composite average statistics. With the spatial restriction of the high-grades in the model estimation, the average block grades fall towards the average of the idealized normal destruction of the population, while allowing the high-grade assays to influence the local area. Domain 4 variance can be attributed to clustered data within the domain, the high-grade composite data clustered to the south of the domain. During estimation the amount of blocks populated from the clustered data is less than the lower grade sparse data to the north, resulting in a lower average block grade compared to the composite average. The composite average does not represent the true spatial population average of the domain. This can be seen visually in Figure 14-51 below.

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14.3.5.10  mineral resource Classification

The classification of the Esmeralda deposit mineral resources was based on outcomes of the estimation processes. The mineral resource has been classified in accordance with the NI43-101 guidelines. Assessment criteria include data integrity, drillhole spacing, sample locations, sampling density, and lode geometry, geological confidence and grade continuity. Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data. A boundary string was generated to separate Indicated from Inferred blocks, on the basis of the above criteria with domains populated by average grades of composites classified as inferred only.

All material within the resource interpretation has been classified to represent the QP’s opinion of the risk in the resource estimated. Cut-off grades have been determined using estimated costs and gold process as outlined above. Domains have been defined on a plus 0.5g/t Au cut-off, it is assumed that this edge material as well as some internal material will form dilution to the mining of the mineralized material.

The mineral resource has been classified as Indicated and Inferred mineral resources. No Measured mineral resources have been identified.

14.3.5.11  Data Spacing and Distribution

The Esmeralda deposit has varying drill hole density and sample locations in relation to the load geometries. During the modeling process the distance to the nearest composite to each block was calculated and utilized during mineral resource classification. Summary statistics of the distances of composites used in the mineral resource modeling are summarized in Table 14-82.

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Domain Nearest
Composite
used (m)
Furthest Composite
used (m)
Average distance of
composites used (m)
1 20.16 82.72 57.53
2 44.31 120.62 81.08
4 16.47 56.30 36.92
5 20.17 82.56 51.71
11 22.92 64.82 44.97
12 18.47 87.68 54.90
All 21.17 75.04 51.89

TABLE 14-82 ESMERALDA PROSPECT DISTANCES TO COMPOSITES USED IN THE MINERAL RESOURCE MODEL

14.3.5.12  Orientation of Data in Relation to Geological Structure

The orientation of the mineralized lodes are generally steep westerly dipping to sub-vertical and strike roughly north - south in direction. The majority of the drilling used in the model estimation is of an east-west direction dipping between -90 to -45 degrees averaging -60 degrees. Drilling is considered to be appropriately targeted for the geological orientations of the Esmeralda deposit.

14.3.5.13  Estimation and Modeling techniques

The estimation methodologies used are considered appropriate based on experience with similar deposit types. The estimation is shown to represent reasonable unbiased reproductions of the input data in areas of adequate sampling. Outside areas of adequate sampling the mineral resource classification is such as to reflect the uncertainty of the estimate.

14.3.5.14  Moisture

The estimate has been made on the basis of dry tonnes.

14.3.5.15  Recommendations

The following points summarize the most relevant recommendations for the Esmeralda lodes:

Undertake infill drilling to further test peripheral inferred domains of narrow but high gold grades to extend and convert mineral resources.

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14.3.6      MINERAL RESOURCES GENERATED PRIOR TO NEWMARKET GOLD

Crocodile Gold previously reported on several other mineral resources in the Union Reefs area. These deposits will be reported without change in this update as no new models have been completed. The mineral resource estimations for Union North, Union Southm Millar/Big Tree and Lady Alice were completed prior to Crocodile/Newmarket Gold’s ownership of these deposits. Each mineral resource has been reviewed by the Author and have been deemed as appropriate for reporting and inclusion in this report.

All block models were generated using MineMap™ software, the same software was used to optimize using the Learch Grossman approach as mentioned above. The block model parameters are outlined in Table 14-83 below.

  Model Base Point and Parameters Algorithm  
Pit/ deposit Easting Northing Top Seam Rotation
degree

 
 
Union North 4900 7670 1210 6 ID  
Union South 4900 5240 1220 3.5 ID  
Millar/Big Tree 5040 4950 1220 2.5 ID  
Lady Alice 5100 7410 1240 5 ID  

  Cell Size Number of Cells
Pit/ deposit X Y Z X Y Z
Union North 2.5 10 2.5 130 90 57
Union South 2.5 10 2.5 80 90 65
Millar/Big Tree 2.5 10 2.5 85 65 65
Lady Alice 2.5 10 2.5 75 60 73

TABLE 14-83 UNION REEFS DEPOSITS - BLOCK MODEL SET UP PARAMETERS

Estimation Parameters and Methodology

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Rotation about axis 1st pass

Rotation about axis 2nd
pass

 
Pit/ deposit X Y Z X Y Z
Union North 7 90 20      
Union South 2.5 90 25      
Millar/Big Tree 2.5 90 30      
Lady Alice 5 90 20 5 90 60

  Search 1st pass Search 2nd pass
Pit/ deposit X Y Z X Y Z
Union North 40 60 6 20 30 6
Union South 40 50 6 20 30 6
Millar/Big Tree 40 50 6 20 30 6
Lady Alice 40 60 10 20 30 10

TABLE 14-84 UNION REEFS MODEL PARAMETERS

Model Validation

Model validation was performed by using historical mining data and reconciling the in-pit mined tonnes and grade.

Overall, ore mined produced 120% of the tonnes at 78% of the grade to produce 94% of the ounces compared to the mineral resource model. This, in part, could be explained by the net effect of dilution (dilution and ore loss).

In all pits, positive tonnage reconciliation was recorded in conjunction with negative grade reconciliations against the mineral reserve. This is mostly due to extra “visual” mineralization being mined, supervised by geologists and pit technicians. The majority of extra ore mined was grab sampled and assayed and returned grades well within the 0.6g/t Au cut-off. In short, “ore” which was not delineated by grade control was mined near the marginal grade, thus reducing the “As mined” predicted grade and increasing tonnes. Results showing positive call factors recorded in the same period illustrate the potential for ore/grade misallocation by grade control (Figure 14-52).

Throughout 2003 reconciliations figures were offset by positive mine to mill call factors (Figure 14-52). January 2003 to July 2003, mine to mill call factors averaged 119%, whilst ounce reconciliations have averaged 95%. This factor illustrates that the ounces lost between mineral resources and grade control are made up in ounces recovered from the mill.

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Classification

All mineral resource defined by Newmarket Gold owned the deposits have been classified as Inferred, mainly due to them not being estimated by Newmarket Gold staff. All estimates have been reviewed by the Author and deemed as appropriate for reporting. Drilling completed in 2011 confirms the grades of the deposits.

Below is a brief summary of other deposits in the Union Reefs area;

14.3.6.1  Low Grade Stockpile

During past mine production at the Union Reefs Gold Mine material of sub-economic grade (0.5 -1.0g/t Au) was stockpiled in close proximity to the eastern waste dump (Figure 14-53). This was placed there as it could have been reclaimed and processed if the economic conditions were sound. However, as the mine closed in 2003 the gold price did not warrant the material to be re-handled and processed so the material was covered with waste material and incorporated into the eastern waste dump. This material is still in the waste dump and could be recovered, however, some work to understand how to mine off the waste would be required, and hence why the material is classified as Inferred.

For this report the previously reported mineral resource estimation will be included until a new mineral resource estimation can be completed. This is due to the Author reviewing the data available and concluding that no material change would occur in generation of a new mineral resource estimation on this deposit so the previously generated estimation is suitable for inclusion in this technical report.

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14.3.6.2  Lady Alice

No additional drilling has been completed at the Lady Alice deposit since 2011, however, a significant change in the mineral resource generated prior to Newmarket Gold’s involvement in the deposit is not expected as most of the drilling completed to date is infill and not expansive in nature. Work is planned to review this drilling to determine if an update is required.

The Lady Alice deposit model used in this reporting is an open pit style estimation, which has been previously optimized. The model is estimated using an ID2 methodology with at top cut in the order of 25g/t Au. Wireframing was done on specific lodes by Bill Makar, who was the Chief Mine Geologist of the operation at the time.

The Lady Alice deposit is a possible location for the exploration portal that would be used to access mineralization from the Prospect deposit. If this were the case then it would allow for the exploration of the Lady Alice deposit from underground.

Drilling in 2011 and 2012 into the Lady Alice deposit showed higher grade mineralization was present but with limited continuity. Further work on the re-modeling process is required to better understand this issue.

For this report the previously reported mineral resource estimation will be included until a new mineral resource estimation can be completed. This is due to the Author reviewing the data available and concluding that no material change would occur in generation of a new mineral resource estimation on this deposit so the previously generated estimation is suitable for inclusion in this technical report.

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14.3.6.3  Millars/Big Tree/Ping Que

The Millars/Big Tree/Ping Que deposit is located on the Lady Alice line to the south of the Crosscourse open pit mine. There is historic evidence of significant underground workings in this area down to the water table with the Millars deposit having one of the largest shafts recorded in the Union Reefs area.

Crocodile Gold drilled this deposit in early 2011 before focus shifted to the Prospect and Crosscourse deposits. The mineralization that was noted in this area seems to be similar to that seen at Prospect with intersections in the order of 1.3m @ 27.8g/t Au and 1.7m @ 10.3g/t Au reported. Follow up work is required to incorporate these results into a new model using both older and newly acquired drilling data.

For this report the previously reported mineral resource estimation will be included until a new mineral resource estimation can be completed. This is due to the Author reviewing the data available and concluding that no material change would occur in generation of a new mineral resource estimation on this deposit so the previously generated estimation is suitable for inclusion in this technical report.

14.3.6.4  Union North

The Union North deposit is of interest to Newmarket Gold as it is directly on strike from the Prospect deposit and any future underground mining could easily access mineralization from this location. Union North is one of the deepest mines completed in the Union Reefs area nearing 100m deep when completed. Some drilling was done in this area in 2011 and 2012 and new mineralization wireframes were completed, however, despite the potential of changing the current mineral resource an update has not been completed.

For this report the previously reported mineral resource estimation will be included until a new mineral resource estimation can be completed. This is due to the Author reviewing the data available and concluding that no material change would occur in generation of a new mineral resource estimation on this deposit so the previously generated estimation is suitable for inclusion in this technical report.

14.3.6.5  Union South/Temple

Limited work has been completed on the Union South deposit; however, during a site visit and limited mapping exercise it was noted that the mineralization and alteration at the Union South open pit would be suitable for future large-scale mining. Work is required to update the mineral resource, however, at the time of writing this report this work had not been completed, hence the mineral resource here is the same as reported in 2013.

For this report the previously reported mineral resource estimation will be included until a new mineral resource estimation can be completed. This is due to the Author reviewing the data available and concluding that no material change would occur in generation of a new mineral resource estimation on this deposit so the previously generated estimation is suitable for inclusion in this technical report.

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14.4 PINE CREEK DEPOSITS

NEWMARKET GOLD PINE CREEK AREA MINERAL RESOURCE STATEMENT - Dec 31 2015

    INDICATED MINERAL RESOURCE       INFERRED MINERAL RESOURCE      
Project Deposit Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
Pine Creek






Cox 0.5 730,000 1.41 33,100 0.5 74,000 1.36 3,300
Czarina 0.5 1,046,000 1.80 60,600  
South Czarina
0.5
0.5
294,000
1,061,000
1.49
2.57
14,100
87,600
Enterprise  
Gandy's 0.5 535,000 1.81 31,100 0.5 482,000 2.92 45,300
International 0.5 5,112,000 1.19 195,600 0.5 197,000 1.29 8,200
Kohinoor 0.5 470,000 1.79 27,100 0.5 331,000 2.67 28,400
South Enterprise 0.5 500,000 1.99 32,000 0.5 101,000 1.35 4,400
  Total   8,393,000 1.41 379,400   2,540,000 2.34 191,300

TABLE 14-85 MINERAL RESOURCES FOR PINE CREEK DEPOSIT AS OF DEC 31 2015

Notes for Table 14-85:

1.

Mineral resources are stated as of December 31, 2015.

   
2.

Mineral resources are inclusive of mineral reserves.

   
3.

Mineral resources are calculated using these parameters.


  a.

Gold Price of $A1,500/oz, metallurgical recovery of 90-92.0% depending on mineral resource.

  b.

Lower cut-off of 0.5g/t Au for open pit mineral resources.

  c.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces.

  d.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.


4.

The mineral resource estimates were prepared by Mark Edwards, B.Sc. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold who has over 18 years of relevant experience and is a qualified person for mineral resources as per the NI43- 101.

At this point in time there are no known events or situations, which would materially affect the mineral resource as stated above, these include metallurgical, social, permitting, political, legal or environmental impacts.

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During 2015 all mineral resource estimations were reviewed. Further work will be required on some of the mineral resources but all are seen by the Author as being suitable for reporting. Newmarket Gold has elected to keep all mineral resources the same as the 2013 statement if no additional on ground work (drilling/sampling/mining) has been completed. Of all the deposits at Pine Creek, only the International deposit has had a mineral resource update completed by Newmarket Gold. The details of that report are outlined below.

The optimization process for each open pit deposit was as follows;

  (ii)

Model imported into MineMap™ software for processing.


  e.

Model was optimized using the Lerch-Grossman Pit Optimizer. This optimizer uses several inputs which are detailed as below:

  f.

Average density (SG) was set as the oxide density value on unpopulated blocks. Assigned density values in the models were used for populated blocks.

  g.

Gold price used was $A1,500/oz

  h.

Process recovery was set at 90% for all oxide and 85% for all fresh material, this reflects the average recovery seen through the process plant.

  i.

Fixed processing cost was set at $A22.50per tonne, this reflects the costs during 2012, which was the last time open pit mining was completed in the NT Operations

  j.

Mining costs were taken from the current contract estimates of $A4.50 per tonne of oxide material and $A5.00 per tonne of fresh material

  k.

The pit wall angle was also utilized; this was set to 40o for material mined within oxide and 50o for material mined below the oxide zone. These figures are generally what are used in Newmarket Gold’s current mining areas but will need more detailed review before mining can commence in new mining areas.

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  l.

When all these parameters are added to the optimizing process, an LG pit shell is generated for reporting.

  m.

For all optimizations of 2012 mineral resources, the LG1450 pit shell was then exported out of MineMap™ and imported into Micromine™ software. At this point the mineral resource estimation was then coded for material above the pit shell and below the current surface (which could be the mined from surface)

  n.

This coded data was then exported into Microsoft™ Excel for processing. The block size and cut-off grade was used to determine the tonnes and grade of material within the shell. All these model calculation spreadsheets have been saved for future use and review.

  o.

The optimized numbers were then entered into the current mineral resource statement above but only for mineral resources that have had new models completed in the past two years


Project Deposit Mineral resources Type New
Model
Drilling
2011/12
QA/QC
2011/12
SG
2011/12
Model
Constructed

by
Year of
Model
Pine
Creek
Cox Ind & Inf OP N N N N Makar 2004
Czarina Ind OP N N N N Geostats 2007
South Czarina Inf OP N N N N Makar 2004
Enterprise Inf OP N N N N Makar 2004
Gandy's Ind & Inf OP N N N N Makar 2004
International Ind & Inf OP Y Y Y Y Newmarket 2012
Kohinoor Ind & Inf OP N N N N Makar 2004
South Enterprise Ind & Inf OP N N N N Makar 2004

TABLE 14- 86 MODEL SUMMARY FOR PINE CREEK DEPOSITS

Project Mineral resource Method Grade cap Au g/t Block size
E x N x RL
(meters)
Pine Creek Cox ID 30 2.5 x 10 x 2.5
Czarina OK 20, 8, 7 (by Lode) 5 x 10 x 5
South Czarina ID 30 2.5 x 10 x 2.5
Enterprise ID 30 2.5 x 10 x 2.5
Gandy's ID 30 2.5 x 10 x 2.5
International OK 10 5 x 12.5 x 2.5
Kohinoor ID 30 2.5 x 10 x 2.5
South Enterprise ID 30 2.5 x 10 x 2.5

TABLE 14-87 PINE CREEK MODEL SUMMARY OF MODEL INPUTS

Overall the bulk density data has been reviewed by the Author and deemed as being suitable to be used in all historic models which has been determined from previously reported mineral resources. The numbers used are reported below, from McGuire et al 2007.

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The Czarina deposit lies adjacent to the excavated Enterprise deposit, with both deposits characterizing similar host rocks and mineralization styles. Density values for Czarina were not obtained as a result of the predominant chip nature of drill samples from RC drilling and thus those used for the Enterprise deposit were also used for the Czarina deposit. Historical density records from Enterprise are scarce, and the density values used for Czarina were obtained from the historic June 1989 Enterprise Reserve Report. A total of five drill core samples in oxide material averaged 2.4g/cm 3, with a total of 16 drill core samples in fresh material giving an average density of 2.7g/cm 3. It is recommended that additional density work be completed

Work completed by Crocodile Gold on the International deposit confirmed the above assumptions for bulk density. A total of 21 sulphide diamond core samples were analyzed for bulk density using the water immersion method with an average density of 2.75g/cm 3 determined in non-mineralized material and 2.80g/cm 3 in mineralized material. Only two oxide samples were tested with an average density of 2.65g/cm 3 recorded, which is slightly higher than that used in historic models. Overall the slight reduction in density used in the historic models is probably sound and appropriate. More test-work is required to confirm these results in other deposits.

14.4.1      INTERNATIONAL DEPOSIT

14.4.1.1  Introduction

During December 2012, Cube Consulting Pty Ltd was requested to undertake a mineral resource estimation update of the International deposit for Crocodile Gold. The estimation was based on historic drill holes as detailed in Section 10. Figure 14-55 shows a plan view of the drilling on the International deposit area used for this mineral resource estimate. The majority of the drilling undertaken during past exploration and mining was RC, shown in green. Some diamond core and RAB drilling was also completed in the area shown in red and light blue respectively.

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Figure 14-56 shows a typical cross section through the deposit at Section 12550 mN in the local grid system, looking north. It shows four mineral resource interpretations including the two most significant, domains 100 and 200. Also shown are the RC and diamond core hole traces utilized in defining the mineral resource.

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14.4.1.2  Data Types

The estimation of contained gold has been based on assays sourced from drilling data. The data available consisted of RC, RAB and diamond core samples from historic exploration and mining definition campaigns. In addition, five diamond and two RC holes drilled by Crocodile Gold were also included.

All data is in local grid co-ordinates.

Drilling provides data in close proximity to the pit area to depths up to 240m below surface. The total database supplied consisted of 570 drill holes, including 506 (for 23,995.8m) RC holes, 36 (for 2,825.59m) diamond core holes and 28 (for 480m) RAB holes.

Within the mineralized domains the drill data consisted of 624 RC intercepts for 9,057 downhole meters, 82 diamond core intercepts for 894.65 downhole meters and 8 RAB intercepts for 66 downhole meters.

A visual comparison in section between diamond and RC data types was completed to test if any material difference was observed between data types. Several cases were observed where grade continuity in between data types was present. Where differences were observed it was not possible to determine whether the grade variability was due to a difference in drilling type or to the short range gold variability observed generally in the deposit and characterized in the variograms. As a consequence, it has been decided to include the diamond data in the estimation to improve data density.

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Similarly, a comparison in section between RAB and RC data types was also completed to test if any material difference was observed between data types. RAB data within mineralized domains is shallow reaching up to 18m below pre-mined surface. In spite of the lower quality associated to this data type, RAB intercepts have been used to guide the end point of some wireframes in the northern part of the deposit. After careful observation no material difference was observed between data types. Where differences were observed it was not possible to determine if the variability was due to a difference in drilling type or to the short range variability observed in the deposit. As a consequence, it has been decided to include the RAB data contained within the mineralized domains in the estimation to improve data density in areas where no other data is available.

The drill cuttings and core were sampled and assayed on varying lengths as summarized within mineralized lodes in Table 14-88.

Mineralised Domain Minimum Length (m) Maximum Length (m)
100 0.35 2.0
200 0.3 2.0
210 0.3 2.0
300 1.0 1.04
400 0.39 3.0
500 0.5 2.0
600 0.6 3.1
900 2.0 2.0

TABLE 14-88 INTERNATIONAL DEPOSIT SUMMARY OF SAMPLE LENGTHS BY MINERALIZED DOMAIN

14.4.1.3  Geological Interpretation

The International deposit is interpreted to be a steeply dipping semi-continuous gold mineralized veining system domain.

The host lithologies of the International deposit have been divided into two main states of weathering/oxidation (oxidized and fresh). A surface interpreted from historic geological logging has been used to flag oxidation state into the three dimensional block model.

An examination of the possible effects the state of oxidation has on the grade tenor has been undertaken. Figure 14-57 shows the comparison of gold sample grades by oxidation state for gold within the mineralized domains. The following inferences can be drawn from the figure:

  a.

The statistics are not calculated on equal support but on samples of varying lengths;

  b.

Of the total mineralized samples 71% are un-oxidized and 29% are oxidized;

  c.

The gold grades present a low variation with oxidation state; a mean of 1.28g/t Au for oxidized and 1.17g/t Au for un-oxidized;

  d.

The statistical variability of un-oxidized and oxidized samples is similar with CVs of 2.2 for oxidized samples and 1.7 for un-oxidized samples.

An examination of the interpreted surface with close inspection of the boundary behavior of grades was undertaken to test support for the application of the oxidation boundaries during estimation. The interpreted boundary shows a degree of variability in position from hole to hole, related to the intensity of shearing and degree of alteration this means the position of the boundary is not exact. Additionally, the grade transition across the boundaries is shown to be graduation.

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As a consequence, the Company has not used the interpreted oxidization surfaces as hard boundaries during the estimation of gold.

14.4.1.4  Mineral Resource Interpretation

Interpretation of mineralized domains has been informed by gold cut-off grade, with a lower limit of 0.4g/t Au used as the basis for defining mineralized veining material. No minimum length criterion has been applied during the interpretations.

The mineral resource domain interpretations were wireframed and nominated 100, 200, 210, 300, 400, 500, 600 and 900. Figure 14-58 below shows the wireframes and drilling traces in an oblique view to the northwest.

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The mineral resource wireframes were used to code the drill intercepts contained within them by flagging into a new table in the database, the zonecode table. This flagging allows the selection of data within domains by codes for the purposes of sample analysis and compositing.

All mineral resource interpretation wireframes have been used as hard boundaries for this estimate.

14.4.1.5  Compositing and Statistics

Compositing of the raw drilling sample data is necessary to establish a single support for the data to avoid bias when calculating statistics and undertaking any estimation of the data into three dimensional volumes. A number of items are considered when selecting an appropriate composite length; they include the original support of the raw sample data, the assumed selectivity (and therefore the block size) of the model and the imposed spatial dimensions of the mineralized domains.

An examination of sample statistics for mineral resource domains reveals that the majority of sampling of the mineralization is on 1 meter downhole support, where sample lengths vary from a minimum of 0.3m to a maximum of 3.1m downhole.

The number of instances of samples over 1 meter are small representing less than 20% of the data. In particular, the mineralized domain 900 includes 2m samples exclusively.

Within the mineralized domains the drill samples were composited to 2m downhole using a best fit algorithm to provide equal support data for estimation. The best fit compositing method was used with a tolerance of 30%, yielding composites of between 0.6m and 2.6m in down hole length.

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The effect of a small number of outlier composite grades or spatially isolated composites may have an undue effect of the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the composite data in 3D visualization.

Complete summary statistical tables and charts are available by individual domain in Table 14-89. A statistical summary of the mineralized domains is shown in Table 14-90, below.

A number of high cuts or limits were identified as necessary within domains as detailed Table 14-90, below. High-grade gold cuts were applied to the 2m composites within the mineralized domains and the cut composites used in the estimation. A general cut of 10g/t Au was applied to all the domains. In addition, cuts to five individual composites included in domains 100, 500 and 900 were also used.

Domain Minimum Gold
Grade g/t Au
Maximum Gold
Grade g/t Au
Mean Gold
Grade g/t Au
High Grade Cut
g/t Au
Mean Cut Gold
Grade g/t Au
100 0.0050 41.73 1.26 10 (7.5) 1.20
200 0.0050 54.50 1.29 10 1.17
210 0.0520 33.10 1.23 10 1.17
300 0.1100 24.51 1.36 10 1.26
400 0.0080 15.80 1.23 10 1.20
500 0.0100 14.62 1.11 10 1.09
600 0.0150 19.24 1.18 10 (5) 1.11
900 0.0100 9.87 0.91 10 (5) 0.89

TABLE 14- 89 INTERNATIONAL DEPOSIT HIGH-GRADE COMPOSITE CUTS BY DOMAIN

Domain Number Cut Mean
g/t Au
Cut Median
g/t Au
Standard
Deviation
Co-efficient of
Variation
100 2,043 1.20 0.74 2.19 1.24
200 1,232 1.17 0.74 2.27 1.29
210 390 1.17 0.73 1.92 1.18
300 148 1.26 0.77 1.88 1.09
400 465 1.20 0.72 2.44 1.30
500 409 1.09 0.75 1.42 1.09
600 310 1.11 0.67 2.02 1.28
900 108 0.89 0.53 1.65 1.45

TABLE 14- 90 INTERNATIONAL DEPOSIT STATISTICAL SUMMARY BY DOMAIN

The general statistics of gold composites within all domains can be described as positively skewed with moderate to high variability as is the case with most gold occurrences. The high variability is reduced somewhat by high cutting of gold grades in those domains most affected.

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Within the mineralized domains a measure of variability, the Co-efficient of Variation (CV) remains at an approximate value of 1.2 for most domains, indicating variability was reduced within the domains after high-grade cutting.

14.4.1.6  Variography

Variography was used to characterize the spatial behavior of the composite data primarily as an aid to establishing estimation parameters. Variogram stability and quality is dependent on the statistical properties of defined domains and the amount of data available within domains. After an initial investigation, the final model established for all domains was based on the modeling of data from domain 100 the most populous with 2,043 composite data. This was considered the most robust solution to very noisy models observed within individual domains as a result of moderate to high domain variability.

The final variogram model is detailed in Table 14-91, below.

Domain Nugget Structure Sill Range
(m)
Azimuth Plunge Dip Major/Semi Major/Minor
100 1.00 1 0.80 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1

TABLE 14- 91 INTERNATIONAL DEPOSIT FINAL VARIOGRAM MODELS BY DOMAIN

The variogram model for the 100 domain was modeled in Gaussian space and then back transformed to real space. It contains three spherical structures with a 48% nugget. The first structure includes 39% of the sill within 5m, the second structure includes 8% of the sill within a range of 30m and the third structure includes the remaining 5% of the sill within 135m. The variogram model appropriately reflects the differences seen in the summary statistics of the composite data. The modeling was unable to determine any distinct axes of preferential continuity and reflects the high variability at short range observed in visual inspection by section.

14.4.1.7  Grade Interpolation Methodology

A standard three dimensional single pass Ordinary Kriging methodology has been used for the estimation of the cut gold 2m downhole composite data within each mineralized domain. Table 14-92 details the variogram models used and Table 14-93 summarizes the estimation parameters by domain. A constant minimum of four and maximum of 40 composites have been set for most domains except for domains 100 (anticline crest) and 500, where a minimum of two composites was used. In addition, a maximum of four composites per hole was also applied. Block discretization of 2 in x, 5 in Y and 1 in Z have been used throughout.

Domain Nugget Structure Sill Range (m) Azimuth Plunge Dip Major/ Semi Major/ Minor
100 1 1 0.8 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1
200 1 1 0.8 5 0 0 0 1 1

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Domain Nugget Structure   Sill Range (m) Azimuth Plunge Dip Major/ Semi Major/ Minor
     2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1
210 1 1 0.8 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1
300 1 1 0.8 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1
400 1 1 0.8 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1
500 1 1 0.8 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1
600 1 1 0.8 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1
900  1 1 0.8 5 0 0 0 1 1
  2 0.16 30 0 0 0 1 1
  3 0.11 135 0 0 0 1 1

TABLE 14- 92 INTERNATIONAL DEPOSIT MINERALIZED DOMAIN ESTIMATION VARIOGRAM MODEL

Domain Search
Radius (m)
Azimuth Plunge Dip Major/Semi Major/Minor
100 (anticline flank) 170 180 0 -65 2.5 10
100 (anticline crest) 170 180 0 0 2.5 10
200 170 180 0 -65 2.5 10
210 170 180 0 -65 2.5 10
300 170 180 0 -65 2.5 10
400 170 180 0 -65 2.5 10
500 180 180 0 -65 2.5 10
600 170 180 0 -65 2.5 10
900 170 180 0 65 2.5 10

TABLE 14- 93 INTERNATIONAL DEPOSIT MINERALIZED DOMAIN ESTIMATION PARAMETERS

Block Model Definition

The primary consideration of the 3D model was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each mineralized zone. All individual mineralized zones were ultimately combined to create a single 443



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block model in the local grid coordinate system. Table 14-94 summarizes the 3D block model “International_Dec12.mdl” definition.

  Northing Easting RL
Minimum 12000 11200 1050
Maximum 13500 11500 1250
Block Size (m) 12.5 5 2.5
Sub-block (m) 6.25 2.5 2.5

TABLE 14-94 INTERNATIONAL DEPOSIT FINAL 3D BLOCK MODEL DEFINITION

The chosen block size represents approximately half the best data spacing in the Northing direction and a choice in the vertical and easting dimension controlled by the need to appropriately represent the volume of the wireframes.

A standard list of field names and descriptions used in the block model are shown in Table 14-95.

Attribute Type Default Description
Au Float 0 estimated Gold ppm
Density Float 0 Density
Zonecode Char BKGR Zonecode
Rescat Char Unclassified Measured, Indicated, Inferred, Exploration, Unclassified
Rescatnum Integer 5 Meas =1; Ind = 2; Inf = 3, Explor = 4; Unclass=5
Depletion Integer 1 Insitu = 1; Mined = 0
Oxidation Char Undefined Undefined, Fresh, Transition, Oxide
Oxidation num Integer 0 0= Undef; 1 = Fresh; 2 = Trans; 3 = oxidised
Ads Float 0 Average distance to composite data
Dns Float 0 Distance to nearest composite data
Kv Float -1 Kriging Variance
Ns Integer 0 Number of composite data
Search_type Integer 1 1= semivertical search; 2= horizontal search

TABLE 14-95 INTERNATIONAL DEPOSIT FINAL 3D BLOCK MODEL ATTRIBUTES

Table 14-96 below confirms the close agreement of the 3D block model volumes and the original interpreted wireframe volumes, supporting the 3D model block size choice as appropriate.

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Domain Wireframe
Volume
Block Model
Volume
Variance
%
100 1,662,501 1,665,859 100.2
200 1,021,760 1,024,180 100.2
210 242,888 243,516 100.3
300 150,087 149,805 99.8
400 244,880 245,273 100.2
500 281,568 281,523 100.0
600 201,858 202,070 100.1
900 121,088 121,055 100.0
Total 3,926,630 3,933,281 100.2

TABLE 14-96 INTERNATIONAL DEPOSIT FINAL 3D BLOCK MODEL TO WIREFRAME VOLUME CHECK

14.4.1.9  Specific Gravity / Bulk Density Assignment

The specific gravity of the waste and mineralized rock of the final 3D block model has been assigned according to oxidation state, using the interpreted surface described in Section 14.2.3 to control the blocks assigned. Crocodile Gold determined the specific gravity based on historic reports of determinations made on drill samples. Table 14-97 summarizes the oxidation state specific gravity assignations.

Oxidation State SG
1                                   Fresh 2.7
3                                   Oxidized 2.5

TABLE 14- 97 INTERNATIONAL DEPOSIT SPECIFIC GRAVITY VALUES BY OXIDATION STATE

Within the final 3D block model blocks coded fresh were assigned a density of 2.7g/cm 3; those coded oxidized below the topographical surface, a density of 2.5g/cm 3; and those blocks below the topographical surface and above the historic pit surface (representing a backfilled volume) a density of 1.9g/cm 3. Blocks located above the topographical surface were assigned a zero density.

14.4.1.10  Model Validation

Model validation has been undertaken to ensure no material error has been made in the estimation of International. The validations include inspection of the audit documentation of the individual estimation runs; visual inspection of the block outcomes and input data; statistical comparisons of input data and block outcomes, swath plots of each of the domains, and comparison with an alternative estimation method.

Statistical comparisons of input data and block model outcomes for the mineralized domains are shown in Table 14-98.

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Domain Composite
Average
Grade
g/t Au
Block Model
Average
Grade
g/t Au
Variance
%
Percentage of
Total Gold
Ounces
Indicated
Model
Average
Grade g/t Au
Variance
%
100 1.26 1.13 90 34 1.13 90
200 1.29 1.22 95 24 1.17 91
210 1.23 1.10 89 6 1.10 89
300 1.36 1.16 85 4 1.19 88
400 1.23 1.30 106 6 1.30 106
500 1.11 1.15 104 7 1.13 102
600 1.18 1.14 97 4 1.15 97
900 0.91 0.83 91 2 - -

TABLE 14- 98 INTERNATIONAL DEPOSIT MINERALIZED DOMAIN AVERAGE GOLD GRADE COMPARISONS

The mineralized domain comparisons display a moderate variation between input and outcome average grades when the total domain is reported. As can be confirmed in the visual inspection and swath plot investigations the comparisons include significant volumes at depth and in border areas in each domain containing a lower density of sample data. This results in extrapolation of the sample data into these volumes and while it is considered a reasonable estimate of the grades in these volumes a simple statistical comparison of total volumes will not result in close comparisons. Figure 14-59 below demonstrates this situation within the mineralized domain 100. Two views of the block model domain 100 are shown, the first with composite data and blocks colored by gold grade and the second with blocks colored by mineral resource classification. Composite data is colored by gold grade distribution as shown in the legend. The contrasting data densities at depths are evident, supporting the assumption that relying only on raw composite to block grade comparisons can be misleading.

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Comparison of the Indicated portions of the mineralized domains in Figure 14-60 above show that for the most significant domains by contained ounces (100 and 200) the comparison to average composite grades agrees within a 10% tolerance.

Swath plot validations for all the mineralized domains are presented in Figure 14-61. Data is analyzed by northing and by elevation for each domain. Reproduced below are four example swath plots, one pair for domain 100, Figure 14-61 and one for domain 200, Figure 14-62.

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The swath plots demonstrate that where there is regularly spaced data the block model reflects those data. The plots also highlight that a paucity of drilling results in parts of the model that rely on only a few measured points and are therefore less likely to match local composite data, and are of reduced certainty and increased risk.

Comparison with an alternative estimation method was also performed as a form of validation. Multiple Indicator Kriging (MIK) was the selected alternative methodology. Table 14-99 shows the total metal obtained for each method at cut-offs 0.5g/t and 1.0g/t Au. It was observe that the difference of total metal for each cut-off is within 3% and 7% respectively with MIK results producing more metal than Ordinary Kriging (OK) as expected for this method.

Au cut off Au Metal - OK
(000’ Oz)
Au Metal - MIK
(000’ Oz)
0.5 g/t 296 306
1.0 g/t 216 231

TABLE 14-99 INTERNATIONAL DEPOSIT COMPARISON OF TOTAL METAL BETWEEN OK AND MIK
METHODS FOR ALL MINERALIZED DOMAINS

Figure 14-63 shows the comparison of grade tonnage curves between both methods with grades being slightly higher in the case of MIK at all cut-offs. On the other hand, tonnage for MIK is slightly lower for cut-offs less than 1.0g/t Au and slightly higher for cut-offs higher than 1.0g/t Au. However, at cut-offs 0.5g/t and 1.0g/t Au the total metal, resulting from each method is almost equivalent as shown in the previous table.

An additional curve indicating a theoretical change of support (using a form of local MIK) from the estimation panel of 12.5m x 5m x 2.5m (Y x X x Z) to an SMU of 6.25m x 2.5m x 2.5m (Y x X x Z) is also presented in Figure 14-64.

Finally a comparison of total metal curves for each method is presented in Figure 14-64.

These results confirm the validity of the methodology used and indicate that OK Block Kriging results are a robust global estimate.

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14.4.1.11  Mineral Resource Classification

The mineral resource has been classified in accordance with NI43-101 guidelines. Assessment criteria include data integrity, drillhole spacing, sample locations, sampling density, and lode geometry, geological confidence and grade continuity. Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data.

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The mineral resource has been classified as Indicated, Inferred, and Exploration target; no Measured Mineral resource has been identified. Exploration target material has not been reported or outlined in anyway in this report.

14.4.1.12  Data Spacing and Distribution

The International deposit model has been shown in validation to be subject to varying drillhole density and sample location in relation to the lode geometry. In most lodes the drilling is regular and have sufficient density within the upper/central parts of lodes but subject to decreasing densities and irregular spacing at depth. The block model outcomes at depth in most lodes are considered to be higher risk and are classified with less confidence than the shallower parts. Each lode was considered individually and for lodes with sufficient data density a boundary limit digitized for the base of indicated.

Within domain 200 a small section between 12400 and 12475mN has been identified that lacks sufficient data density to be classified as mineral resource. Geological continuity has been reasonably assumed however the blocks estimated all display elevated grades based on the limited data available. This portion has been adjudged to be exploration target only as the risk and uncertainty associated with the estimated grade is high.

14.4.1.13  Orientation of Data in Relation to Geological Structure

The orientation of the deposit is interpreted to be close to vertical and the drilling is considered to be appropriately targeted for this geological orientation.

14.4.1.14  Deposit Dimensions

The mineralized portion of the International deposit extends within drill testing from 12,150 to 13,450m in Northing with extension beyond drilling of up to 12.5m; within the Easting plane the dimensions of the mineralization are tightly constrained by drilling extending from 11250 to 11450m; in the vertical the deposit extends within drilling from surface (at approximately 1240mRL) to 1100mRL, extension beyond drilling was up to 12.5m from the last data point. The dimensions of the mineralization are adequately defined by the available drilling with limited and acceptable extensions beyond data.

14.4.1.15  Estimation and Modeling Techniques

The estimation methodology used for the mineralization style is considered appropriate by the CP based on experience with similar deposit types. It is shown to represent reasonable unbiased reproductions of the input data in areas of adequate sampling. Outside areas of adequate sampling the mineral resource classification is such as to reflect the uncertainty of the estimate. The validation methods used also demonstrate the adequacy of the methodology used.

14.4.1.16  Moisture

The estimate has been made on the basis of dry tonnes.

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14.4.1.17  Classification

All material within the mineral resource interpretation has been classified to represent the opinion of the risk in the mineral resource estimated. Within the mineralized domains that have been defined on a plus 0.4g/t Au cut off, it is assumed that some of the material will form dilution to the mining of higher grade material. For reporting purposes the mineralized material has been reported with a lower cut off of 0.5g/t Au within the interpreted wireframes. The classification of the International mineral resource into Indicated and Inferred mineral resources and Exploration Target as set out below reflects the view of this deposit, as it is currently defined.

14.4.1.18  Selectivity Assumptions

The mineral resource estimate contains implicit assumptions of mining selectivity represented by the block size of 12.5m x 5m x 2.5m (Y x X x Z).

14.4.1.19  Discussion of Relative Accuracy/Confidence

The only available measure of accuracy of the estimate is the comparison with historic production data from the International Mine. Survey data of the final pit is available while the original topography was inferred based on collar data. Table 14-100 presents the results of historic production against a mineral resource estimate of the estimated mined volume at a cut-off of 1.2g/t Au. An overall recovery of 79% was used to calculate metal. The results compare well with the current model staying within a 4% difference in metal from historic figures. The variations observed in the grade and tonnes are sufficiently small to be accounted for by a change of support from mineral resource block to mine SMU as demonstrated in the curves of Figure 14-63.

Tonnes Au (ppm) Recovered
Metal (Oz)
Historic Prod. 745,000 1.62 30,681
Current Model 763,199 1.52 29,368
Difference 2% -6% -4%

TABLE 14- 100 INTERNATIONAL DEPOSIT COMPARISON OF HISTORIC PRODUCTION DATA AND CURRENT MODEL (OVERALL RECOVERY OF 79%)

14.4.1.20  Recommendations

A significant quantity of original supporting data is available in hard copy form. It is recommended that this information be collated, reviewed and digitally data based. This would have a number of benefits, firstly, it preserves the work undertaken so far, which has some considerable value to the Company; secondly this facilitates the Company to ascertain if additional QA/QC data has been overlooked and if further QA/QC data is required.

In order to improve the quality of the estimated mineral resource the following actions are also recommended:

  1.

Undertake infill diamond drilling of the deeper extends of priority lodes to confirm the assumptions of geological continuity inherent in the current estimate;

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  2.

Undertake infill drilling to confirm the continuity of Lode 200 in the vicinity of Northing 12,400 to 12,475m;

  3.

Undertake density measurements on diamond core drilling to build a database of density values;

  4.

Undertake a mining study to determine the limits of economical extraction of the deposit by open pit;

  5.

Undertake a check/validation process of sampling and assaying by resampling and assaying some of the available drilling sample material.

14.4.1.21  Cox

The Cox mineral resource model was produced by Makar in 2004. It uses a 30g/t Au top cut. The model has not been mined from surface. Wireframing was completed in two zones, the first concentrated on the well drilled main zone and the second looked at the poorly drilled area to the southwest of the main zone. No new drilling has occurred on this model since Crocodile Gold/Newmarket Gold took ownership of the deposit. Additional drilling may be required for metallurgical and geotechnical studies, in the manner drilling was conducted at the International deposit in 2012.

The modeling technique used was ID2, with search ellipses around 20m down dip, 30m along strike and 5m RL. The rotation on the search ellipsoid was 15° on strike, 90° down drip and 10° down plunge. This model was previously reported in both 2009 and 2011 by Crocodile Gold and is deemed by the Author as being appropriate to include in this technical report.

14.4.1.22  Czarina

Below is taken from McGuire et al 2007; Geostat Services Pty Ltd (Geostat) was commissioned by a previous owner to complete an updated mineral resource estimate for the Czarina deposit, Northern Territory, Australia in 2007. Mineral resource estimation was undertaken in compliance with CIM mineral resource and mineral reserve definitions that are referred to in National Instrument (NI) 43-101, Standards of Disclosure for Mineral Projects. All data has been reviewed by the Author and deemed as being appropriate for inclusion in this technical report. This review has included open discusions with the consultant for Geostat who also consulted to Newmarket Gold for a period between 2009 and 2013.

The Czarina deposit comprises a meta-sedimentary hosted gold deposit, with mineralization sub-parallel to bedding. The mineralization is along the western axis of the Czarina Anticline, striking about 315° and plunging gently to the south. The fold is parallel to the Enterprise Anticline, host to the previously operating Enterprise Mine.

Three dimensional (3D) modeling methods and parameters were adopted in accordance with best practice principles. A 3D lode wireframe interpreted from drillhole geology and assays was completed. Statistical and grade continuity analyzes were completed to characterize the mineralization and subsequently used to develop grade interpolation parameters. These were applied to the supplied 3D lode wireframes.

Surpac mining software was used for generating the 3D block model and subsequent grade estimate. Top-cuts were used to restrict the influence of statistical outliers during Ordinary Kriging of block grades. A bulk density model was generated by Geostat using data previously collected by previous owners. A mineral resource classification scheme consistent with CIM guidelines was applied. The estimate is categorized as Indicated and Inferred mineral resources and reported above a gold grade cut-off that is appropriate for a potentially mineable deposit.

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14.4.1.23  South Czarina

The South Czarina deposit model was produced by Makar in 2004. It uses a 30g/t Au top-cut. The model has not been mined from surface. Wireframing was completed on both the South Czarina deposit and the Czarina deposit main area. No new drilling has occurred on this model since Crocodile Gold/Newmarket Gold took over ownership of the deposit. Additional drilling may be required for metallurgical and a geotechnical study in the manner drilling was conducted at the International deposit in 2012. All data has been reviewed by the Author, including having Makar conduct several site visits for Newmarket Gold to review the work completed. It is in the opinion of the Author that this work remains suitable for inclusion in this technical report.

The modeling technique used was ID2, with search ellipses around 30m down dip, 60m along strike and 5m RL. The search ellipsoid was rotated at 3° along strike, dipping 90° and plunging at 14° towards the south. Additional QA/QC data is available for the South Czarina deposit as previous owners have drilled it in the past.

14.4.1.24  Enterprise

The Enterprise deposit mineral resource was produced by Makar in 2004. The mineralization was defined on one major zone with several smaller, less continuous lodes wireframed in the footwall of the deposit. The majority of the mineralization is located in the main zone. All data has been reviewed by the Author, including having Makar conduct several site visits for Newmarket Gold to review the work completed. It is in the opinion of the Author that this work remains suitable for inclusion in this technical report.

The main mineralized zone is continuous through the deposit and plunges towards the south end of the existing open pit. Due to a lack of recent drilling this mineral resource was only classified as inferred but the grade and potential of the deposit suggests additional drilling would be required to up-grade the confidence in the model.

The methodology of this model follows the process used at the Cox deposit. Generally, it was produced using the ID2 method. There is no rotation on the model with the searches being 50m down dip, 50m along strike and 5m down RL. The search ellipsoid is 0° along strike, 90° down dip and 15° south down plunge

14.4.1.25  Gandy’s

The Gandy’s deposit mineral resource was constructed by Makar in 2004 using a 30g/t Au top-cut. It had the mineralization constructed using two zones, the first looking at the South Gandy’s area and the other the North Gandy’s area. The methodology used was also ID2 using a search ellipsoid of 30m down dip, 70m along strike and 5m down RL. All data has been reviewed by the Author, including having Makar conduct several site visits for Newmarket Gold to review the work completed. It is in the opinion of the Author that this work remains suitable for inclusion in this technical report.

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The rotation of the search ellipsoid was 0o along strike, 90° down drip and 14° south down plunge. Both deposits have been previously mined so the material mined has been removed from the calculated mineral resource quoted in this report. South Gandy’s has been partially backfilled and the North Gandy’s pit has been completely backfilled. This has been considered when looking at the economic viability of mining these deposits.

In previous reports the North and South Gandy’s deposits were reported separately, in this report the zones have been combined for simplicity.

In late 2012 Crocodile Gold re-interpreted the Gandy’s deposit to identify the potential for up-grading the deposit with what was learnt from the drilling at the International deposit. While this mineral resource estimate is not reported in this document but it could be used to assist with targeting future drilling. It is anticipated that a similar path can be taken with Gandy’s deposit as has been concluded with the International deposit, with the additional of a few well-placed holes allowing for an updated model to be produced.

14.4.1.26  Kohinoor

The Kohinoor deposit is located to the south of the Cox deposit. It sits along the southern extent of the Czarina line of gold deposits. Mineralization at Kohinoor is slightly rotated, resulting in a rotated mineral resource. The rotation is around 9° towards the east.

The northern part of the deposit is well drilled and is classified mainly as indicated, the southern part of the deposit is poorly drilled in comparison and is there classified as inferred. The methodology used on this model is ID2. No drilling has been conducted by Crocodile Gold/Newmarket Gold since taking over ownership of the project

The searches used on the Kohinoor model are 40m down dip, 90m along strike and 5m down RL. For the second pass this was reduced to 20m down dip, 50m along strike and 5m down RL. The search ellipsoid rotation is 9° for strike, dipping 75° towards the west and plunging at 0°.

14.4.1.27  South Enterprise

The Burnside Joint Venture drilled 34 RC holes into the South Enterprise deposit in 2004 and 2005. This drilling was used in a mineral resource update; however, this has not been previously reported.

The methodology and searches from the South Enterprise deposit are the same as the Enterprise deposit outlined above. All information has been reviewed by the Author and deemed suitable for reporting.

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14.5 BURNSIDE AREA

Deposit Indicated mineral resource Inferred mineral resource
Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
Cut-off
(Au g/t)
Tonnes Grade
(Au g/t)
Ounces
Gold
HowleyB 0.7 5,836,000 1.22 228,900 0.7 1,351,000 1.41 61,200
MottramsB 0.7 204,000 1.17 7,700 0.7 169,000 1.14 6,200
North PointB 0.7 139,000 1.43 6,374 0.7 117,000 1.31 4,900
Princess LouiseB 0.7 394,000 1.30 16,500  
Rising Tide 0.7 292,000 1.45 13,600 0.7 372,000 1.49 17,800
Fountain Head 0.7 273,000 1.79 15,700 0.7 99,000 1.95 6,200
Tally Ho* 2.0 221,000 4.71 33,400 2.0 114,000 4.86 17,900
Kazi   0.7 410,000 1.95 25,700
Western Arm   0.7 3,383,000 1.11 120,300
Bon's Rush   0.7 805,000 2.33 60,400
Sub-total   7,358,000 1.36 322,200   6,820,000 1.46 320,600

TABLE 14-101 RESOURCE ESTIMATIONS NEWMARKET GOLD DEPOSITS NORTHERN TERRITORY

Notes for Table 14-101:

5.

Mineral resources are stated as of December 31, 2015.

   
6.

Mineral resources are inclusive of mineral reserves.

   
7.

Mineral resources are calculated using these parameters.


  e.

Gold Price of $A1,500/oz, metallurgical recovery of 90-92.0% depending on mineral resource.

  f.

Lower cut-off 2.0g/t for all underground mines and 0.7g/t Au for open pit mineral resources.

  g.

All tonnes are rounded to the closest 1,000t and ounces are rounded to the closest 100 ounces.

  h.

Mineral resources that are not mineral reserves do not have demonstrated economic viability.


8.

The mineral resource estimates were prepared by Mark Edwards, B.Sc. MAusIMM (CP) MAIG, General Manager Exploration for Newmarket Gold who has over 18 years of relevant experience and is a qualified person for mineral resources as per the NI43- 101.

At this point in time there are no known events or situations, which would materially affect the mineral resource as stated above, these include metallurgical, social, permitting, political, legal or environmental impacts.

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During 2015 all mineral resource estimations were reviewed and are deemed suitable for reporting. The only notable changes to the Newmarket Gold mineral resource statement has been the removal of the Glencoe and Bridge Creek deposits’ mineral resources, as these deposits have been divested to third parties.

The optimization process for each open pit deposit was as follows;

  (iii)

Model imported into MineMap™ software for processing.


  a.

Model was optimized using the Lerch-Grossman Pit Optimizer. This optimizer uses several inputs which are detailed as below:

  b.

Average density (SG) was set as the oxide density value on unpopulated blocks. Assigned density values in the models were used for populated blocks.

  c.

Gold price used was A$1,500/oz.

  d.

Process recovery was set at 90% for all oxide and 85% for all fresh material, this reflects the average recovery seen through the process plant.

  e.

Fixed processing cost was set at $A22.50 per tonne, this reflects the costs during 2012. These costs match milling costs during open pit mining operations.

  f.

Mining costs were taken from the current contract estimates of $A4.50 per tonne of oxide material and $A5.00 per tonne of fresh material.

  g.

The pit wall angle was also utilized; this was set to 40° for material mined within oxide and 50° for material mined below the oxide zone. These figures are generally what are used in Newmarket Gold’s current mining areas but will need more detailed review before mining can commence in new mining areas.

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  h.

When all these parameters are added to the optimizing process, an LG pit shell is generated for reporting.

  i.

For all optimizations of 2012 resources, the LG1450 pit shell was then exported out of MineMap™ and imported into Micromine™ software. At this point the mineral resource estimation was then coded for material above the pit shell and below the current surface (which could be the mined from surface).

  j.

This coded data was then exported into Microsoft™ Excel for processing. The block size and cut-off grade was used to determine the tonnes and grade of material within the shell. All these model calculation spreadsheets have been saved for future use and review.

  k.

The optimized numbers were then entered into the current mineral resource statement above but only for mineral resources that have had new models completed in the past two years.


Deposit Mineral
resources
Type New
Model
Drilling
2015
Mining 2015 QA/QC 2015 SG 2015
Howley Ind & Inf OP N N N N N
Mottrams Ind & Inf OP N N N N N
Princess Louise Ind & Inf OP N N N N N
Rising Tide Ind & Inf OP N N N N N
Fountain Head Ind & Inf OP N N N N N
Tally Ho Ind & Inf UG N N N N N
Kazi Inf OP N N N N N
Western Arm Inf OP N N N N N
Bon's Rush Inf OP N N N N N

TABLE 14-102 COMMENTS ON MINERAL RESOURCES ESTIMATIONS OF NEWMARKET GOLD DEPOSITS, NORTHERN TERRITORY

Deposit Method Grade Cap Au g/t Block size E x N x RL
(meters)
Howley OK 2 to 18 (by lode) 4 x 10 x 5
Mottrams OK 10 4 x 5 x 5
Princess Louise OK 3.9 to 12.5 (by lode) 4 x 5 x 2.5
Rising Tide OK 5 or 10 (by Lode) 10 x 5 x 2.5
Fountain Head OK 4 to 40 (by lode) 5 x 2 x 2.5
Tally Ho OK 10 to 30 (by lode) 10 x 5 x 5
Kazi OK 15,10,5 (by lode) 2 x 10 x 5
Western Arm MIK NA 10 x 20 x 5
Bon’s Rush MIK NA 25 x 100 x 5

TABLE 14-103 MINERAL RESOURCE SUMMARY FOR BURNSIDE AREA

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14.5.1      RISING TIDE

14.5.1.1  Introduction

During November 2011 Geostat Services was commissioned to undertake a mineral resource estimation updating the Rising Tide deposit. The estimation was based mainly on historic drill holes and some holes drilled in 2011 by Crocodile Gold, as detailed in Section 10. Figure 14-66 shows a plan view of the drilling in the Rising Tide area used for this mineral resource estimate. The majority of the drilling undertaken during past exploration and mining was reverse circulation (RC), shown in blue. In addition, some diamond core drilling (DD) was also completed in the area, shown in red.

Figure 14-67 shows a typical cross section through the deposit at Section 10080mE in the local grid system, looking north. It shows several mineral resource interpretations including the most significant, domains 10, 80 and 110. Also shown are the RC and diamond core hole traces defining the mineral resources.

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14.5.1.2  Data Types

The estimation of contained gold has been based on assays sourced from drilling data, detailed in Section 10, above. The data available as at November 2011 consisted of 383 RC and diamond holes. In particular 365 RC holes, 13 RC holes with diamond tailing, and five diamond holes were contained in the database. Most drilling is historic (pre-1998) although most recent drilling includes 21 RC holes and two diamond holes drilled in 2006 by GBS Gold International and 88 RC holes and three diamond holes drilled in 2011 by Crocodile Gold.

All data is in local grid co-ordinates.

Drilling provides data in close proximity to the existing pits area to depths up to 150m below surface. The total database supplied consisted of 21,895m of RC drilling and 930.36m of diamond drilling.

A visual comparison in section between DD and RC data types was completed to assess if any material difference could be observed between data types. Generally, grade continuity between data types was present. Where differences were observed it was not possible to determine whether the grade variability was due to a difference in drilling type or to the short range gold variability observed in the deposit and characterized in the gold variograms. As a consequence, Geostat Services decided to include the diamond data in the estimation to improve data density.

The drill cuttings and core were sampled and assayed mostly at 1 meter intervals, although the database contains intervals at varying lengths within mineralized lodes as summarized in Table 14-104.

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Mineralized
Domain
Minimum Length
(m)
Maximum Length
(m)
10 1 2
20 0.21 2
30 1 1
40 1 2
50 1 1
60 1 1
70 0.2 1.02
80 0.8 2
90 1 1
100 1 1
110 0.21 2
120 0.35 2
130 1 1.1

TABLE 14-104 RISING TIDE DEPOSIT SUMMARY OF SAMPLE LENGTHS BY MINERALIZED DOMAIN

14.5.1.3  Geological Interpretation

The Rising Tide deposit is interpreted to be a series of gently dipping semi-continuous gold mineralized lodes.

The host lithologies of the Rising Tide deposit have been divided into three main states of weathering/oxidation (oxidized, transition, and fresh). Two surfaces interpreted from geological logging (base_ox_2011_tri.dtm and top_fresh_2011_tri.dtm) have been used to define the oxidation state in the deposit. The interpreted oxidization surfaces have not been used as hard boundaries during the estimation of gold. However, they have been used during the assignation of bulk density to the block model.

14.5.1.4  Mineral Resource Interpretation

Interpretation of mineralized domains utilized for this mineral resource estimate, was based upon a lower limit 0.4g/t Au cut-off grade that defines the mineralized veining material.

The resultant estimation domain interpretations were wireframed and nominated 10, 20, 30, 40, 50, 60, 70, 80, 90, 100, 110, 120, and 130. Figure 14-68 below shows the wireframes and drilling traces in an oblique view to the northeast.

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The estimation domain wireframes were used to code the drill intercepts contained within them by flagging into a new table "solids" in the database. This flagging allows the selection of data within domains by codes for the purposes of sample analysis and compositing.

All estimation domain interpretation wireframes have been used as hard boundaries for this November 2011 mineral resource estimate.

14.5.1.5  Compositing and Statistics

Compositing of the raw drilling sample data is necessary to establish a single support for the data to avoid bias when calculating statistics and undertaking any estimation of the data into three dimensional volumes. A number of items are considered when selecting an appropriate composite length; they include the original support of the raw sample data, the assumed selectivity (and therefore the block size) of the model and the imposed spatial dimensions of the interpreted mineral resource estimation domains.

An examination of sample statistics reveals that 95% of sampling within the mineralized domains is on 1m downhole support, although sample lengths vary from a minimum of 0.2m to a maximum of 2m downhole.

Within the mineralized domains the drill samples were composited to 1m downhole to provide equal support data for statistical evaluation and estimation.

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The effect of a small number of outlier composite grades or spatially isolated composites may have an undue effect on the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the composite data in 3D visualization.

A number of high cuts or limits were identified as necessary within the mineral resource estimation domains. A statistical summary of these domains and their corresponding high cuts is shown in Table 14-105 below.


Domain
Minimum
Gold Grade
g/t Au
Maximum
Gold Grade
g/t Au
Mean Gold
Grade
g/t Au
High Grade
Cut
g/t Au
Mean Cut
Gold Grade
g/t Au
10 0.005 29.40 1.75 10 1.70
20 0.010 6.93 1.23 - 1.23
30 0.005 21.60 1.86 10 1.60
40 0.005 7.36 1.18 - 1.18
50 0.140 3.54 1.17 - 1.17
60 0.005 25.70 1.74 10 1.49
70 0.005 10.04 1.04 5 0.96
80 0.005 23.40 1.69 10 1.57
90 0.390 5.06 1.58 - 1.58
100 0.070 4.71 1.79 - 1.79
110 0.010 8.95 1.44 - 1.44
120 0.040 18.60 2.14 10 2.06
130 0.090 3.90 0.68 - 0.68

TABLE 14-105 RISING TIDE DEPOSIT STATISTICAL SUMMARY FOR GOLD IN PPM BY MINERAL RESOURCE ESTIMATION DOMAIN

High-grade gold cuts were applied to the 1m composites within the mineral resource estimation domains and the cut composites used in the estimation. Summary statistics for cut composites are shown in Table 14-106.

Domain Number Cut Mean
g/t Au
Cut
Median
g/t Au
Cut Standard
Deviation
Cut Co-
efficient of
Variation
10 687 1.70 1 1.97 1.16
20 201 1.23 0.75 1.27 1.03
30 44 1.60 0.905 1.91 1.19
40 187 1.18 0.7 1.32 1.12
50 20 1.17 0.71 1.04 0.89
60 108 1.49 0.83 1.87 1.26
70 95 0.96 0.65 1.07 1.11
80 295 1.57 0.98 1.84 1.17
90 13 1.58 0.875 1.45 0.92
100 26 1.79 1.53 1.33 0.74

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Domain Number Cut Mean
g/t Au
Cut
Median
g/t Au
Cut Standard
Deviation
Cut Co-
efficient of
Variation
110 305 1.44 0.975 1.44 1.00
120 191 2.06 0.99 2.57 1.25
130 23 0.68 0.42 0.82 1.21

TABLE 14-106 RISING TIDE DEPOSIT HIGH- GRADE COMPOSITE CUTS FOR GOLD IN G/T BY MINERAL RESOURCE ESTIMATION DOMAIN

The general statistics of gold composites within all domains can be described as positively skewed with moderate to high variability as is the case with most gold occurrences. The high variability is reduced somewhat by high cutting of gold grades in those domains most affected.

The co-efficient of variation, which is a measure of variability, remains at an approximate value of one for most of the domains, indicating variability was reduced within the domains after high-grade cutting.

14.5.1.6  Variography

Variography was used to characterize the spatial behavior of the composite data primarily as an aid to establishing estimation parameters. Variogram stability and quality is dependent upon the statistical properties and the amount of data available within the defined domains. After an initial investigation of the gold data, one variogram model was defined for all mineralized zones. The general variogram model is detailed in Table 14-107 below.

Domain Nugget Stuct Sill Major
(m)
Semi
(m)
Minor
(m)
Major/
Semi
Major/
Minor
Surpac Rotation
Bearing Plunge Dip
All Mineralized
Lodes
0.67 St1 0.12 30 30 7 1 4.29 69 -3 -20
St2 0.2 200 100 8 2 25 69 -3 -20

TABLE 14- 107 RISING TIDE DEPOSIT VARIOGRAM MODELS FOR GOLD BY MINERALIZED DOMAIN

The features of the variogram model for gold can be summarized as moderately high relative nugget of about 70% for the mineralized domains with a significant amount of variability demonstrated over a short range. This reflects the high variability at short range observed in visual inspection by section. Maximum range extends to 200m.

14.5.1.7  Grade Interpolation Methodology

A standard three dimensional two pass Ordinary Kriging methodology has been used for the estimation of the cut gold 1m down hole composite data within each estimation domain. Table 14-108 summarizes the estimation parameters by domain. A constant minimum of four and maximum of 25 composites have been set for all domains on the first pass whilst a minimum of two and a maximum of 25 composites were used for all domains on the second pass. A discretization array of 5 (X) by 5 (Y) by 2 (RL) was used to refine the Kriging weights for each model block. In addition, the search orientation was adjusted to the specific orientation of each estimation domain.

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Domain Pass Search Radius
(m)
Bearing Plunge Dip Major
/Semi
Major/
Minor
All Mineralized Domains 1st 60 69/95 -3 -20/-30 2 6
2nd 90 69/95 -3 -20/-30 2 6

TABLE 14- 108 RISING TIDE DEPOSIT ESTIMATION PARAMETERS FOR GOLD BY ESTIMATION DOMAIN

14.5.1.8  Block Model Definitions

The primary consideration of the 3D model was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each estimation domain. All individual estimation domains were ultimately combined to create a single block model in the local grid coordinate system. Table 14-109 summarizes the 3D block model “RisingTide_Nov2011.mdl” definition.

  Northing Easting RL
Minimum 3680 9500 990
Maximum 4300 10350 1170
Block Size 5 10 2.5
Sub-block 2.5 5 1.25

TABLE 14-109 RISING TIDE DEPOSIT 3D BLOCK MODEL DEFINITION (M)

The chosen block size represents approximately half the best data spacing in the Northing and Easting directions and a choice in the vertical dimension controlled by the need to appropriately represent the volume of the wireframes that define the estimation domains.

A standard list of field names and descriptions used in the block model are shown in Table 14-110.

Attribute Type Default Description
au Float -99 Au Grade
au_oz Calculated - (_xext*_yext*_zext*density*au)/31.10347
density Float -99 SG
krig_var Float -99 Kriging Variance
lodecode Integer -99 Lode code
mined Integer -99 1=yes, 2=no, 0=air
no_samp Integer -99 No of samples to inform a block
res_class Integer -99 Classification code

TABLE 14-110 RISING TIDE DEPOSIT 3D BLOCK MODEL ATTRI BUTES

Table 14-111 below confirms the close agreement of the 3D block model volumes and the original interpreted wireframe volumes, supporting the 3D model block size choice as appropriate. The total volume for the estimation domains in the block model stays within three percent from the wireframe volumes defining such domains.

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Domain Wireframe
Volume
Block Model
Volume
Variance %
10 377,338 364,141 96.5
20 91,473 88,359 96.6
30 35,807 35,516 99.2
40 143,627 143,891 100.2
50 8,619 9,156 106.2
60 61,960 61,516 99.3
70 37,728 36,813 97.6
80 189,322 184,859 97.6
90 9,180 9,188 100.1
100 16,406 16,234 99
110 141,627 140,359 99.1
120 72,812 68,344 93.9
130 7,960 7,734 97.2
Total 1,193,859 1,166,110 97.7

TABLE 14-111 RISING TIDE DEPOSIT 3D BLOCK MODEL TO WIREFRAME VOLUMES CHECK

14.5.1.9  SPECIFIC GRAVITY/ BULK DENSITY ASSIGNMENT

The bulk density of the waste and mineralized rock of the final 3D block model has been assigned according to oxidation state, using the interpreted surfaces described in Section 14.2.4 to control the blocks assigned. Crocodile Gold has determined the specific gravity based on 811 measurements made by Crocodile Gold on RC drill samples using the Vacuum Pyncnometer method. The data analysis excluded historic data measured mainly by the water invasion method (293 samples) because of its higher variability and incompatibility with current geologic domains. Table 14-112 summarizes the oxidation state bulk density assignations.

Oxidation State SG
g/cm3
Oxidized 2.79
Transition - Non Mineralized 2.78
Fresh - Non Mineralized 2.96
Transition/Fresh - Mineralized 3.11

TABLE 14-112 RISING TIDE DEPOSIT SPECIFIC GRAVITY VALUES BY OXIDATION STATE

Within the final 3D block model, blocks within the mineralized domains coded fresh or transitional were assigned a value of 3.11g/cm 3 in the “density” field; those outside the mineralized domains coded fresh were assigned a value of 2.96g/cm 3 whilst those coded transition were assigned a value of 2.78g/cm 3; finally all blocks within the oxidized zone and below the topographical surface were assigned a specific gravity of 2.79g/cm 3. Blocks located above the topographical surface were assigned a zero specific gravity.

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14.5.1.10  Model Validation

Model validation has been undertaken to ensure no material error has been made in the estimation of the Rising Tide November 2011 mineral resource estimate. The validations include inspection of the audit documentation of the individual estimation runs; visual inspection of the block outcomes and input data; statistical comparisons of input data and block outcomes, and swath plots of each of the domains.

Statistical comparisons of input data and block model outcomes for the estimation domains are shown in Table 14-113.

Domain Cut Composite
Average Grade
g/t Au
Block Model
Average Grade
g/t Au
Variance
%
Percentage of
Total Gold Oz
Indicated Model
Average Grade
g/t Au
Variance
%
10 1.70 1.67 98 35.3 1.64 97
20 1.23 1.22 99 6.2 1.23 100
30 1.60 1.57 98 3.4 1.61 100
40 1.18 1.16 98 10.1 1.17 99
50 1.17 1.05 90 0.6 1.05 90
60 1.49 1.44 97 5.4 1.44 97
70 0.96 1.11 115 2.5 1.11 115
80 1.57 1.29 82 14.3 1.35 86
90 1.58 1.56 99 0.8 1.62 103
100 1.79 1.81 101 1.8 1.82 102
110 1.44 1.51 105 12.3 1.5 105
120 2.06 1.95 95 7.2 1.97 96
130 0.68 0.64 95 0.3 0.64 95

TABLE 14-113 RISING TIDE DEPOSIT MINERALIZED DOMAIN A VERAGE GOLD GRADE COMPARISONS

The estimation domain comparisons display a reasonable variation between input and outcome average grades when the total domain is reported. As can be confirmed in the visual inspection and swath plot investigations, the comparisons include small volumes in border areas of some domains containing a lower density of sample data. This results in extrapolation of the sample data into these volumes and while it is considered a reasonable estimate of the grades within these volumes, a simple statistical comparison of total volumes will not result in close comparisons for all cases. Two views of the block model for domain 10 (the most significant one in terms of contained metal) are shown to demonstrate this situation. Figure 14-69 shows composite data and blocks colored by gold grade and Figure 14-70 shows blocks colored by mineral resource classification and composites colored by gold grade. The contrasting data densities at depths are evident, supporting the assumption that relying only on raw composite to block grade comparisons can be misleading.

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Comparison of the Indicated portions of the estimation domains shows that the comparison to average composite grades agrees within a 15% tolerance.

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Data is analyzed by northing and easting for each domain. Reproduced below are some example swath plots for domain 10 (Figure 14-71) and domain 80 (Figure 14-72).

The swath plots demonstrate that where there is regularly spaced data the block model reflects that data. The plots also highlights that a paucity of drilling results in parts of the model that rely on only a few measured points are therefore less likely to match local composite data, and are of reduced certainty and increased risk.

14.5.1.11  Mineral Resource Classification

The Rising Tide mineral resource estimate has been classified in accordance with the CIM guidelines and National Instrument NI 43-101. This classification was based upon information provided by Crocodile Gold and outcomes of the estimation processes undertaken by Geostat Services. Assessment criteria include data integrity, drillhole spacing, sample locations, sampling density, and lode geometry, geological confidence and grade continuity. Consideration has been given to the estimation technique and the risks associated with extrapolation of sample data.

The mineral resource has been classified as Indicated and Inferred; no Measured mineral resources have been identified.

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14.5.1.12  Data Spacing and Distribution

The Rising Tide model has been shown in validation to be subject to varying drillhole density and sample location in relation to the lode geometry. In most lodes the drilling is regular and of sufficient density but subject to decreasing densities in border areas. The block model outcomes in low density areas are considered to be higher risk and are classified with less confidence than denser parts.

14.5.1.13  Orientation of Data in Relation to Geological Structure

The orientation of the deposit is interpreted to be close to horizontal and the drilling is considered to be appropriately targeted for this geological orientation.

14.5.1.14  Deposit Dimensions

The mineralized portion of the Rising Tide deposit extends within drill testing from 3,725 to 4,225m in Northing with extension beyond drilling of up to 12.5m; within the Easting plane the deposit extends within drill testing from 9,550 to 10,325ms with extension beyond drilling of up to 12.5m; in the vertical the dimensions of the mineralization are tightly constrained by drilling extending from surface (at approximately 1155mRL) to 1015mRL. The dimensions of the mineralization are adequately defined by the available drilling with limited and acceptable extensions beyond data.

14.5.1.15  Estimation and Modelling Techniques

The estimation methodology used for the mineralization style is considered appropriate by the QP’s based on experience with similar deposit types. It is shown to represent reasonable unbiased reproductions of the input data in areas of adequate sampling. Outside areas of adequate sampling the resource classification is such as to reflect the uncertainty of the estimate. The validation methods used also demonstrate the adequacy of the methodology used.

14.5.1.16  Moisture

The estimate has been made on the basis of dry tonnes.

14.5.1.17  Classification

All material within the mineral resource interpretation has been classified to represent the QPs’ opinion of the risk in the mineral resource estimated. Within the mineralized estimation domains that have been defined on a plus 0.4g/t gold cut off, it is assumed that some of the material will form dilution to the mining of higher grade material. For reporting purposes the mineralized material has been reported with a lower cut off of 0.7g/t Au within the interpreted wireframes. The classification of the Rising Tide mineral resource into Indicated and Inferred as set out below reflects the Company’s view of this deposit, as it is currently defined.

14.5.1.18  Selectivity Assumptions

The resource estimate contains implicit assumptions of mining selectivity represented by the block size of 5m x 10m x 2.5m (Y x X x Z).

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14.5.1.19  Discussion of Relative Accuracy / Confidence

At this stage no quantitative testing of the accuracy of the estimate or establishment of confidence limits has been undertaken.

14.5.1.20  Recommendations

The Rising Tide deposit is considered to be defined sufficiently well with a high confidence in the mineral resource model.

However, in order to improve the quality of the estimated mineral resource, it is recommended that current bulk density assignations be confirmed with measurements on core samples that would take into account the natural porosity of the rock

In addition, the re-assay of selected samples included in the historic portion of the database (prior 1998) is also recommended to compensate for the lack of QA/QC results for that period.

14.5.2      MINERAL RESOURCES GENERATED PRIOR TO NEWMARKET GOLD

Crocodile Gold previously reported on several other deposits in the Burnside area. These will be reported without change in this update as no new data has been capture and no updated models have been completed.

Below is an extract from the 2013 technical report on mineral resources and mineral reserves. This summarizes the methodology of the past mineral resources previously reported.

14.5.2.1  Western Arm

Introduction

In 2012 Crocodile Gold requested Cube Consulting (“Cube”) to undertake an audit of the current resource estimate for the Western Arm deposit, located in the Howley Field, of the Pine Creek Geosyncline, NT. Cube has conducted a review of the available reports, the digital database, the estimation data, the mineral resource model and grade tonnage reports pertaining to the Northern Gold NL (NGNL) March 2001 Resource estimate (Hardy and Hague 2001a). All data has been sourced from Crocodile Gold and a complete file listing is attached for reference.

This review is desktop in nature and focused primarily on identifying material issues pertaining to the current reported mineral resource (NGNL 2001).

The database available can be confirmed as the one used by NGNL in 2001. It contains digital summary geological and oxidation logging and a gold value with lab repeats and associated sample number.

The drillhole database was extensively reviewed and verified by Hardy and Hague (2001a) (Hardy and Hague 2001a).

Details of the NGNL 2001 block model are documented in Hardy and Hague 2001 (Hardy and Hague 2001a). Hardy and Hague (2001) (Hardy and Hague 2001a) used MIK to estimate gold grade for Western

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Arm, inclusive of a block support correction to 5mE x 8mN x 2.5mRL SMU’s using the lognormal-normal method. They further subdivided the four aforementioned domains into Oxide and Fresh portions for gold grade estimation, resulting in a total of 8 distinct estimation domains.

Cube has made the following observations, which are supported by the Author, with respect to the parameters and methodology used by Hardy and Hague (2001) (Hardy and Hague 2001a) for the geostatistical gold grade estimation:

The mineral resource summary provided by Hardy and Hague (2001) (Hardy and Hague 2001a) was compared to a block model report generated by Cube. Cube concludes that the block model file provided reflects the result of Hardy and Hague’s (2001) (Hardy and Hague 2001a) study.

Cube reviewed the block model visually and using swath plot comparison to composite data. The swath plot comparison of the block model e-type grades to composite grades were made (see Figure 14-73 below) and the comparison is reasonable, although the estimate is consistently higher by a small margin than the composite data. The composite file used for this comparison includes all composite data – both mineralized and un-mineralized. This indicates that the block model has the un-mineralized domains estimated and diluted into the panel grades (e-type grades). Cube is not clear exactly how this has been done from the documents reviewed. The result is however that the e-type grades are on large support and unsuitable for any meaningful mining evaluation studies. The MIK results with their change of support assumptions should be used for this purpose.

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Conclusions and Recommendations

Cube suggested and the Author agrees with the primary recommendations made by Hardy and Hague (2001a) (Hardy and Hague 2001a), which are as follows:

The work undertaken by Hardy and Hague in 2001 is a diligent and thorough study with a significant effort put into the validation of the data and analysis of the available sample QA/QC. The modeling approach is reasonable and diligent as far as Cube can tell. There appears to be no fatal flaws in the mineral resource estimation methods applied by Hardy and Hague (Hardy and Hague 2001a), however, there are a number of questions, which cannot be answered around the details of some of the parameters that were used. A number of parameters (e.g. search neighborhood rotations) appear to be sub-optimal and could, in Cube’s opinion, be improved upon. It is possible, using modern geostatistical methods, to better deal with the folded geometry of the deposit. Finally, Cube is uncertain as to how the component parts of the geology and gold grade estimates were utilized and combined.

A significant proportion of the 2001 mineral resource model has been classified class 1 – equivalent to Indicated. With the supporting data available (using the work done by Hardy and Hague and the sources detailed in Hardy and Hague) and after a check estimate the class 1 portions of the mineral resource would in all likelihood be classified as Indicated.

Cube understood that a significant quantity of original supporting data is available in hard copy form. It is recommended that this information be collated, reviewed and digitally data based. This would have a number of benefits, firstly, it preserves the work undertaken so far which has some considerable value to the company; secondly this facilitates the QP to ascertain if additional QA/QC data has been overlooked and if further QA/QC data is required. This has been completed during 2015 with the Author confident

in the data matching from hard copies into the database. This process will continue before a model update is completed.

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Cube recommends, and the Author agrees that, a thorough search is conducted for any remaining core from this deposit. If the core can be located and catalogued it can be used for additional confirmation sampling and assaying and if of suitable quality facilitate additional density determinations. This work has commenced.

It is Cube’s opinion, to which the Author agrees that, a re-estimation of the Western Arm model would be of benefit in providing confirmation of the existing model, by comparison to an alternate modeling method. Such a re-estimate would benefit greatly from additional bulk density readings, especially in the oxide zone.

14.5.2.2  Bon’s Rush

Introduction

In 2012 Crocodile Gold requested Cube Consulting (“Cube”) to undertake an audit of the current mineral resource estimate for the Bon’s Rush deposit, located approximately 28km east of Adelaide River in the Northern Territory, Australia. Cube has conducted a review of the available reports, the digital database, the estimation data, mineral resource model and grade tonnage reports pertaining to the Northern Gold NL (NGNL) June 2001 Resource estimate (Hardy and Hague 2001b). All data has been sourced from the Crocgold data room.

This review is desktop in nature and focused primarily on identifying material issues pertaining to the current reported mineral resource (NGNL 2001).

The database available can be confirmed as the one used by NGNL in 2001. It contains digital summary geological and oxidation logging and a gold value with lab repeats and associated sample number.

The drillhole database for Bon’s Rush deposit has been extensively reviewed and verified by Hardy and Hague (2001) (Hardy and Hague 2001b). During a site visit in 2015 the Author was able to identify the drilling collars to confirm their location in relation to the database.

Hardy and Hague (2001b) (Hardy and Hague 2001b) used MIK to estimate probabilities for two domains – these two domains were based on a subdivision of lode and non-lode assays. The lode portion consists of three wireframed domains trending north-northeast and dipping steeply to the northwest. The cut-off grade or other possible criteria used to model these wireframes are unclear, the lower cut-off used appears to be low and Cube would interpret this to indicate geology logging attributes were also used to define the mineralization. It is unclear from the documentation how the estimation of the domain probabilities was used in the subsequent estimation of gold grade.

Hardy and Hague (2001) (Hardy and Hague 2001b) used Multiple Indicator Kriging (MIK) to estimate gold grade for Bon’s Rush, inclusive of a block support correction to 5mE x 5mN x 3mRL SMU’s using the lognormal- normal method. They further subdivided the four domains into Oxide and Fresh portions for gold grade estimation, resulting in a total of 8 distinct estimation domains.

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Cube has made the following observations with respect to the parameters and methodology used by Hardy and Hague (2001) (Hardy and Hague 2001b) for the geostatistical gold grade estimation:

Hardy and Hague (2001b) (Hardy and Hague 2001b) have categorized the entire mineral resource model as Inferred due to the lack of quality control data, wide data spacing and inadequate bulk density data. Cube agrees with this classification and the reasons given for it.

Conclusions and Recommendations

Cube suggested, and the Author agrees with, the primary recommendations made by Hardy and Hague (2001) (Hardy and Hague 2001b), which are as follows:

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The work undertaken by Hardy and Hague in 2001 is a diligent and thorough study with a significant effort put into the validation of the data and analysis of the available sample QA/QC. The modeling approach is reasonable and diligent as far as Cube can tell. There appear to be no fatal flaws in the mineral resource estimation methods applied by Hardy and Hague (2001) (Hardy and Hague 2001b). However, there are a number of questions, which cannot be answered, including exact details of the parameters that were used. Some of the parameters (e.g. search neighborhoods) appear to be sub-optimal and could, in Cube’s opinion, be improved upon. In addition, Cube is uncertain as to exactly how the component parts of the geology and gold grade estimates were utilized and combined.

The Auther agrees with Cube in that a re-estimation of the mineral resource model using an alternate method to MIK would be of benefit in providing an independent check on the existing model outcome.

With a summary of the supporting information (drilling, surveying and QA/QC) that is available and a check estimation by an alternate method, Cube see no obstacles to the reporting of this mineral resource in accordance with the CIM guidelines, in 43-101 format.

14.5.2.3  Kazi

Introduction

In 2012 Crocodile Gold requested Cube Consulting (“Cube”) to undertake an audit of the current mineral resource estimate for the Kazi deposit, located approximately 160km south of Darwin in the Northern Territory, Australia. Cube has conducted a review of the available reports, the digital database, the estimation data and the mineral resource model and grade tonnage reports pertaining to the Burnside Operations Ltd (BOL) March 2005 Mineral resource estimate, reported by Harris and Dyer in March 2005 (Harris 2005). All data has been sourced from the Crocgold data.

This review is desktop in nature and focused primarily on identifying material issues pertaining to the current reported mineral resource.

The database available can be confirmed as the one used in 2005. It contains digital summary geological and oxidation logging and a single gold value.

Drill lines are oriented roughly east-west with a hole spacing along the lines of 20m, with the lines themselves spaced 20 to 40m in the north-south direction.

The 2005 report appears to be incomplete as it states the mineralization wireframes are subdivided into 4 broad lodes due to reverse faulting, (Lodes; rt100, rt200, rt300 & rt400) and is based upon a 0.4g/t Au cut-off for better continuity. In fact, there are 12 lodes interpreted with apparent consistency in the cutoff used for the interpretations for the 12 lodes. The cut-offs are quite variable with 0.1, 0.3, 0.4, 0.9 & 1.0g/t Au cut-offs used.

Cube also notes that a number of significant assay intervals are not included in mineralization wireframes.

There are 12 lodes named; 10, 20, 30, 40, 50, 60, 70, 80, 90, 100, 110, 120.

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Cube has replicated the basic statistics documented in the supplied report by coding a database using the supplied lode wireframes. As the drill holes and wireframes did not exactly match, there are some minor immaterial differences.

Top cuts have been applied to individual lodes;

The selection of top cuts by Harris and Dyer are reasonable for the individual lodes modeled. Sensitivity analysis by Cube of various top-cut scenarios indicate the choice of top-cuts is appropriate in reducing the overestimation of metal produced from higher grade spatial outliers in the composited data.

Estimation of block grades was by Ordinary Block Kriging of composite data into individual wireframed domains.

The block model dimensions are summarized in Table 14-116 below.

Model Limits Extent of Model No of Blocks Block Size
68800N – 69150N 350m 35 10m
46150E – 46400E 250m 125 2m
-120mRL – 70mRL 190m 38 5m

TABLE 14-114 BLOCK MODEL SUMMARY FOR KAZI DEPOSIT

Sub-blocking was to a minimum of 5m north, 1m east and 2.5m in RL. The interpolation used the following parameters:

  1.

A minimum of four composites - maximum of 10 composites per block estimate;

  2.

A initial search of 60m x 40m x 10m with a second pass expanded search (lodes 100&110 doubled, all other expanded by 1/3);

  3.

Ordinary Kriging was used for the estimation.

  4.

Cube sees no material issues with the estimation parameters adopted for the estimate.

Block model validation graphs have been produced by Cube for domains 90, 100, 110 and 120. The graphs do not indicate any material over or under estimation in the grades, and the mean grade block estimates are close to the mean composite grades. This is the result of the small search neighborhood used for the estimation and the regular spaced (non-clustered) data.

These graphs are presented in Figure 14-74 to Figure 14-77 below.

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The visual/graphical validation of the estimate compared with the Cube generated composite data shown above looks sound overall. There is some smoothing evident which is expected from Ordinary Kriging. The estimation in the wider spaced data regions (greater than 20m x 20m spaced drilling) that are proximal to higher grade composite data may represent a higher risk in overcalling contained metal; however, these areas have been classified as Inferred Resources by Harris and Dyer.

Conclusions and Recommendations

During the review completed by Cube no information regarding drilling methodology, sampling procedures or recoveries was available with the supplied data. This information or at least a commentary on the lack of it is essential for reporting.

Since the Cube review was completed a more detailed set of information has been sourced and provided to the Auther for review. This information outlines the drilling methodology and available sampling QA/QC data. This information has been reviewed and the commentary is below;

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Western Mining Corporation (WMC) conducted a number of drilling programs within the Kazi area during their tenure of the lease (1988 -1993) to test the gold bearing quartz-arsenopyrite zone. A Warman 1000 rig was used to drill both HQ and NQ diamond holes during their 1988/89 drilling season, which saw a total of six diamond holes drilled for 595.6 m. Only five of these holes had percussion precollars. If core was found to be mineralized, a half core sample would be taken, which would then be coarsely crushed to 1-3mm, to produce a 100-200g sub-sample. The samples were then sent to WMC laboratories in Kalgoorlie for pulverization prior to aqua-regia digest with an AAS finish. Recent efforts to locate a more detailed description for the sampling, splitting and gold analysis techniques have been unsuccessful. A similar method was used for sampling non-mineralized samples. Chip samples were collected every 10-20cm from the core section, over 1m intervals. This sample would then be coarsely crushed to 1-3mm, to create a 200g sample which would then be sent to the WMC laboratories in Kalgoorie. The sample would then undergo pulverization, prior to a 25g charge, aqua-regia digest and an AAS finish. Gold was reported in parts per million (ppm) with a detection limit of 0.02ppm Au.

WMC conducted an exploration and grade control reverse circulation drilling program, between 1988 and 1991. The programs consisted of 79 drill holes for a total of 2,475m drilled on the lease. It is believed a “T-3” drill rig was used for the program (Hardy and Hague 2001c), using a cyclone and a two tier riffle splitter. Little is known on the specific details of the rig drill size, sample techniques or sample recoveries. A 2-5kg sample was collected, which was then crushed down to 1-3mm. A 100-200g sample was produced which was then sent to WMC laboratories. The samples were then pulverized down for a 25g sample to be tested using aqua-regia digest with an AAS gold determination.

It is unclear if WMC conducted any quality control measures on any of the samples obtained during their drilling programs.

WMC drilled their holes on the Paqualin grid, which was then transformed onto the Northern Gold NL (NGNL) grid. Downhole survey points were taken, presumably by a downhole camera and it is unknown if the collar coordinates were also surveyed.

In 1993, Northern Gold NL gained ownership of the lease, which saw a number of drilling campaigns being undertaken during 1994 and 1996. This included 119 reverse circulation percussion (RCP) holes for 10,973m and five diamond holes for 1,037.5m.

Diamond drilling was undertaken by Gaden Drilling, using a Universal 650 drill rig, which drilled HQ3, NQ and NQ2 core. Core was sampled through the use of a diamond saw, which provided half a core sample for gold assaying at Assaycorp (Pine Creek). A FA50 fire assay was used for gold analysis.

Holes were geologically logged and had both downhole survey points and surveyd collar positions.

NGNL contracted out their RCP drilling to Tennant Creek drilling company Gomex Drilling. Both a truck mounted Rotmac 50 and a RCD150 drill rig was used for the drill campaigns which used a 4½ inch face sampling hammer and a 4:1 cascade rig mounted riffle splitter for sample extraction. One meter samples would be collected into two calico bags, with one bag used for assaying, while the other remained on site as reference material. Samples would be sent to Assaycorp in Pine Creek for a 50g fire assay to test for gold. Geological logs, downhole surveys and surveyed collar data are available for the holes. Field duplicates were also collected and sent to Amdel (Darwin) for gold analysis, using the FA, 50g fire assay methology. Results from the internal QC programme of laboratory duplicates are seen in Figure 14-78 and Figure 14-79 below.

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This additional information gives the Author confidence that the drilling was completed to a standard that is appropriate for reporting mineral resources and therefore this mineral resource is suitable for inclusion in this technical report.

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It was found that there was a normal distribution associated with the Q Q plot from the assay data.

A summary of both exploration and mineral resource development drilling undertaken at Kazi deposit from both WMC and NGNL is given in Table 14-117 below:

 Drillhole Summary for the Kazi Project Area
Company Sample Type No. Holes Total Meters (m)
WMC DDH 5 595.6
RCP 79 2,475
NGNL DDH 5 1,037.50
RCP 119 10,973
  TOTAL 208 15,081.10

TABLE 14-115 DRILLHOLE SUMMARY FOR KAZI 1989 -1996

It is the opinion of the Author that the sampling is of sufficient quality that the use of an Inferred mineral resource is appropriate.

Cube has assessed the grade interpolation outcome of the mineral resource estimate and found it to be a sound global estimate of contained metal assuming the interpreted domains and data utilized are correct. There are, however, some issues regarding the interpretation and classification of the mineral resource.

Cube recommends, and the Aurther agrees that, the following steps be undertaken to bring the Kazi estimate to an acceptable standard for reporting.:

A broader wireframe/envelope definition would be favored by Cube at this stage, as the current interpretation assumes a high degree of continuity between high-grade intercepts, which often results in a conditionally biased estimate. Some "lode switching" is evident in the interpretation, where the interpretation crosses natural trends to obtain higher grade intersections.

14.5.2.4  Howley Deposit

The Howley deposit mineral resource estimate was completed in early 2011 after significant amounts of drilling conducted by both GBS Gold and Crocodile Gold. This drilling was used to up-date the mineral resource. A summary of the work included is outlined below;

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Domaining

Since the Howley deposit is interpreted to be a complex overturned fold, with mineralization sub-parallel to the dolerite units within this fold complex, consideration was given to domaining of the deposit as to enable representative variography analysis of the respective lodes.

The Howley deposit was divided into 11 broad zones, on the basis of overall lode dips. Not all 11 zones are present in this illustration, with other domains present in other areas of the Howley deposit (e.g. Big Howley deposit to the north).

Block Model Creation and Extents

A 3D block model, howley_feb2011.mdl was generated using Surpac software with origin, extents and attributes defined below in Table 14 85. Parent blocks of 10m x 4m x 5m size (Y x X x Z) were subdivided into sub-blocks of 5m x 1m x 2.5m in order to fill areas adjacent to wireframe boundaries. The wireframes were used to limit the blocks available for grade interpolation, with block centroid locations used to define the blocks and sub-blocks for interpolation.

Since the Big Howley area of the Howley deposit is located 500m north of the main Howley deposit, it was decided to create a separate Surpac block model for this area, bighowley_feb2011.mdl. Table 14 118 tabulates the origin and extents of this area.

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Estimation Parameters and Methodology

Ordinary Kriging using parameters derived from the back-transformed log variograms was used to estimate Au grades for the Howley deposit. The skewed nature of the data distribution makes this technique ideal, whereas other techniques such as inverse distance interpolation assume a normal distribution, which can lead to errors if the data is not cut appropriately. Inverse distance techniques also do not utilize the information obtained from the variogram in interpolation of blocks, and thus the spatial correlation between composites is not taken into account.

Grade Interpolation

Each lode was treated as a separate hard boundary, restricting the grade interpolation to drillhole data located within each lode. A minimum of 4 samples and a maximum of 25 samples were used to interpolate grades into each block. A discretization array of 5 (north) by 2 (east) by 5 (RL) was used to refine the Kriging weights for each model block.

Two interpolation passes were conducted for all lodes. Search ellipse parameters were guided by variography ranges, with 40m x 20m x 10m used for all lodes in the initial pass of interpolation. For the second interpolation pass, search ellipses for all lodes were expanded to 80m x 40m x 20m. Only the blocks not filled by the first pass were interpolated by the second pass, and grades estimated from the first interpolation pass were left unchanged. All non-filled blocks within the lodes after these interpolation passes were left un-estimated and will not be included in the resource.

Model Validation

The Howley block model was validated by several methods, including visual validations on screen, global statistical comparisons of input and block grades, and local grade/depth relationships. The model was validated visually by viewing vertical sections and plans with spatial comparison of interpolated block grades against input composite grades to ensure grade trends were represented correctly.

Grade/Northing Validations

Figure 14-80 illustrates the Au grade/northing relationship for the Howley deposit. Both input composite data and model grade data for all lodes were averaged within 20m northing increments, and plotted together with the number of composites to assess the reliability of the block model.

Comparisons of model grades with composite grades illustrate a good reconciliation, with model grades reproducing the broad trends in composite grades. Deviations between the model and composite grades occur in areas of low data density, illustrating the need for more composites in these areas.

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Mining Reconciliation

Mining continued from the commencement of operations for Crocodile Gold until April 2012 when the small open pit called Howley West was completed. During that time a total of just over 2 million tonnes of ore (greater than 0.7 g/t Au) was mined at an average grade of 1.26 g/t Au for 84,000 ounces. During the same period the mineral resource estimate predicted a mined tonnes and grade of 1.9 million tonnes at 1.40 g/t Au for 87,500 ounces. This suggests the estimate predicted 6% less tonnes at 10% more grade for 4% less ounces. Some of this difference would have to be explained by the net effect of dilution and ore loss. It is therefore understood that the Howley mineral resource estimate predicts the tonnes and grade of the deposit well and can be used in the future with confidence.

14.5.2.5  Mottrams Deposit

Domaining

Statistics were run within the drillhole database for all constrained uncut composite data by lode and are presented in Table 14-119. No other mineralization indicators were used, as data was extracted from within wireframes.

Log histograms and lognormal probability plots of lode domains are presented below.

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Grade populations were analyzed statistically by lode, and top-cuts applied to each lode together with cut lode statistics. Some lodes, being of predominant low-grade did not necessitate the application of a top-cut.

Lode 100 200 300 400
Number 457 550 1176 959
Minimum 0.01 0.01 0.005 0.005
Maximum 18.3 8.66 15.2 12
Mean 1.17 1.18 1.29 1.15
Median 0.73 0.77 0.81 0.79
Std Dev 1.59 1.27 1.48 1.21
Variance 2.54 1.61 2.19 1.47
Coeff Var 1.36 1.08 1.15 1.06

TABLE 14- 117 MOTTRAM DEPOSIT - SUMMARY STATISTICS

Block Model Creation and Extents

A 3D block model, mottrams_nov2010.mdl was generated using Surpac software with origin, extents and attributes defined below. Parent blocks of 5m x 4m x 5m size (Y x X x Z) were subdivided into sub-blocks of 5m x 1m x 2.5m (along-strike x across-strike x RL) in order to fill areas adjacent to wireframe boundaries. The solid wireframes were used to limit the blocks available for grade interpolation, with block centroid locations used to define the blocks and sub-blocks for interpolation.

Parameter Y X Z
Minimum Coordinates 5200 5150 1000
Maximum Coordinates 6200 5650 1150
User Block Size 5 4 5

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Parameter Y X Z
Min. Block Size 5 1 2.5

TABLE 14-118 BLOCK MODEL SET UP AND DIMENSIONS

Attribute
Name
Type
Decimals
Background
Description
au_cut Float 2 -99 estimated Au value with top cut
au_oz Calculated - - Au ounces per block
density Float 2 -99 SG value
krig_var Float 3 -99 Kriging variance
lodecode Integer - -99 Wireframe lode code
no_samp Integer - -99 No of samples used for block estimation

TABLE 14-119 MOTTRAMS DEPOSIT - BLOCK MODEL SUMMARY

Estimation Parameters and Methodology

Ordinary Kriging using parameters derived from the back-transformed lognormal variograms was used to estimate gold grades for the Mottrams deposit.

Grade Interpolation

Each lode was treated as a separate hard boundary, restricting the Au grade interpolation to drillhole data located within each lode.

Two interpolation passes were conducted for all lodes, with a search ellipse of 25m x 20m x 5m used in the initial pass of interpolation. For the second interpolation pass, search ellipses were expanded by 50%, with only those blocks unfilled by the first pass interpolated by the second pass, and grades estimated from the first interpolation pass were left unchanged.

Model Validation

The Mottrams deposit block model was validated by several methods, including the following:

Table 14-122 illustrates the reconciliation of global average input composite grades with mean block grades by lode. All lodes show a robust reconciliation between composites and model grades, with the only deviation occurring in areas of low drill density.

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Rockcode No of Average Model Average Cut % Difference
100 457 1.20 1.14 6%
200 550 1.19 1.18 1%
300 1176 1.30 1.29 1%
400 959 1.13 1.15 -1%

TABLE 14-120 MOTTRAMS DEPOSIT GLOBAL VALIDATION, G/T AU

Figure 14-85 illustrates the Au grade/Northing relationship for the Mottrams deposit. Both input composite data and model grade data for all lodes were averaged within 10m northing increments, and plotted together with the number of composites to assess the reliability of the block model.

Grade/Northing Validations

Comparisons of model grades with composite grades illustrate an excellent reconciliation, with model grades reproducing the broad trends in composite grades. Deviations between the model and composite grades occur in areas of low data density, illustrating the need for more composites in these areas.

Mining Reconciliation

Mining commenced at the Mottrams deposit in December 2010 with minimal material moved prior to January 1st 2011. Mining was completed on the pit in December 2011 with a total of just under 1.3 million tonnes of material reconciled and processed at the Union Reefs mill at an average grade of 1.09 g/t Au for 44,300 ounces. This is above a 0.7 g/t Au lower cut-off. Reconciling this against the model reported in 2011 shows the estimated tonnes for the period were around 1.15 million tonnes at an average grade of 1.16 g/t Au, this suggested the mining was around 10% higher in tonnes and 6% lower in grade. This is most likely the net effect of dilution, this would need to be confirmed if more work is required on this deposit.

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14.5.2.6  North Point

Domaining

Compositing of the raw drilling sample data was necessary to establish a single support for the data to avoid bias when calculating statistics and undertaking any estimation of the data into 3 dimensional volumes. A number of items were considered when selecting an appropriate composite length; they include the original support of the raw sample data, the assumed selectivity (and therefore the block size) of the model and the imposed spatial dimensions of the interpreted mineral resource estimation domains.

An examination of sample statistics revealed that 93% of sampling within the mineralized domains is on 1 meter downhole support, with around 5% of composites around 1.5m, although sample lengths vary from a minimum of 0.1 to a maximum of 5 meters downhole.

Within the mineralized domains the drill samples were composited to 1m downhole to provide equal support data for statistical evaluation and estimation.

The effect of a small number of outlier composite grades or spatially isolated composites may have an undue effect on the estimated block grades within individual domains. The identification of outliers was undertaken using statistical tables, statistical summary charts and an investigation of the composite data in 3D visualization.

A number of high cuts or limits were identified as necessary within the mineral resource estimation domains. A statistical summary of these domains and their corresponding high cuts is shown in Table 14-123 below.

Domain Minimum Grade g/t
Au
Maximum Grade
g/t Au
Mean Gold
Grade g/t
High Grade Cut
g/t Au
Mean Cut Gold Grade
g/t
100 0.005 104 3.1 15 1.24
200 0.005 97.36 2.92 25 1.11
300 0.005 62.66 2.04 20 0.96
400 0.005 58.8 1.77 7 0.66
500 0.005 2.57 0.35 2 0.35
600 0.01 3.39 0.66 3 0.65
700 0.02 35 1.37 2.5 0.79

TABLE 14-121 STATISTICAL SUMMARY FOR GOLD IN G/T BY MINERAL RESOURCE ESTIMATION DOMAIN

Block Model Creation and Extents

The primary consideration of the 3D model was to provide an adequate level of resolution to cope with all volume related complexity. The 3D wireframes were used to create block model volume constraints for each estimation domain. All individual estimation domains were ultimately combined to create a single block model in the local grid coordinate system. Table 14-124 summarizes the 3D block model “Northpoint-120405.mdl” definition.

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  Northing Easting RL
 Minimum 9000    9800  1000
Maximum 10070 10205 1150
Block Size 5 4 2.5
Sub-block 2.5 1 1.25

TABLE 14- 122 3D BLOCK MODEL DEFINITION (M)

The chosen block size represents approximately half the best data spacing in the Northing and Easting directions and a choice in the vertical dimension controlled by the need to appropriately represent the volume of the wireframes that define the estimation domains.

A standard list of field names and descriptions used in the block model are shown in Table 14-125.

Attribute Type Default Description
 au Float  -99   Au Grade
density Float -99 SG
krig_var Float -99 Kriging Variance
lodecode Integer -99 Lode code
no_samp Integer -99 No of samples to inform a block
res_class Integer -99 Classification code

TABLE 14- 123 3D BLOCK MODEL ATTRIBUTES

The resource estimate contains implicit assumptions of mining selectivity represented by the block size of 5m x 4m x 2.5m (Y x X x Z).

The block model used a primary block size of 4m EW x 5m NS x 2.5m vertical with sub-cells of 2.5m x 1m x 1.25m. The parent block size was selected on the basis of around 50% of the average drill hole spacing.

Estimation Parameters and Methodology

Variography was used to characterize the spatial behavior of the composite data primarily as an aid to establishing estimation parameters. Variogram stability and quality is dependent upon the statistical properties and the amount of data available within the defined domains. After an initial investigation of the gold data, one variogram model was defined for all mineralized zones. The general variogram model is detailed in Table 14.126 below.

Domain Nugget Stuct Sill Major
(m)
Semi
(m)
Minor
(m)
Major/
Semi
Major/
Minor
Surpac Rotation
Bearing  Plunge  Dip 
All Mineralized
Lodes
0.72  St1 0.08 30 60 75 2.0 3.75 355 0 -65
St2 0.2 30 60 75 2.0 3.75 355 0 -65

TABLE 14-124 VARIOGRAM MODELS FOR GOLD BY MINERALIZED DOMAIN

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The features of the variogram model for gold can be summarized as moderately high relative nugget of about 70% for the mineralized domains with a significant amount of variability demonstrated over a short range. This reflects the high variability at short range observed in visual inspection by section. Maximum range extends to 75m. This is shorter than usual mineral resource estimations as this is using grade control drilling, which is generally closer spaced than exploration or resource definition drilling

Grade Interpolation

A standard 3 dimensional, 2 pass Ordinary Kriging methodology has been used for the estimation of the cut gold 1 meter downhole composite data within each estimation domain. Table 14-127 summarizes the estimation parameters by domain. A constant minimum of 4 and maximum of 10 composites have been set for all domains on the first pass whilst a minimum of 2 and a maximum of 10 composites were used for all domains on the second pass.

Domain Pass Search Radius
(m)
Bearing Plunge Dip Major/Semi Major/Minor
All Mineralized Domains 1st 20 355 0 -65 2 3.75
2nd 40 355 0 -65 2 3.75

TABLE 14-125 ESTIMATION PARAMETERS FOR GOLD BY ESTIMATION DOMAIN

Model Validation

The North Point deposit block model was validated by a couple of methods, including the following:

Mining Reconciliation

The North Point model has been constructed above the current mining area and can be used to reconcile against mill and mining production. This is the ultimate test for these models. During 2010 a total of 477,500 tonnes of material was mined at an average grade of 1.30g/t Au. The current mineral resource estimate predicted a total tonnage of 371,000 tonnes at an average grade of 1.56g/t Au. The difference between the models is significant but in the early days of mining some 0.5 -0.7g/t Au material was mined and processed. There would also be some dilution in the models, which is not determined at this stage. Overall ounce estimation is close with the model predicting 18,600 ounces and the mining predicting just under 20,000 ounces.

14.5.2.7  Princess Louise

Domaining

No top-cut was applied to zone 300, whereas a top-cut of 5.5g/t Au was applied to zone 200. Zones 100 and 500 had 12g/t Au applied as a top-cut and zone 400 a value of 12.5g/t Au. The global statistics of the top-cut values are depicted below in Table 14-128.

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Zones 100, 300, 400 and 500 have an average grade of approximately 1.0g/t Au. The variance and skewness values for zones 100, 400, 500 are similar with values of around 3 and 4 respectively.

Block Model Creation and Extents

A Surpac block model was created with Table 14-129 outlining the model extents:

Type Y X Z
Minimum Coordinates 6450 9375 1000
Maximum Coordinates 7475 10203 1200
User Block Size 5 4 2.5
Min. Block Size 2.5 2 1.25
Rotation 0 0 0

TABLE 14-127 PRINCESS LOUISE DEPOSIT – BLOCK MODEL EXTENTS

Attributes were created to account for the various parameters. The details of these attributes are given in Table 14-130 below:

Attribute
Name
Type Decimals Description
au_ok Float 2 OK estimation
av_dist Real 3 Average distance of samples used in interpolation
dist Real 3 Anisotropic distance to the first sample used
kv Real 3 Kriging Variance
material Character - Material: Waste, Air, 100, 200, 300, 400, 500
num Integer - Number of samples used in the interpolation
pass Integer - Pass number
sg Float 2 Specific gravity
weather Character - Weathering zone: Air, Fresh, Weathered

TABLE 14-128 PRINCESS LOUISE DEPOSIT – BLOCK MODEL ATTRIBUTES

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Estimation Parameters and Methodology

Ordinary Kriging was used to interpolate grades into the model. Only the cut composite data inside each zone was used to interpolate grades into the zone. Table 14-131 below depicts the parameters used during the interpolation:

Parameter Pass 1 Pass 2 Pass 3
Min samples 5 5 5
Max samples 15 10 10
Max samples per drillhole 5 5 5
       
Search distance 25 50 200
       
Nugget   2.1752  
Sill first structure   0.577  
Range first structure   8  
Sill second structure   0.382  
Range second structure   25  
       
Major: Semi major   1.5  
Major: Minor   2  

TABLE 14- 129 PRINCESS LOUISE DEPOSIT – INTERPOLATION PARAMETERS

Grade Interpolation

From the statistical analysis above, the global statistics (average, variance and standard deviations) for zones 100, 400 and 500 are very similar with lots of data in each zone. These zones were combined into one overall domain for the purpose of variography analysis. Zones 200 and 300 have few samples to obtain meaningful variograms so these zones were also included into the domain above.

An omni-directional variogram was extracted to obtain the nugget value and a value of 2.1752 was obtained with an overall sill value of 3.1333 giving a high nugget of approximately 70% of the total sill value. This shows high variability in the grade values in the zones.

The direction that was chosen for the variogram analysis was along a strike of 20 degrees (major) and a dip of -65° towards 290°. A spread of 10° was used with a spread limit of 15m to approximate the width of the mineralized zones.

Model Validation

Figure 14-86 to Figure 14-90. Figure 14-86 below depicts the validation of the model against input composite grades. The model was divided into 50m vertical slices from south to north and compared directly against average composite grades within the same slices.

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The two lines represent the grade for the composite file (blue) and the interpolated grade (red). These two need to follow the same trend.

Mining Reconciliation

Crocodile Gold carried out two campaigns of mining at Princess Louise. The first was a trial period between September 2010 and November 2010. Mining was stopped during the wet season of 2010-11 due to the local roads being closed to heavy haulage. The second period started in June 2011 and finished in December of the same year. In that period a total of 434,000 tonnes of material was mined and trucked to Union Reefs for processing. The average grade of the material was 1.17 g/t Au for 16,300 ounces. During the same period the reconciled mineral resource estimate predicted around 424,000 tonnes of material at an average grade of 1.38 g/t Au for 18,800 ounces, therefore the tonnes predicted well but the grade call was around 14% higher than achieved. This would need to be reviewed before any future mining was planned.

14.5.2.8  Fountain Head

Domaining

Statistics were run within the drillhole database for all constrained uncut composite data by lode, and are presented in Table 14-132. No other mineralization indicators were used, as data was extracted from within wireframes.

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Block Model Creation and Extents

A 3D block model, FountainHead_sept2006.mdl was generated using Surpac software with origin, extents and attributes defined below in Table14-133. Blocks were subdivided into sub-blocks of 2.5m x 1m x 1.25m (along-strike x across-strike x RL) in order to fill areas adjacent to wireframe boundaries. The solid wireframes were used to limit the blocks available for grade interpolation, with block centroid locations used to define the blocks and sub-blocks for interpolation.

Estimation Parameters and Methodology

Ordinary Kriging using parameters derived from the lognormal variograms was used to estimate gold grades for the Fountain Head Project.

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Kriging Neighbourhood Analysis (KNA)

A Kriging neighbourhood analysis was undertaken for all Fountain Head lodes with the aim of minimizing conditional bias issues and determining the optimal block size, maximum search range and number of composites for use in the block model. The KNA results, the lode domain geometry and previous model parameters were all used as a guide in the determination of the final parameters. Estimation parameters selected as a result of this KNA analysis include the following:

An estimation block size of 5mE x 2mN x 2.5mRL;
   
A maximum of 25 composites used to estimate a block for all lodes;
   
A minimum of 4 composites used to estimate a block for all lodes; and
   
A maximum search range of 40m for all lodes.

Grade Interpolation

Each lode was treated as a separate hard boundary, restricting the Au grade interpolation to drillhole data located within each lode. As an example, a block in the footwall could use hinge composites for interpolation as well as footwall composites, provided they were within the applicable search ellipse, and composites were located within the same lode. A discretization array of 5 (east) by 2 (north) by 2 (RL) was used to refine the Kriging weights for each model block.

Two interpolation passes were conducted for all lodes, with an initial search pass of 40m x 20m x 8m, and a second search of either 60m x 30m x 12m or 80m x 40m x 16m. Only those blocks unfilled by the previous pass were interpolated by the second pass, and grades estimated from the previous interpolation pass were left unchanged.

Model Validation

The Fountain Head deposit block model was validated by several methods, including the following:

  on-screen visual validation via vertical sections and plan views, showing block grades;
  against input composites;
  global statistical comparisons of average input composites and tonnage-weighted;
  block grades; and
  local grade/depth and grade/easting relationship plots.

A global grade validation was conducted by comparing input average top-cut composite grades with mean block grades by lode.

A good overall global reconciliation between average input composite grades and mean block grades is present, apart from the central and fault domains which show a slight underestimation of global grades. These domains contain large uninformed areas at depth and to the west, with drillhole composites bordering this area comprising low grades. These low grades have had a disproportionate influence on surrounding block grades, hence lowering the average block grade relative to the composite grade.

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Model and composite grade/depth relationships were compiled for all lodes within the Fountain Head deposit, and are presented in Figure 14-94 & Figure 14-95. Both input average composite data and tonnage-weighted mean model grade data were averaged within 10m RL slices for each lode, and plotted together with the number of composites to assess the reliability of the block model.

Comparisons of model grades with composite grades for all lodes illustrate a relatively good reconciliation, with model grades reproducing the fluctuations in composite grades with respect to depth. Deviations of block grades from composite grades generally occur in areas of low composite numbers, where these composites influence a disproportionate number of model blocks, as is the case at depth and in the western area of the model.

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Grade/Easting Validations

Figure 14-96 and Figure 14-97 illustrate the model and composite grade/Easting relationships, based on 20m vertical slices from east to west.

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14.5.2.9  Tally Ho

Domaining

Six lode domains were delineated for the Tally Ho deposit by GBS Gold, using the approximate geological boundaries of broad quartz veins, which equates to a 0.4 g/t Au lower cut-off. The 200 Lode, being the largest lode is subdivided into 4 domains (201 to 204) on the basis of geometry for statistical and geostatistical analysis. A nominal surface was constructed separating the shallow-dipping and steep-dipping quartz veining of this lode into two sub-domains. These two sub-domains were further divided by another sub-vertical surface at approximately 9775mE to represent different lode trends.

Lode 300 was sub-divided at 9675mE and 9810mE into three domains (301 to 303) based on different trends within the lode. It represents the second largest lode in the deposit. Lode 100 comprises two discrete wireframes and thus formed separate domains. Lode 500 was separated at 9675mE into two domains, again on the basis of differing trends within this lode. The remaining 2 lodes (400 and 600) are small lodes, which run parallel to one of the four main lodes. Neither of these lodes has been subdomained.

Statistics were run within the drillhole database for all constrained uncut composite data for the Tally Ho deposit, and are presented in Table 14-134. No other mineralization indicators were used, as data was extracted from within wireframes.

Block Model Creation and Extents

A 3D block model, tallyho_april2008.mdl was generated using Surpac software with origin, extents and attributes defined below in Table 14 96Table 17.5. Blocks were subdivided into sub-blocks of 2.5m x 1.25m x 1.25m (along-strike x across-strike x RL) in order to fill areas adjacent to wireframe boundaries.

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The solid wireframes were used to limit the blocks available for grade interpolation, with block centroid locations used to define the blocks and sub-blocks for interpolation.

Model Parameters Y X Z
Minimum Coordinates 9700 9500 870
Maximum Coordinates 10000 9950 1120
Model Extent 300m 450m 250m
Parent Block Size 5m 10m 5m
Minimum Block Size 1.25m 2.5m 1.25m
Attribute Type Description
au au_oz Float Au estimate grade
resclass Calculated Au ounces per block
density Integer Classification code
krig_var Float SG
domcode Real Kriging variance
lodecode Integer Domain code
pass Integer Lode code
no_samp Integer Interpolation pass no
  Integer No of samples

TABLE 14-133 TALLY HO DEPOSIT BLOCK MODEL PARAMETE RS

Estimation Parameters and Methodology

Ordinary Kriging using parameters derived from the lognormal variograms was used to estimate Au grades for the Tally Ho deposit.

Kriging neighbourhood analysis was not completed for this resource estimate, as Kriging neighbourhood parameters from the previous resource estimate were deemed to be appropriate for this current model. These estimation parameters include the following:

Grade Interpolation

Each lode was treated as a separate hard boundary, restricting the Au grade interpolation to drillhole data located within each envelope. Some lodes were sub-divided into domains on the basis of changes in strike and/or dip with respect to easting (Table 14-136). Soft boundaries were created between domains within each lode, with blocks in each domain able to utilize composites in the other domains of the same lode if located within the search ellipse. A discretization array of 5 (east) by 5 (north) by 5 (RL) was used to refine the Kriging weights for each model block.

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Three interpolation passes were conducted for all lodes with an initial search ellipse of 30m x 15m x 10m used. A second pass with search dimensions increased to 45m x 23m x 12.5m and a third pass with search dimensions increased to 60m x 30m x 15m were conducted, with only those blocks unfilled by the previous pass interpolated. Grades estimated from the previous interpolation pass were left unchanged.

Four domains (110, 240, 330 and 520) had some unfilled blocks after all interpolation passes. These domains included extrapolated areas at depth with no drill coverage, and blocks in these areas were not filled despite using a doubled search ellipse on the third interpolation pass. Hence, it was deemed appropriate to leave these blocks unfilled.

Model Validation

The Tally Ho deposit block model was validated by several methods, including the following:

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LODE NO OF
COMPOSITES
COMPOSITE
MEAN AU
MODEL MEAN
AU
% DIFFERENCE
100 144 1.68 1.56  -7.4%
 200  954 1.79 2.23 6.9%
 300 363 2.05 2.19 24.4%
400   22 0.52 0.55  5.6%
500 113 1.29 1.24 -4.5%
600 21 1.14 1.17 1.8%

TABLE 14-135 TALLY HO DEPOSIT GLOBAL STATISTICAL VALIDATION OF AU INTE RPOLATED GRADES G/T

Global reconciliation between average input composite grades and mean block grades is around 5% for most lodes, with the exception of lode 200. High-grade composites from THRC106 have spread high-grades down-plunge without any surrounding constraints, and thus these composites have had a disproportionate influence on surrounding blocks.

Model and composite grade by depth and by easting relationships were compiled for all lodes within the Tally Ho deposit, and are presented in Figure 14-98 to Figure 14-100. Both input average composite data and tonnage-weighted mean model grade data were averaged within 10m RL slices for depth comparison and 20m sections for easting comparison of all lodes, and plotted together with the number of composites to assess the reliability of the block model.

Comparisons of model grades with composite grades illustrate a robust reconciliation apart from the 500 lode, with model grades reproducing the fluctuations in composite grades. Deviations of block grades from composite grades generally occur in areas of low composite numbers (as in lode 500), where these composites influence a large number of model blocks, as expected. The graphs for model grades extend beyond the limits of composite grades for several lodes, particularly at depth, as a result of the wireframes being extrapolated beyond composite data limits.

The Easting and depth validation plots for lodes 400 and 600 are not displayed, due to the limited number of composites and the small size of these lodes.

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15 MINERAL RESERVES

A summary of all mineral reserves for the NT Operations can be seen in Table 15-1 below. Descriptions on how the mineral reserves were calculated can be seen in Table 15-2 to Table 15-5.

 Mineral Reserves as of Dec 31, 2015 
Deposit Category Tonnes (T) Gold Grade
(g/t)
Oz Gold
(Oz)
Cosmo UG Proven 487,000 3.47 54,400
Probable 445,000 3.28 46,900
Sub-Total 932,000 3.38 101,300
Union Reefs OP
(Esmealda)
Probable 244,000 1.61 12,700
Union Reefs UG
(Prospect)
Probable 276,000 4.42 39,200
Pine Creek OP Probable 1,245,000 1.55 62,100
Sub-Total Proved 487,000 3.47 54,400
Sub-Total Probable 2,210,000 2.26 160,900
Total Reserves 2,697,000 2.48 215,300

TABLE 15-1 NT OPERATIONS MINERAL RESERVE SUMMARY – EFFECTIVE DEC 31, 2015

15.1 COSMO MINE

The mineral reserve estimate for the Cosmo Mine is based on a combination of modified Avoca stoping and longhole open stoping techniques. These mining methods are described in detail in Section 16.1.

Classification Tonnes (t) Gold (g/t) Gold (oz)
Proven      
Underground 479,000 3.50 53,800
Stockpile 8,000 2.38 600
Proven Subtotal 487,000 3.47 54,400
Probable      
Underground 445,000 3.28 46,900
Total mineral reserve 932,000 3.38 101,300

TABLE 15-2 MINERAL RESERVE CLASSIFICATION FOR COSMO AS AT DECEMBER 31, 2015

Notes to accompany Table 15-2:

1.

The mineral reserve is stated as of December 31, 2015.

   
2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the National Instrument 43-101.

   
3.

Mineral reserves were estimated using the following mining and economic factors:

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  a.

14% dilution at 0.5g/t Au is added to all stopes, based on reconciled 2015 production.

  b.

Minimum stope width of 3.0m.

  c.

Stope recovery of 90%, based on reconciled 2015 production.

  d.

Crown pillar recovery of 50%.

  e.

15% dilution at the mineral resource grade is added to all development.

  f.

Mineralization development recovery of 100% is assumed.

  g.

A gold price of $A1,450/oz.

  h.

An overall processing recovery of 92.0% at a cost of $28.90/t.

  i.

Total mining cost used of $68.72/t.

  j.

Stockpiles include Cosmo material at the Mine and Union Reefs processing facility.

  k.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100 oz.


4.

The cut-off grade for mineral reserves has been estimated at 2.3g/t Au.

   
5.

Mineral reserve estimates were reviewed by Murray Smith who is a consultant with Mining Plus Pty Ltd. Mr Smith is a Member and Chartered Engineer of the Australasian Institute of Mining and Metallurgy, has over 20 years of relevant engineering experience and is the Qualified Person for mineral reserves for Cosmo Mine as per the National Instrument 43-101.


15.2

UNION REEFS UNDERGROUND - PROSPECT

The mineral reserve estimate for the Prospect underground mine is based on uphole benching with backfill. This mining method is described in detail in Section 16.2. Table 15-3 shows the mineral reserve classification figures which are inclusive of the modifying factors for mining recovery and dilution.

Classification Tonnes (kt) Gold Grade
(g/t)
Gold (ozs)
Proven      
Probable 276,000 4.42 39,200
Total mineral reserve 276,000 4.42 39,200

TABLE 15-3 MINERAL RESERVE CLASSIFICATION PROSPECT DEPOSIT UNDERGROUND AS AT DECEMBER 31, 2015

Notes for Table 15-3:

1.

The mineral reserve is stated as of December 31, 2015.

   
2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the NI 43-101.

   
3.

Mineral reserves were estimated using the following mining and economic factors:


  a.

A 0.2m hangingwall and footwall skin has been added to the economic stope shape to allow for dilution .

  b.

Minimum stope width is 2m.

  c.

Stope recovery is 95%.

  d.

A gold price of $A1,450/ oz.

  e.

An overall processing recovery of 93% at a cost of $28.90/t.

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  f.

Total mining cost of $87.10/t.

  g.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100 oz .


4.

The cut-off grade for mineral reserves has been estimated at 2.7g/t Au.

   
5.

Mineral reserve estimates were prepared by Murray Smith, who is a consultant with Mining Plus Pty Ltd. He is a Member of the Australasian Institute of Mining and Metallurgy and has over 20 years of relevant engineering experience and is the Qualified Person for mineral reserves for Prospect underground as per the National Instrument 43-101.

The mining sequence includes permanent rib pillars that separate individual stopes. If a consolidated fill was introduced into the mining sequence a high proportion of these pillars could be extracted thereby potentially increasing the mineral reserve.

Sensitivity analysis conducted as part of the economic assessment shows that a 10% decrease in grade, recovery or gold price will still result in a positive NPV being maintained by the project.

A 5% decrease in metallurgical recovery or a 10% increase in costs would increase the cut-off grade from 2.7g/t to approximately 2.9g/t Au. As the mineral reserve relies on a high-grade core, an increase in cutoff grade of this range will have very little effect on the reserve.

15.3 UNION REEFS OPEN PIT – ESMERALDA

The mineral reserves for the Esmeralda deposit are based on the open pit mining techniques. The mining method is described in detail in Section 16.3.

Classification Tonnes (t) Gold Grade (g/t) Gold (ozs)
Proven      
Probable 244,000 1.61 12,700
Total mineral reserve 244,000 1.61 12,700

TABLE 15-4 MINERAL RESERVE CLASSIFICATION ESMERALDA OPEN PIT AS AT DECEMBER 31, 2015

Notes for Table 15-4:

1.

The mineral reserve is stated as of December 31, 2015

   
2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the National Instrument 43-101

   
3.

Mineral reserves were estimated using the following mining and economic factors:


  a.

Dilution of 10% and mineralization loss of 5%

  b.

Mining costs of $4.50/t and processing costs of $26.00

  c.

A gold price of $A1,450/oz

  d.

An overall processing recovery of 90%

  e.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100 oz

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4.

The cut-off grade for mineral reserves has been estimated at 0.7g/t Au

   
5.

Mineral reserve estimates were prepared by Mark Edwards who is a Member of the Australasian Institute of Mining and Metallurgy and has over 18 years of relevant experience and is the Qualified Person for mineral reserves at Esmeralda open pit as per NI 43-101.


15.4 PINE CREEK OPEN PITS

Mineral reserves for Pine Creek are set forth below. The mining method is describe in more detail in Section 16.4.

Deposit Classification Tonnes
(t)
Gold Grade
(g/t)
Gold
(ozs)
Cox Proven      
Probable 133,000 1.61 6,900
International Proven      
Probable 860,000 1.30 35,900
Kohinoor Proven      
Probable 129,000 2.39 9,900
South Enterprise Proven      
Probable 123,000 2.37 9,400
Total mineral reserve 1,245,000 1.55 62,100

TABLE 15-5 MINERAL RESERVE CLASSIFICATION FOR PINE CREEK AS AT DECEMBER 31, 2015

Notes for Table 15-5:

1.

The mineral reserve is stated as of December 31, 2015

   
2.

All mineral reserves have been estimated in accordance with the JORC code and have been reconciled to CIM standards as prescribed by the NI 43-101

   
3.

Mineral reserves were estimated using the following mining and economic factors:


  a.

Dilution of 15% and mineralization loss of 5% for all pits excluding International which used a mining dilution of 10%

  b.

Mining costs of $4.80/t and processing costs of $33.24

  c.

A gold price of $A1,450/oz

  d.

An overall processing recovery of 90% for all pits excluding International, which used a recovery of 85%

  e.

Tonnes are rounded to the closest 1,000t and ounces are rounded to closest 100oz


4.

The cut-off grade for mineral reserves has been estimated at 0.9g/t Au.

   
5.

Mineral reserve estimates were prepared by Mark Edwards who is a Member of the Australasian Institute of Mining and Metallurgy and has over 18 years of relevant experience and is the Qualified Person for mineral reserves at Pine Creek as per the NI 43-101.

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15.5 CONCLUSION ON MINERAL RESERVES

There are no known situations where the mineral reserves outlined above could be materially affected by environmental, permitting, legal, title, treatment, socio-economic or political issues. There is however some risk with any gold mineral reserve where the gold price realized may affect the overall economic viability of a mining operation.

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16 MINING METHODS

While each area outlined below has a separate mining method described all sites still utilize the Union Reefs processing facility as a single infrastructure requirement. Therefore this is described as a single project for reporting.

16.1 COSMO MINE

16.1.1      INTRODUCTION

Newmarket Gold is currently mining the Cosmo deposit as an underground mine at a current depth of 600mRL (presented in Figure 16-4). The Cosmo workings are accessed via a decline commencing from the southern wall of the Cosmo open pit. The current mining activity is centered on the Eastern Lode, utilizing Avoca-style cut and fill stoping with regional sill and crown pillars.

16.1.2      GEOTECHNICAL

16.1.2.1  Geotechnical Zones

Mineralization at Cosmo Mine generally occurs within a package of metamorphosed sediments between the Zamu Dolerite Sill and an outer thick carbonaceous mudstone unit (Pmc). The main mineralization zone on the Eastern Limb has been divided into four lodes (100, 200, 300 & 400) with three waste zones (not shown) separating the mineralization. Whilst the majority of the mineralization has come from the 100 Lode, stoping has occurred on all four lodes throughout the mine.

Figure 16-1 outlines the mineralization lodes on the Eastern Limb beneath the open pit sectioned on Northing 1520mN.

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16.1.2.2  Stope Design Criteria

Stope spans are designed using Mathew’s Stope Stability Method (Mathews, et al. 1980), with a stope stability chart, which has been calibrated from site experience (Figure 16-2).

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Input data for the stability number for individual stopes is based on drillhole information situated in the hangingwall of the stope. Stope spans are typically in the range of 12–25m in length. The main impact on stope stability is the location of the F10 Fault, which sits in the hangingwall of the 100 Lode. Dependent on the distance of the F10 Fault from the hangingwall the stope will be designed to either include the F10 inside the stope design or leave a skin of mineralization to hold the hangingwall in place. Cable bolts of 6–9m in length are also installed into the hangingwall from the level accesses to minimize over-break and damage to the mineralization drives.

Once a stope reaches its design length, the stope is backfilled with run of mine waste before the next stoping panel is opened up.

16.1.2.3  Ground Support Requirements for Decline and Level Development

In 2014 a review of the ground support requirements for Cosmo Mine was conducted by AMC (A.M.C 2014). This coincided with a change of mining contractor and a move to using mesh instead of fibrecrete as the main form of surface support. The summary of rock mass properties is presented in Table 16-1. Recommended ground support patterns are presented in Table 16-2, which are based on Barton’s Q rock mass classification system (Figure 16-3). The standard ground support pattern in use at the mine is 5.6mm diameter, 100x100mm aperture weld mesh brought down to 3.5m from the floor with a pattern of grouted split sets on a 1.4x1.5m spacing. All intersections are also pattern cable bolted with 6m long twin-strand cable bolts.

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Lithology Q
minimum
Q
maximum
Qualitative Description
Code Name
Pgt Greywacke 0.7 4.5 Very Poor to Fair
Psl Siltstone 0.7 6.3 Very Poor to Fair
Pdz Dolerite 0.8 3.2 Very Poor to Poor
Pmc Carbonaceous mudstone 0.3 1.0 Very Poor
Pca Carbonate 9.0 38.0 Good
Pca/Psl Carbonate/Siltstone contact 0.03 0.13 Extremely Poor
Pdz/Pgt Dolerite/Greywacke contact 0.1 0.3 Very Poor
F10 Fault 0.05 0.2 Very Poor to Very Poor

TABLE 16-1 SUMMARY OF ROCK MASS QUANITIES OF COSMO GEOTECHNICAL DOMAINS (A.M.C 2014)

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Lithology Alternative
Systems
Rock Bolts Surface Support
Pgt, Psl, Pdz Alternative 1 2.4m galvanized split sets installed on 2.3m spacing FRS 75 mm nominal thickness
Alternative 2 2.4m galvanized split sets installed on 1.4m spacing Weld mesh 5.6mm wire, 100mm aperture, galvanized, 3.5m from floor
Pmc Alternative 1 2.4m galvanized split sets installed on 1.8m spacing FRS 100mm nominal thickness
Alternative 2 2.4m galvanized split sets installed on 1.0m spacing Weld mesh 5.6mm wire, 100mm aperture, galvanized, 1.5m from floor
Pca/Psl contact
Pdz/Pgt contact
F10 Fault
   2.4m galvanized split sets installed on 1.4m spacing FRS 120mm nominal thickness
Pca   2.4m galvanized split sets installed on 2.3m spacing Weld mesh 5.6mm wire, 100mm aperture, galvanized, 3.5m from floor

TABLE 16-2 RECOMMENDED PRIMARY GROUND SUPPORT SYSTEM AT COSMO MINE (A.M.C 2014)

16.1.2.4 Decline Location

The decline is located in the hangingwall of the Eastern Lodes (Figure 16-4). This allows the majority of the decline to be hosted in the more competent Zamu Dolerite and also provides the best access to exploit both limbs of the Cosmo fold structure, being located centrally between the east and west limbs of the Cosmo Anticline.

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16.1.3      MINE DESIGN

16.1.3.1  Mining Method Description

Ore is mined by a combination of Modified Avoca and Long-hole Open Stoping when isolated stopes are extracted, which often do not require backfilling. The Modified Avoca mining method is most commonly used. Figure 16-5 shows existing development and stopes mined, plus those included in the mineral reserves for future extraction.

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16.1.3.2  Modified Avoca

Avoca mining is sequenced from the bottom up, which delays the production of stope mineralization until the lowest development drive has been completed. To reduce the delay of developing to the bottom of the mine before stope tonnes are available, sill pillars are designed at intervals within the mining sequence to allow earlier production. The sill pillars provide support for stoping below already open stopes. The location of the sill pillars are flexible to allow for variations in mineralization grade from the block model and for changes in scheduling due to the rate of decline development or mineralization extraction. Figure 16-6 to Figure 16-7 outline the development, stoping and fill sequence for the Cosmo Mine. This method allows for the stockpiling of potentially acid producing waste rock underground. The number of sill pillars remaining in mineralization may be reduced in areas of good grade by the placement of cemented rock fill (CRF) in the lowest level of the panel sequence. Stoping of the top level will be by longhole open stope using upholes from the top drill drive once extraction and backfilling of the level below is complete. Figure 16-8 outlines the sequence including CRF.

The production cycle for Modified Avoca stoping includes the following activities:

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16.1.3.3  Cut-off Grade

Table 16-3 shows the cut-off grade assessment that has been undertaken for the Cosmo Mine. The inputs are based on the 2016 budget mining costs at the Cosmo Mine and milling costs at the Union Reef process plant (Table 16-4). This cut-off grade is applied to all stopes and covers all operating costs. Each stope or group of stopes on each mining level is tested to ensure that the level generates a positive cash flow after accounting for all operating and capital development required to access and to recover these stopes.

Any marginal development mineralization, which is mined in the process of accessing these economic stopes is also included in the mineral reserves. This is only applied if the development material had to be trucked to surface anyway and does not displace higher-grade mineralization from the mill.

Cut-off Grade
Calculation
Units $/t (mineralization)
Metal Price ($/oz) A$ 1,450
Mining Dilution % 14%
Mining Recovery % 90%
Processing Recovery % 92%
A$/g Recovered A$/g 42.89
Total Mining Costs A$ $68.72
Total Processing Costs A$ $28.90
Total Costs A$ $97.62
Cut-off grade g/t Au g/t 2.30

TABLE 16-3 COSMO MINE CUT-OFF GRADE CALCULATIONS

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Operating Costs (Mining) $/t
(mineralization)
Mining Fixed Costs $12.29
Power & Fuel $7.52
Stope Production Costs $19.99
Technical Services $1.75
Geology $6.32
Mineralization Haulage $6.41
Camp Costs- Cosmo $4.59
Processing $28.90
Administration $8.12
Cosmo Maintenance $1.75
Total $97.62

TABLE 16-4 MINE OPERATING COSTS

16.1.3.4  Material Handling

All underground production mineralization is trucked via the decline to the surface ROM stockpile adjacent to the Cosmo open pit. Mineralization is then loaded into on-highway haul trucks to be transported 67km to the Union Reefs processing plant.

Development waste is used for stope backfill as dictated by the production schedule. Waste not immediately required for backfill is typically stockpiled underground close to stope backfill locations. Occasionally when there is insufficient capacity to stockpile waste underground it is brought to surface and stockpiled in the Cosmo pit for future use as stope backfill.

16.1.4      MINE DESIGN GUIDELINES

16.1.4.1  Design Parameters

The mine design parameters used in the design of the Cosmo Mine are summarized in Table 16-5.

Item Size Gradient
Decline 5.5m H x 5.5m W 1:7 down
Decline stockpile 5.5m H x 5.5m W 1:50 up
Level Access 5.5m H x 5.5m W 1:50 up
Level Stockpile 7.5m H x 5.5m W 1:50 up
Vent Access, Escapeway access 5.0m H x 5.0m W 1:50 up
Ore Drives 5.0m H x 5.0m W 1:50 up
Sumps 5.0m H x 5.0m W 1:7 down
Ventilation Rises (Long hole) 5.0m H x 5.0m W vertical
Ventilation Rises (Raise bore) 4.5m diameter vertical
Escapeway Rise (Raise bore or Airleg rise) 4m diameter 60 < Ø < 75 degrees

TABLE 16-5 DESIGN PARAMETERS

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16.1.4.2  Mining Sequence

The mining sequence is bottom up in panels of three or four levels. The sequence for each mining block commences once the decline has reached the level of the lower access. The lower levels have priority as stoping of the top level can only commence upon completion of the lower two levels. Stoping on each level retreats from the northern and southern extents back to a central access. Stopes on the upper levels of a panel with no stopes beneath can commence prior to the lower levels.

16.1.4.3  Decline Development

Access to the mineral resource would be via a continuation of the current decline commencing from the 600mRL down. Decline standoff is kept at a minimum of 30m from planned stope voids.

16.1.4.4  Level Development

Truck-loading stock-piles (20m in length) are mined mid-way along these accesses, necessitating decline dimensions to be maintained for the first 25m to 30m of level access development. Beyond this point level access dimensions are reduced to mineralization drive dimensions. A sump is mined prior to the access intersecting the mineralization. The Level access development gradient of 1:50 allows mine production water to report to sumps on the level access. Life of Mine lateral development is shown in Table 16-6.

Cosmo Mine Development (m)
Decline 0
Access 335
Drill Drive 181
Stockpile 80
Ore drive 1,165
Total 1,761

TABLE 16-6 LATERAL DEVELOPMENT

16.1.4.5  Crown Pillar

The crown pillar includes all mineral reserves that currently sit above the 955mRL and below the base of the Cosmo pit at the 1000mRL. Development of an intermediate level at the 975mRL will be required prior to stoping of crown pillar. Stopes above the 975 drive cannot be mined until the tailings (2.25M m3) are removed from the Cosmo open pit. The associated capital cost for the tails removal has been considered and is included in the summary of capital costs in Table 22-1. A conservative recovery factor of 50% has been applied to the crown pillar stopes above 975mRL.

16.1.5      VENTILATION

16.1.5.1  Ventilation Circuit

The mine is ventilated by drawing fresh air down the decline and also down a main intake rise in the Cosmo Pit (Figure 16-9). A 250kW fan located on an exhaust rise in the Cosmo pit and twin-500kW exhaust fans located at the north end of the mine draw the air into the mine. The main air intake in the pit is also equipped with 3 x 500kW chiller plants, which are used in the wet season months (October to April) in order to reduce wet bulb air temperatures in the mine. The return air rise in the center of the mine returns air from the 855mRL to surface. The northern exhaust system returns air from the lowest active part of the mine (605mRL) to surface.

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16.1.5.2  Airflow Requirements

Airflow requirements for the mine are based on an allowance of 0.05m 3/s per kW of diesel-powered equipment. Currently, 460m3/s of air is exhausted through the twin 500kW exhausting system at the northern end of the mine and one 250kW fan in the center of the mine. Total mine requirements based on dilution of diesel exhaust was 250m3/s in January 2016 (Table 16-7).

Equipment Maximum
Number
kW
Rating
Airflow Required
(m3/s) per Unit
Total Airflow
Required (m3/s)
Twin Boom Development Jumbo- Sandvik DD421C 1 110 5.5 5.5
Production Long Hole Drill Rig- Sandvik DL431C 1 110 5.5 5.5
Cable bolting Drill Rig- Sandvik DL421C 1 110 5.5 5.5
Loader- Sandvik LH621 2 352 17.6 35.2
Loader- Sandvik LH517 1 285 14.25 14.3
Underground Dump Truck- Sandvik TH663 3 567 28.35 85.1
Charge Up unit- Normet Charmec 1 110 5.5 5.5
Integrated Tool Carrier- Volvo L120/CAT IT28G 2 180/107 9/5.4 14.4

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Equipment Maximum
Number
kW
Rating
Airflow Required
(m3/s) per Unit
Total Airflow
Required (m3/s)
Grader - Caterpillar 12M 1 180 9 9
Spraymec 6050W 1 75 3.75 3.75
Concrete Truck 1 110 5.5 5.5
LVs - V8 6 151 7.55 45.3
LVs -6cyl 4 75 3.75 15
Total Required       250

TABLE 16-7 AIRFLOW REQUIREMENTS

16.1.5.3  Raise Sizes

The velocity of air in a return air rise is recommended not to exceed 20m/s. Shorter raises are excavated by longhole blasting and longer raises using a raisebore. The minimum size of any raise is 4mx4m square or 4m in diameter. The ventilation system is capable of servicing the current ventilation requirements and currently anticipated future requirements.

16.1.5.4  Backfill

Stopes are backfilled with rock generated from waste development in the decline, stockpiles and access drives. Stopes requiring CRF backfill have cement slurry delivered to a waste rock stockpile for mixing prior to placement in stope voids. Any waste rock deficit is supplemented with surplus waste rock material from surface.

16.1.5.5  Mine Services & Infrastructure

The current mine services and infrastructure at Cosmo Mine are adequate for the continuation of operations with future extensions to the electrical and ventilation circuits being completed as required.

16.1.5.6  Dewatering

The current pumping system will be maintained and extended as development progresses to access future resource additions.

16.1.6      MINING SCHEDULE

16.1.6.1  Scheduling Strategy

The scheduling strategy for the mine is:

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The mining sequence is generally backfill constrained and requirements for rib and sill pillars are considered based on stope geometry and the specific sequencing of stopes.

16.1.6.2  Development Schedule

The lateral development quantities are presented in Table 16-8

  Cosmo Mine Development (m)
Description / Year 2016 2017 Total
Access 111 224 335
Stockpile 40 40 80
Mineralized Development Drive 537 628 1,165
Drill Drive 181   181
Total 869 892 1,761

TABLE 16-8 COSMO MINE DEVELOPMENT SCHEDULE

16.1.6.3  Production Schedule

The production schedule for the Cosmo Mine is summarized in Table 16-9.

  2016 2017 2018 Total
Mining Method tonnes Grade
(g/tAu)
tonnes Grade
(g/tAu)
tonnes Grade
(g/tAu)
tonnes Grade
(g/tAu)
Au Oz
Production 601,713 3.37 167,353 3.63 77,357 3.50 846,423 3.43 93,394
Development 43,152 2.66 34,758 3.32 0 0.00 77,911 2.95 7,398
Total 644,865 3.32 202,111 3.57 77,357 3.50 924,333 3.39 100,792

TABLE 16-9 COSMO MNE PRODUCTION SCHEDULE

16.1.6.4  Equipment

Table 16-10 presents the current mining fleet for the Cosmo Mine. No expansion of the fleet will be required beyond these levels to enable extraction of the ,mineral reserves.

Machine Number
Twin Boom Development Jumbo- Sandvik DD421C 1
Production Long Hole Drill Rig- Sandvik DL431C 1
Cable bolting Drill Rig- Sandvik DL421C 1

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Machine Number
Loader- Sandvik LH621 2
Loader- Sandvik LH517 1
Underground Dump Truck- Sandvik TH663 3
Charge Up unit- Normet Charmec 1
Integrated Tool Carrier- Volvo L120/CAT IT28G 2
Grader - Caterpillar 12M 1
Fibrecrete Machine- Spraymec 6050W 1
AGI Concrete Truck 1
Diamond Drill LM75/90 2
Light Vehicle- Toyota 4WD utility 10

TABLE 16- 10 MINING FLEET

16.1.6.5  Manpower and Supervision

Cosmo Mine is a continuous mining operation running 24 hours per day 365 days per year. Supervisors, operators and maintenance personnel work a two weeks on one week off roster with 12 hour shifts alternating between day shift and night shift. Newmarket Gold support staff work either nine days on five days off, or five days on 2 days off and on either 10 or 12 hours/day.

Labor costs are incorporated in the unit costs for mining and have been included in the financial evaluation. Newmarket Gold’s Cosmo staff and the underground shift mining personnel are presented in Table 16-11 and Table 16-12 respectively.

Newmarket Gold Staff Number
Mine Operations Manager 1
Technical Services Superintendent 1
Senior Production Engineer 1
Mine Planning Engineer 1
Geotechnical Engineer 1
Drill and Blast Engineer 2
Mining/ Ventilation Engineer 1
Mine Surveyor 2
Geology Superintendent 1
Senior UG Geologist 2
UG Mine Geologist 3
Graduate Geologist 1
Underground Geology Technician 3
Electrical Superintendent 1
Mine Electrician 1
Maintenance Supervisor 1
Light Vehicle Fitter 1
Dewatering Fitters 2
HSEC Manager 1
Medic/Safety Officer 2
Environment Officer 3
Site Administration 1
Total Staff 33

TABLE 16- 11 NEWMARKET GOLD PERSONNEL REQUIREMENTS

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 Underground Contract Personnel Contractor Number
Project Manager Downer 1
Mine Foreman Downer 2
UG Supervisors Downer 3
Safety and Training Manager Downer 1
Safety and Training Officer Downer 2
Maintenance Superintendent Downer 1
Maintenance Planner Downer 2
Electrical Supervisor Downer 1
Site Clerk Downer 2
Store person Downer 3
Jumbo Operator Downer 3
Cable bolt Operator Downer 2
Loader Operator Downer 9
Truck Operator Downer 9
Shotfirer Downer 3
Production Drillers Downer 3
Leading Hand Service Crew Downer 3
Service Crew Downer 3
Grader Operator Downer 2
Electrician Downer 3
Leading Hand Fitter Downer 3
Shift Fitter Downer 6
Light Vehicle Fitter Downer 2
Auto Electrician Downer 2
Specialist Drill Fitter Downer 1
Sandvik Product Support Technician Downer 2
Diamond Drill supervisor Boart 1
Diamond Drill Field fitter Boart 1
Diamond Driller Boart 7
Diamond Drill Assistant Boart 7
Total Contractor Staff   90

TABLE 16-12 CONTRACTOR PERSONNEL REQUIREMENTS – COSMO MINE

16.2 UNION REEFS UNDERGROUND – PROSPECT

16.2.1      INTRODUCTION

The mining of the Prospect underground mineral resource at Union Reefs is both technically and economically possible. The Prospect underground mineral resource provides the potential for a source of high-grade mineralization situated right on the Union Reef mine leases within a 2.0km haul to the ROM stockpile at the mill. This report has been compiled using the current Prospect Block model, updates to economic parameters as of December 31, 2015 and information and advice supplied by:

Wayne Chapman- 2012 Prospect Underground Mine Prefeasibility Study (Chapman 2013)

Ian McEnhill – Report Geotechnical Consulting Pty Ltd (GCPL)-Prospect-310113 (McEnhill 2013)

Newmarket Gold Management and Geology

The 2013 prefeasibility study recommends one mining method to be applied to the narrow vein high grade core Lode 40 lens. The mining method applied is bottom up, up hole stope and fill in three panel increments, with uphole retreat stoping of the sill pillars. The mine plan has also included some inferred mineral resource in the lower grade Lode 30 zone due to its proximity to the core mining activity. Material is only classified as ore and included within reserves if the indicated resource grade alone is sufficient to provide an average grade greater than or equal to the relevant cut-off grade.

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Since there are no material changes to the mining parameters and the mine plan, references are made within Section 16.2 of this document to the previous technical report, ‘Report on the Mineral Resources and Mineral Reserves of the Union Reefs Gold Project in the Northern Territory Australia’ with effective date December 31, 2012.

16.2.2      GEOTECHNICAL

16.2.2.1  Overview

Geotechnical Consulting Pty Ltd (GCPL) carried out a geotechnical review of the Prospect deposit in November 2012 (McEnhill 2013). The results of this review including the geotechnical zones, stope design parameters, development ground support requirements and decline location are described within the Prospect technical report with effective date December 31, 2012. These results are summarized below:

 

The Prospect deposit is hosted in a metasiltstone hangingwall and footwall, comprising greywacke and fissile shale rock units.

 

A zone of weathered siltstone is of variable depth (20 to 50 m), and stoping can commence below approximately 1150mRL.

 

The significant structure sets that have been geotechnically reviewed include the Union Fault 1 and 2 and two easterly dipping faults.

 

The Norwegian Geotechnical Institute’s (NGI) Rock Mass Quality (Q) System was used for assessment of Prospect ground conditions. Based on this, 31% of decline, access and mineralization development will occur in poor ground (Q<4) with the remaining 69% in fair and good ground (Q>4).

 

There are four geotechnical zones;

  o The weathered zone - classed as weak rock.
  o Fresh hangingwall zone – classed as strong to very strong rock
  o Fresh footwall zone - classed as strong to very strong rock
  o Mineralization zone - classed as strong rock.

Gold mineralization is associated with the Pine Creek Shear Zone, a 250m wide N-NW trending zone of deformation and shears that is the major conduit for mineralizing fluids in the region. The main mineralization zone of the mineral reserve comprises two lodes of which the narrow high grade 40 Lode occurs within the 400 Lode. There are stopes above the cut-off grade on the 30 and 31 Lodes that occur within the 300 Lode. The 200 and 300 Lodes are in the hangingwall of the 400 Lode at the southern end.

Figure 16.10 outlines the Prospect Lodes and the decline position as viewed from the south. The 200 and 300 Lodes are 100m to the south of the level access drives and are accessed from mineralization drives along the 400 Lode. The decline commences from the Lady Alice pit.

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Figure 16-11 outlines the 400 Lode stopes position within the geotechnical domains described above, also viewed from the south.

16.2.2.2  Mine Design Criteria

The stope design criteria are summarized as follows.

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16.2.3      MINE DESIGN

16.2.3.1  Mining Method Description

The selected mining method is a combination of up-hole benching and long hole open stoping, in a three lift bottom-up sequence. The two lower lifts are mined as up-hole bench stopes separated by rib pillars, with unconsolidated waste backfill placed in these stopes. The third lift is mined by longhole open stoping under a previously constructed cement rock fill (CRF) sill pillar in the floor of the overlying mineralization drive. There is a crown section under the previously mined Prospect Open Pit that will be mined from the 1135mRL at the end of mine life. Stopes at the northern and southern extremities of the lodes that are not under the pit will be mined as soon as access is complete along strike.

This method requires the decline to be advanced at least three levels prior to production commencing. Development driving along the mineralization utilizing a second jumbo will access the stopes at the extremities prior to production commencing on the bottom level of the sequence.

The production cycle for Benching and LHOS includes the following:

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16.2.3.2  Cut-off Grade

Table 16-13 shows the stoping cut-off grade assessment that has been undertaken for the Prospect underground mineralization. The cut-off grades have been calculated using current and estimated operating costs provided by Newmarket Gold. Processing costs are based on current costs for Cosmo Deeps mineralization (which is considered similar to Prospect mineralisationmineralization in terms of comminution and metallurgical response). Mining costs assume similar cost savings that have been experienced at the Cosmo operation between 2013 and 2015.

    Reserve Cut-off grade
Gold price $ 1,450 / oz  
Mill Recovery 93%  
Operating Costs    
 Mining $ 9,363,844 $ 93 / t
Processing $ 6,796,906 $ 23 / t
Total $ 47,189,086 $ 116 / t
Cut-off grade (Stoping) g/t Au 2.7

TABLE 16-13 PROSPECT STOPING CUT-OFF GRADE CALCULATION

A development-only cut-off grade has been calculated based on the assumption that all fixed site costs are covered by stoping (Table 16-14).

Cut-off Grade
Calculation
Units Reserve Cut-off
(Development)
Metal Price ($/oz) A$ 1450
Mining Dilution % 0%
Mining Recovery % 95%
Processing Recovery % 93%
AUD $/g Recovered A$/g 43.36
Power and Fuel A$/t 7.6
Ore Haulage A$/t 6.3
Processing A$/t 28.9
Prospect Maintenance A$/t 1.6
Total Operating Costs A$/t 44.4
Cut-off grade
(Development)
g/t Au 1.0

TABLE 16-14 PROSPECT DEVELOPMENT CUT-OFF GRADE CALCULATION

16.2.3.3  Material handling

As the Union Reefs mill facility is only 2km from the Prospect Mine, all underground production mineralization will be trucked via the decline and a surface haul road directly to a ROM stockpile adjacent to the mill.

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Initially waste will be hauled up the decline to the portal and tipped into the open pit to extend the area outside the portal. Prior to the commencement of production stoping requiring back fill waste is to be placed in the southern end of the Prospect pit. Waste rock from the lower areas of the mine will be used as backfill as the schedule requires. Backfill short falls will be trucked back underground from surface stockpiles.

16.2.4      MINE DESIGN GUIDELINES

16.2.4.1  Development and Stope Design

Figure 16-13 displays development commencing from the current Lady Alice open pit that indicates the proposed mining layout for the Prospect Mine viewed, from the footwall looking east.

Ore drives will be developed at 3.4m W x 4.0m H to allow for the use of production drills with full electronic boom movement, 3.0m 3 loaders, installation of ground support coverage with a single boom jumbo and the use of 1,067mm diameter ventilation ducting.

Table 16-15 summarizes the mineralization development mineral reserve. Mineralization development that fell below the development cut-off grade has been excluded from the mineral reserve. The proposed lateral and vertical development are shown in Figure 16-14, and waste development quantities are summarized in Table 16-16.

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Level

Meters
In situ
Tonnes
(t)
In situ
Grade
(g/t Au)
Recovered and
Diluted Tonnes
(t)
Recovered and
Diluted Grade
(g/t Au)
Recovered and
Diluted Metal
(Oz Au)
1135 502 18,624 3.38 20,487 3.07 2,021
1115 312 11,596 3.60 12,756 3.27 1,342
1095 398 14,797 3.44 16,277 3.12 1,634
1075 373 13,898 4.42 15,288 4.02 1,977
1055 368 13,706 3.76 15,077 3.28 1,592
1035 242 9,006 4.35 9,907 3.91 1,244
1015 228 8,468 2.83 9,315 2.57 770
995 217 8,075 2.48 8,883 2.25 643
Ore Development 2,639 98,171 3.58 107,988 3.23 11,223

TABLE 16-15 DEVELOPMENT MINERALIZATION INVENTORY

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Technical Report Newmarket Gold Inc.
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Prospect Development Meters (m) Waste Tonnes (t)
Decline 1,219 90,815
stockpiles 318 23,788
Level Access & Lode Cross-Cuts 632 44,437
Return airway (RAW) 116 8,643
Sumps 57 3,805
Total Lateral 2,266 166,215
RAR 170 7,082
Escape way 219 1,139
Total Vertical 389 8,221

TABLE 16-16 WASTE DEVELOPMENT QUANTITIES

A ventilation airway circuit will be established from the surface. The upper 90m of the return air rise (RAR) will be a 3.5m diameter raisebore with a shaft pre-sink through the surface weathered zone to allow pilot hole boring to commence prior to the 1109mRL access being completed. Subsequent legs of the RAR between each level will be developed as long-hole drill and blast raises.

An escape way system will be established to surface, utilizing each redundant decline stockpile. The first stockpile is 200m from the portal. An 84m long shaft, 1.5m in diameter will be raisebored from the surface with a shaft pre-sink through the weathered zone. This raise will initially be used for exhaust ventilation to provide primary ventilation to the two upper levels until the first RAR section is complete. The fan will then be removed and a ladder way installed to form the escape way. The escape way system will remain a down cast fresh air source for the remainder of the mine life.

Table 16-17 summarizes the underground mineral reserve for the Prospect deposit with respect to the layout shown in Figure 16-12. The stopes are not designed to a cut-off grade but are designed around the narrow high-grade zones associated with the 30, 40 and 41 Lodes. The stopes are designed to a minimum stope width of 2.0m with a 0.2m over-break on both hangingwall and footwall to allow for unplanned dilution. Stopes that fell below the stoping cut-off grade are excluded from the mineral reserve.

Level In situ
Tonnes
(t)
In situ Grade
(g/t Au)
Recovered and
Diluted Tonnes
(t)
Recovered and
Diluted Grade
(g/t Au)
Recovered and
Diluted Metal
(Oz Au)
1135
1115
1095
1075
1055
1035
1015
995
26,188
18,944
26,053
33,583
31,343
17,044
14,694
9,100
5.09
5.92
4.73
4.75
5.75
5.39
5.58
4.52
24,879
17,997
24,751
31,904
29,775
16,192
13,959
8,645
5.09
5.91
4.73
4.75
5.67
5.19
5.57
4.52
4,073
3,418
3,764
4,871
5,423
2,702
2,500
1,257
Mineralization
Development
176,949 5.22 168,101 5.18 28,009

TABLE 16- 17 STOPE INVENTORY

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16.2.5      MINE SERVICES

Ventilation, backfill, mine services, dewatering and other infrastructure requirements and methodology are the same as described in the Prospect technical report with effective date December 31, 2012. These results are summarized below.

Ventilation:

Backfill:

Dewatering:

16.2.2      MINING SCHEDULE

The Prospect scheduling strategy remains the same as described in the Prospect technical report with effective date December 31, 2012. Small changes to scheduled quantities only have resulted from the updates to the mineral reserve estimate detailed in Section 15.2.

The lateral and vertical development quantities are presented in Table 16-18 Development Schedule. There is 340m of decline and level development before mineralization driving commences. After a further 160m the mineralization driving on the second level commences.

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Prospect Development Meters
Year 3 Year 4 TOTAL
Decline
Stockpiles
Level Access & Lode Cross-Cuts
Return airway (RAW)
Sumps
Mineralisation Drive
909
213
477
94
36
1,988
310
105
155
22
21
869
1,219
318
632
116
57
2,856
Total Lateral 3,716 1,481 5,197
RAR & Escapeways 273 116 389
Total Vertical 273 116 389

TABLE 16- 18 DEVELOPMENT SCHEDULE

The production schedule for the Prospect Mine is presented in Table 16-19 mineralization Production Schedule.

Mining Method Year 3 Year 4 TOTAL
Tonnes (t) Grade (g/t
Au)
Tonnes (t) Grade (g/t
Au)
Tonnes (t) Grade (g/t
Au)
Au Oz
Production 18,741 4.16 149,360 5.31 168,101 5.18 28,009
Development 73,292 3.37 34,696 2.94 107,988 3.23 11,223
Total 92,033 3.53 184,056 4.86 276,089 4.42 39,232

TABLE 16-19 MINERALIZATION PRODUCTION SCHEDULE

Equipment and manpower requirements are the same as described in the Prospect technical report with effective date December 31, 2012.

  16.3 UNION REEFS OPEN PIT – ESMERALDA

16.3.1      INTRODUCTION

The Esmeralda deposit is located 7 kilometers to the south of the Union Reefs processing facility. It has had a significant amount of drilling completed over the years, which was updated with a drilling campaign in October-November 2015. This new drilling has been used to develop an updated mineral resource estimate (see Section 14), which has been used to generate this mineral reserve.

The Esmeralda deposit is located in undulating topography with a significant oxidized zone, which makes the deposit a highly ranked mining area to complement material from the sulphide ore sources such as the Cosmo Mine.

The deposit is partially overlain by the Australian Pipeline Association (APA) owned Amadeus Gas pipeline and an associated 50m wide protection corridor that runs north south through the mining lease. This gas pipeline limits the potential size of any economic open pit to extract mineralization from the deposit.

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Under the Energy Pipelines Act (NT) 2015 there are specific restrictions to mining in proximity to any gas pipelines, related to blast vibration and depth of any excavations adjacent to the pipeline. These restrictions have been strictly followed in the development of the mineral reserve pit designs. Discussions have commenced with the pipeline owner as to how operational risks would be managed during the project and if there is potential of changes to these restrictions conditional on a favourable site specific assessment and approvals from relevant government agencies.

16.3.2      GEOTECHNICAL

There is currently (March 2016) a geotechnical assessment program underway to determine final geotechnical parameters to be used for the Esmeralda open pits and test the currently applied design assumptions. This program consisted of drilling and logging 6 diamond holes through projected pit locations and submitting rock core samples for rock strength testing as inputs to analysis.

The assumptions used for the Esmeralda pit designs presented in this report were based on the performance of historically mined oxide open pits around the Union Reefs deposit area north of Esmeralda. Overall wall angles in oxidised zones of 50° with 5m wide berms every 15m resulting in an overall wall angle of 40° were used for the Esmeralda design. Generally oxide pits around Union Reefs have overall wall angles of around 44° (Crosby, et al. 2003), so the conservative approach used for the Esmeralda pits is seen as appropriate until the geotechnical assessment is complete.

16.3.3      MINE DESIGN

16.3.3.1  Mining Method Description

The proposed mining method at Esmeralda is conventional truck and excavator mining with mobile diesel fleet and blasting of mineralization and waste. While it is probable there is freely diggable material close to ground surface presently 100% of all rock is assumed to be required to be blasted.

A typical mining cycle would involve:

16.3.3.2  Open Pit Optimization

The optimization methodology adopted for the Esmeralda Project used the latest geological block models developed for the deposit. The block models contained mineable resource codes to which Whittle specific fields have been added for use in GEOVIA Whittle-4X optimization software. Whittle-4X utilizes the Lerchs-Grossman algorithm to provide the optimum mining pit shell for a given set of mining, metallurgy and economic parameters. While the deposit is classified into one mineral resource estimation there are two distinct mining areas, Esmeralda A (eastern lode) and Esmeralda B (western lode), which have been optimized separately.

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Technical Report Newmarket Gold Inc.
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Pit optimization for open pit mines using the Lerchs-Grossman algorithm is an industry-standard approach for defining an optimum open pit shape and development of a mining sequence. The methodology relied on the preparation of a 3D block model to represent all parts of the mineralization and host rock that can reasonably influence the pit shape. A single cash surplus for each block was estimated as the difference between the revenues derived from each block, at a nominated product price, and the costs required to realize the revenue from that block. For mineralized blocks with a grade above the economic cut-off grade, the net cash flow is positive reflecting the profit that can be made by mining and treating the block to recover the product. For all the other blocks, the net cash flow is negative, reflecting the cost of mining the block to access blocks of positive cash flow.

A summary of the various input parameters and assumptions used to develop the Whittle base case costs (mining, processing and selling costs) are presented in Table 16-20 and Table 16-21. The base costs were subsequently used to generate the optimum pit shells.

Whittle Processing Parameters

The processing costs presented in Table 16-20 are inclusive of transport costs to mill, administration and maintenance costs and actual milling costs.

Deposit Processing Cost
(A$/ t)1
Au Recovery (%)
Esmeralda A 30 90 for oxide rock, 85 for transitional and fresh rock
Esmeralda B 30 90 for oxide rock, 85 for transitional and fresh rock

TABLE 16- 20: WHITTLE PROCESSING PARAMETERS

Whittle Revenue Parameters

Deposit Base Au Price
(A$/troy oz.)
Royalty
(A$/troy oz.)
Whittle Au Price (A$/troy
oz.)
Esmeralda A 1,450 29 1,421
Esmeralda B 1,450 29 1,421

TABLE 16-21: WHITTLE REVENUE PARAMETERS

16.3.3.3  Open Pit Design

Pit designs were prepared using the optimized pit shells as templates. Mine design software including Surpac and MineRP Mine 2-4D were used to prepare practical pits which incorporate haul roads and ramps with the appropriate inter-ramp slope angles.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

The mining method for the deposits within the Esmeralda project assumed a configuration of Caterpillar 777F rear dump trucks and Hitachi EX1200 hydraulic excavators for removing the overburden and mining the mineralization. The bench height selected for the deposits were 15m with mining to be completed in three 5m flitches

Newmarket Gold has employed industry-standard safe operating tolerances to design final pits for the equipment assumed. The design parameters are presented in Table 16-22 and images of the completed designs for Esmeralda A are presented in Figure 16-15.

Parameter Unit Esmeralda A Esmeralda B
Overall Wall Angle deg. 40 on main ramp wall, 45 all other walls 40 on main ramp wall, 45 all other walls
Bench Height m 15m (3 x 5m flitch) 15m (3 x 5m flitch)
Berm Width m 5 5
Ramp Width m 14 14
Ramp Gradient   1 in 10 1 in 10
Mining Recovery (ore) % 95 95
Mining Dilution % 10 10

TABLE 16-22: PIT DESIGN PARAMETERS

Table 16-23 presents the inventory and operating cash flow analysis of the Esmeralda A pits.

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Technical Report Newmarket Gold Inc.
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Item Unit Esmeralda A
Waste t 845,000
Ore t 108,000
Strip Ratio   7.81
Cut-Off Grade (Au) g/t 0.72
Grade (Au) g/t 1.91
Ounces Mined (Au) oz. 6,650
Total Movement t 953,000
Mining Cost A$ 4,606,000
Processing Cost A$ 3,248,000
Revenue A$ 8,680,000
Total Cash A$ 826,000

TABLE 16- 23 ESMERALDA A PIT DESIGN RESULTS

Images of the completed design for Esmeralda B are presented in Figure 16-15.

Table 16-24 presents the inventory and operating cash flow analysis of the Esmeralda B pits.

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Technical Report Newmarket Gold Inc.
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Item Unit Esmeralda B
Waste t 493,000
Ore t 135,500
Strip Ratio   3.64
Cut-Off Grade (Au) g/t 0.72
Grade (Au) g/t 1.35
Ounces Mined (Au) oz. 6,000
Total Movement t 628,500
Mining Cost A$ 3,037,000
Processing Cost A$ 4,066,000
Revenue A$ 7,829,000
Total Cash A$ 726,000

TABLE 16-24 ESMERALDA B PIT DESIGN RESULTS

16.3.4      MINE EQUIPMENT ASSUMPTIONS

The mining equipment list assumed for the Esmeralda project is presented in Table 16-25. It was assumed that the fleet used for mining the Pine Creek open pits would be transferred when mining at that site was complete.

Plant Make Model Qty.
Drill Rig Atlas Copco ROCD65 2
Excavator Hitachi EX1200-6 2
Truck Caterpillar 777F 11
Track Dozer Caterpillar D10T 1
Grader Caterpillar 16 m 1
Water Truck Caterpillar 773FWC 1
Light Vehicle Toyota Toyota-Ops 4
Light Vehicle Toyota Toyota-TechServices 2
Lighting Plants Generic Generic 4
Stemming Loader Komatsu WA430-6 1

TABLE 16- 25: MINING EQUIPMENT LIST

16.3.5      MINING SCHEDULE

16.3.5.1  Scheduling Strategy

When developing a mining strategy, a systematic approach was undertaken with consideration to practical limitations and regulatory constraints.

In addition the mining strategy has been developed considering all mineral reserve deposits within NT Operations. The primary source of feed is from existing and proposed underground operations with open pit projects scheduled to supplement these mineralization sources. It is therefore required that the economics for the all deposits in the NT Operations are reported.

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Technical Report Newmarket Gold Inc.
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A joint strategy involving Pine Creek, Union Reefs and Cosmo areas allows for synergistic gains such as the sharing of assets and capital.

Cosmo will be the main ore source at the start of the mining schedule as it is the current operations. Ore will be sourced from Pine Creek, Union Reefs and Esmeralda in a logical procession that provides the best economic outcomes for Newmarket Gold.

The Esmeralda deposit has a short mine life with operations only expected to continue for less than one year. This short life means it will not be affected by the wet season which can slow open pit mining activities within the NT Operations.

16.3.5.2  Production Schedule

The production schedule for both Esmeralda A and B assumes a mining recovery of 95% with 10% ore dilution.

The key outcomes of the open cut production schedules are:

16.3.6      MANPOWER AND SUPERVISION

It was assumed that the Esmeralda project would be mined on a 24/7 basis to complete mining operations in the nine month window of the dry season thus limiting surface water management issues with active open pits.

It was assumed that the majority of production staff would work on a 2 weeks on 1 week off roster with 12 hour shifts alternating between day shift and night shift thus requiring a total of three crews. Technical support staff would work either 9 days on 5 days off, or 5 days on 2 days off and on either 10 or 12 hours/day. A total of 60 people would be directly employed during the peak of production.

16.3.7      RESTRICTIONS ON MINE DESIGN

Due to the proximity of the Amadeus gas pipeline, which transects the Esmeralda A deposit there are some restrictions on the design of open pits. There is a 50m exclusion zone (25m either side of the center of the pipeline) where no excavation can occur. Outside this exclusion zone all pits mined have to be at least three times the distance away from the pipeline corridor as their maximum depth e.g. the base of a 10m deep pit must be at least 30m away from the edge of the pipeline exclusion zone. The south wall of the southernmost of the Esmeralda A pits was flattened to meet this requirement.

During conversations with the APA they have indicated that there is potential to loosen this 1:3 batter constraint if the geotechnical risk/issues with wall stability to the pipeline can be managed to their satisfaction. The proposed plan is to backfill this pit with waste from other areas and reducing the duration that walls are exposed could aid this. At the time of reporting this work had not been completed so the conservative approach of strictly following the 1:3 batter angle has been used for calculating mineral reserves. It is a recommendation that this work is progressed with the potential to increase the mineral reserves and improve overall economics of the Esmeralda A area.

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Technical Report Newmarket Gold Inc.
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16.4 PINE CREEK OPEN PITS

16.4.1      INTRODUCTION

The proposed Pine Creek open pits are located on Newmarket Gold leases in the historic mining precinct to the west of Pine Creek.

Generally the deposits are located in slightly undulating topography adjacent to existing mined out and flooded open pit and rehabilitated waste dumps.

16.4.2      GEOTECHNICAL

The views and comments presented by Newmarket Gold geotechnical engineers have been based on previous pit design recommendations rather than from site specific geotechnical parameters or previous pit performance.

Due to limited data availability, the following have not been considered in the review:

As the pits are largely situated in the oxide zone the impact of these site characteristics is expected to be minimal.

The International Pit is 5km away from the Cox, Kohinoor and South Enterprise pits and is also situated in the Mount Bonnie and Burrell Creek Formations. It is understood that all pits have a similar geology, comprising interbedded siltstone, mudstone phyllites and greywacke units with dolerite sill intrusions, but the degree of oxidization, ground water conditions and residual properties are not well understood for any of the pits discussed.

Pit Design Overall Wall Angles

The overall pit wall angle of 37° used in these pit designs is comparable to the 43° used in the International Pit design. It is acknowledged that the new pits are generally shallower pits with shorter mine lives. Without complete geotechnical information it is not possible to fully understand how the pit walls will perform, but an overall wall angle of 37° is considered reasonable.

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Pit Design Batter Angles

The batter angles of the reviewed pit designs vary between 40° and 60° which may be at the upper end of the preferred range through the oxide material. The International design used slightly shallower angles, however, most of the pits reviewed in this report are shallower with shorter lives and it is understood they will only be mined in the dry season. This should minimize any concerns with batter stability.

Given the limited amount of geotechnical data available, the batter angles used in these designs are considered to be appropriate. In most pits there is some opportunity to alter the design once ground conditions are better understood, should it become necessary to control any instabilities.

Pit Design Catch Berms

Catch berms designed at 5 m wide are at the lower end of the preferred range, and any over break will reduce or limit their capacity to retain rock falls. The International pit uses 6m wide catch berms under 10-15m high batters. The pit designs reviewed in this report generally have 20m high batters but catch berms rarely exceed 6m. Due to the shallow depth of the pits and the short mining life of the pits, this is not expected to be an issue in weathered or oxidized material, however, where crests cannot be free dug, pre-splitting is recommended to minimize over break.

Waste Dump Design

Waste dumps designed with 35° lift slope angles are unlikely to have major stability issues, provided they are founded on suitable topography and materials

16.4.3      MINE DESIGN

16.4.3.1  Mining Method Description

The proposed mining method at Pine Creek is conventional truck and excavator mining with mobile diesel fleet and blasting of mineralization and waste. While it is probable there is some freely diggable material close to ground surface presently 100% of all rock is assumed to be required to be blasted for all pits except for International pit where there is backfill material within the pit footprint to be removed from when this pit was originally rehabilitated.

The bench height selected for the pits within the deposits are 10m with mining to be carried out in three flitches.

A typical mining cycle would involve:

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Technical Report Newmarket Gold Inc.
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In the case of the Pine Creek deposits it was assumed that Newmarket Gold would lease the equipment rather than purchase mining fleet given the short duration of the projects and employ production staff directly as an owner miner. This approach would need to be assessed closer to production commencing, as contractor mining may be more favorable. In this study all mining activities are assumed to be carried out by Newmarket Gold, establishment of facilities are expected to be sub-contracted under the supervision of Newmarket Gold.

16.4.3.2  Open Pit Optimization

The optimization methodology adopted for the Pine Creek deposits used the latest geological block models developed for each deposit. The block models contained mineable mineral resource codes to which Whittle specific fields have been added for use in GEOVIA Whittle-4X optimization software. Whittle-4X utilizes the Lerchs-Grossman algorithm to provide the optimum mining pit shell for a given set of mining, metallurgy and economic parameters.

Pit optimization for open pit mines using the Lerchs-Grossman algorithm is an industry-standard approach for defining an optimum open pit shape and development of a mining sequence. The methodology relied on the preparation of a 3D block model to represent all parts of the mineralization and host rock that can reasonably influence the pit shape. A single cash surplus for each block was estimated as the difference between the revenues derived from each block, at a nominated product price, and the costs required to realize the revenue from that block. For mineralized blocks with a grade above the economic cut-off grade, the net cash flow was positive reflecting the profit that can be made by mining and treating the block to recover the product. For all the other blocks, the net cash flow was negative, reflecting the cost of mining the block to access blocks of positive cash flow.

Whittle-4X structure arcs were used to define the precedence of block removal, such that a block cannot be considered for mining unless certain overlying blocks are also mined. This effectively defines the slope geometry for an open pit operation.

The optimization then consisted of finding the combination of positive and negative cash flow blocks, consistent with the slope precedent rules, which accumulate to a maximum positive cash flow.

A series of pit optimization shells are produced in Whittle, which are regarded as concentric pits, each generating the maximum undiscounted cash surplus for the set of economic parameters used to develop the optimized shell. The shells are created by varying the product price, but once defined they are all evaluated at the base case product price.

Only blocks with mineral resource categories of 1 (Measured) and 2 (Indicated) were considered as potential mineralization blocks in the generation of the optimum pit shell. A process flow of the optimization logic and mineral reserves estimation process is summarized in Section 14.4.

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16.4.3.3  Open Pit Mine Design

Pit designs were prepared using the optimized pit shells as templates. Mine design software including Surpac and MineRP Mine 2-4D were used to prepare practical pits which incorporate haul roads and ramps with the appropriate inter-ramp slope angles.

The mining method for the deposits within the Pine Creek area has assumed a configuration of Caterpillar 777F rear dump trucks and Hitachi EX1200 hydraulic excavators for removing the overburden and mining the mineralization. The bench height selected for the deposits were 20m with mining being carried out in three flitches. Newmarket Gold has employed industry-standard safe operating tolerances to design final pits for the equipment assumed. The design parameters used are presented in Table 16-26.

Parameter Unit Cox Kohinoor South Enterprise International
Wall Angle deg. 55 40/60* 55 50/55^
Bench Height m 20 20 20 20
Berm Width m 5 10/6* 5 5
Ramp Width m 15 15/12* 15 15
Ramp Grade % 10 10 10 10
Mining Recovery % 95% 95% 95% 95%
Mining Dilution % 15% 15% 15% 10%

* Applies to bottom 20m of pit ^ Applies to Eastern side of pit

TABLE 16-26: PINE CREEK PIT DESIGN PARAMETERS

Newmarket Gold validated the pit designs for Cox, International, Kohinoor and South Enterprise deposits by reporting against the Surpac mineral resources. These reports were constrained so only material above the cut-off grade applied in Whittle was classed as ore.

The cut-off grades was calculated using the processing cost and total metal recovery, mining dilution factor, and the gold price and represents the minimum grade for each mining block that can be economically extracted.

An image of the completed Kohinoor pit design is presented in Figure 16-16 and Table 16-27 presents the pit inventory and associated operating cash flow evaluation.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Item Unit Kohinoor
Waste t 792,000
Ore t 128,000
Strip Ratio   6.17
Cut-Off Grade (Au) g/t 0.91
Grade (Au) g/t 2.40
Ounces Mined (Au) oz. 9,900
Total Movement t 920,000
Mining Cost A$ -4,418,000
Processing Cost A$ -4,269,000
Revenue A$ 12,905,000
Total Cash A$ 4,218,000

TABLE 16-27: KOHINOOR PIT DESIG N RESULTS

The completed Cox pit design is presented in Figure 16-17 Figure and Table 16-28 presents the pit inventory and associated operating cash flow evaluation.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Item Unit Cox
Waste t 451,000
Ore t 132,000
Strip Ratio   3.41
Cut-Off Grade (Au) g/t 0.91
Grade (Au) g/t 1.62
Ounces Mined (Au) oz. 6,900
Total Movement t 583,000
Mining Cost A$ -2,798,000
Processing Cost A$ -4,397,000
Revenue A$ 8,970,000
Total Cash A$ 1,775,000

TABLE 16-28: COX PIT DESIGN RESULTS

The completed International pit design is presented in Figure 16-18 and Table 16-29 presents the pit inventory and associated operating cash flow evaluation.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Item Unit International
Waste t 1,650,000
Ore t 861,000
Strip Ratio   1.92
Cut-Off Grade (Au) g/t 0.93
Grade (Au) g/t 1.30
Ounces Mined (Au) oz. 35,900
Total Movement t 2,511,000
Mining Cost A$ -12,054,000
Processing Cost A$ -29,016,000
Revenue A$ 44,581,000
Total Cash A$ 3,511,000

TABLE 16-29: INTERNATIONAL PIT DESIGN RESULTS

The completed South Enterprise pit design is presented in Figure 16-19 and Table 16-30 presents the pit inventory and associated operating cash flow evaluation.

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Item Unit South Enterprise
Waste t 1,251,000
Ore t 123,000
Strip Ratio   10.16
Cut-Off Grade (Au) g/t 0.91
Grade (Au) g/t 2.36
Ounces Mined (Au) oz. 9,350
Total Movement t 1,374,000
Mining Cost A$ -6,732,000
Processing Cost A$ -4,094,000
Revenue A$ 12,180,000
Total Cash A$ 1,354,000

TABLE 16- 30: SOUTH ENTERPRISE PIT DESIGN RESULTS

16.4.4      MINE EQUIPMENT ASSUMPTIONS

The mining equipment list assumed for the Pine Creek area is presented in Table 16-31.

Plant Make Model Qty.
Drill Rig Atlas Copco ROCD65 2
Excavator Hitachi EX1200-6 2
Truck Caterpillar 777F 11
Track Dozer Caterpillar D10T 1
Grader Caterpillar 16 m 1

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Technical Report Newmarket Gold Inc.
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Plant Make Model Qty.
Water Truck Caterpillar 773FWC 1
Light Vehicle Toyota Toyota-Ops 4
Light Vehicle Toyota Toyota-TechServices .2
Lighting Plants Generic Generic 4
Stemming Loader Komatsu WA430-6 1

TABLE 16- 31: MINING EQUIPMENT LIST

16.4.5      MINING SCHEDULE

16.4.5.1  Scheduling Strategy

The production schedule also incorporates a mining recovery of 95% with 10 and 15% mineralization dilution for the International pits and all other areas respectively.

16.4.5.2  Production Schedule

The key outcomes of the open cut production schedules include:

The open pit production is scheduled to produce 1.24Mt of mineralization with an average grade of 1.58g/t Au

16.4.6      MANPOWER AND SUPERVISION

It was assumed that the Pine Creek area would be mined on a 24/7 basis to maximize productions in the nine month window of the dry season thus limiting surface water management issues with active open pits. However given the proximity of the mining operations to the existing Pine Creek Township, this would have to be possibility reviewed in terms of impact to the local population.

It was assumed that the majority of production staff would work on a 2 weeks on 1 week off roster with 12 hour shifts alternating between day shift and night shift thus requiring a total of three crews. Technical support staff would work either 9 days on 5 days off, or 5 days on 2 days off and on either 10 or 12 hours/day. A total of 60 people would be directly employed during the peak of production.

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17 RECOVERY

17.1 UNION REEFS GOLD PLANT

Any ore production from the Cosmo Mine is processed at the Union Reefs processing facility, which is approximately 67km from the project via the haul road and public roads. Details of the Union Reefs processing plant are below;

The Union Reefs Gold Mine was commissioned in 1994 by Acacia Resources Ltd (Acacia), an entity spun out of the Shell/Billiton Group. AngloGold Australia Ltd (AngloGold) acquired the mine through its successful takeover bid for Acacia in December 1999. Until late 2000, Union Reefs formed one half of what was known as AngloGold’s Pine Creek operations, which also included the smaller Brocks Creek Project.

The Union Reefs CIL treatment plant was commissioned with a throughput capacity of 1.25Mtpa. It included a gravity circuit to extract coarse gold. It was designed by Kinhill and commissioned in December 1994. JR Engineering carried out an upgrade in 1998 that involved the installation of a tertiary crusher, second ball mill, and two additional leach tanks.

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The plant currently has a maximum capacity for 2.5mtpa, depending on ore types and is configured with three-stage crushing and two single-stage milling circuits. Prior to the plant being placed on care and maintenance in 2003, the milling rate at Union Reefs was typically 335tph at a P80 of 75µm. Plant availability was typically 96-98%.

In August 2004, and before they were acquired by GBS Australia, the Burnside JV partners purchased the Union Reefs Gold Project for A$4 million on a walk-in, walk-out basis.

In August 2006, GBS Australia re-commissioned the Union Reefs plant on the larger of the two mills while leaving the other smaller ball mill in a care and maintenance state. The first source of feed material was low-grade stockpiles from Cosmo Mine and an alluvial tailings deposit from the Union Reefs site. Following commissioning, mineralization was sourced from a blended mix of oxidized and fresh underground and open pit mines.

In June 2010, Crocodile Gold announced commercial production for the Union Reefs plant. In the period from December 2009 until June 2010 the plant was in commission. Production at Union Reefs has continued since commercial production was declared in 2010.

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17.2 UNION REEFS PLANT OPERATIONS

For more details on the operations of the Union Reefs processing facility please refer to section 13.1. A plan showing the location of required infrastructure is shown below and a detailed flow sheet is also shown. Details on the operation of the plant is described in detail in section 13.

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A schematic flow sheet of the plant as currently configured is shown below.

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17.2.1      UNION REEFS PLAN POWER/WATER/MATERIAL

The processing facility at Union Reefs requires several important consumables to continue operations. In the past there have been no issues with being supplied this material but below is a summary.

17.2.1.1  Water

Water is sourced from site; there is ample water currently located within the tailing storage facility within the Crosscourse pit, other sources could be utilized if required such as Dam A or Dam C, which are also located on the Mineral Lease. Crosscourse pit currently contains around 2,000 megalitres of water while Dam A holds 270 megalitres and Dam C around 891 megalitres. It is estimated that the milling operation uses around 1,300 megalitres a year, most of which can be recycled from the tailings facility.

17.2.1.2  Power

Power is currently supplied under contract from the Northern Territory Government owned Power and Water Corporation (Power and Water Corporation). Power consumption is currently low while the mill is not on full production but in previous years when the mill has been in full production, Power and Water Corporation was able to supply sufficient power requirements. Up to 94 million KWhrs were supplied to Crocodile Gold and could be again when required.

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17.2.1.3  Processing Materials

Other consumables used in the processing facility include Diesel Fuel, Lime, Cyanide, Sodium Hydroxide, Flocculent, Antiscalants. Carbon, Grinding Material, Hydrochloric Acid and Liquid Oxygen. All are and have been provided by local suppliers based in Darwin or surrounding areas. In the four years of production at the Union Reefs facility no supply issues have been encountered.

17.2.2      UNION REEFS PLANT INFRASTRUCTURE

17.2.2.1  Laboratory

The laboratory is used for sample preparation and for analysis of gold in solution only. All solids and carbon samples are assayed at an external laboratory. The laboratory is complete but it is not set up to cater for grade control samples or metallurgical test work. External contract laboratories provide these services. This process is currently under review to suitability to current mineralization types, some work may be required using fire assay, which can be completed in Pine Creek at NAL.

17.2.2.2  Buildings

The offices are all re-locatable buildings that are all fully functional. The offices have a complete IT system with on site server, computers, phone systems, office furniture and other equipment to ensure that all the staff can function productively. A preventative maintenance schedule is in place to protect the value of the assets.

17.2.2.3  Accommodation

Union Reefs staff is accommodated at the Cosmo Mine Village near the Cosmo Mine. The buildings are all fully functional transportable type structures. The whole camp is maintained and there is a planned maintenance schedule in place to protect the value of the asset.

17.2.2.4  Maintenance and stores

The workshops and facilities are typical of a remote gold plant in Australia and are all fully functional with the capacity and ability to perform the majority of the work required on site. The spare parts holding is extensive and is maintained with the assistance of the Pronto Accounting System that has been installed on all Newmarket Gold operations. The planned maintenance schedule and area costing is also handled through the Pronto Accounting System.

17.2.2.5  Drawings and Intellectual Property

All plant drawings are on CAD files with hard copies stored on site.

17.2.3      UNION REEFS PLANT HISTORICAL PERFORMANCE

Total gold production from the commencement of operations at Union Reefs in 1994 to June 2003 was estimated by AngloGold to be 953,294oz from the treatment of approximately 21Mt of ore grading an average 1.5g/t Au. Total production since November 2009 when Crocodile Gold commenced operations has totaled 445,700oz from 6.94Mt at an average grade of 2.0g/t Au.

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These production figures were achieved on a variety of ore types, including oxide, sulphide and low-grade ore from Cosmo, Brocks Creek, various open pits and historic stockpiles from deposits such as Moline, Golden Dyke and Glencoe.

The historical performance of the Union Reefs plant indicates that it was a reliable, efficient and low cost ore processing facility. It has previously demonstrated a capability to reliably and efficiently process in excess of 2.5Mtpa of free milling ore from a blend of oxide, transition and fresh ore types. The plant also has a high level of flexibility and can be operated efficiently at a lower throughput rate with the use of only one of the installed ball mill circuits.

  Milled
Period Tonnes Grade
g/t Au
Ounces
Total 2009   29,000 1.8   1,600
Total 2010 1,855,000    1.5 92,300
Total 2011 1,886,000    1.2 73,100
Total 2012 916,000 1.4 40,700
Total 2013 720,000 3.6 82,200
Total 2014 809,000 3.3 85,900
Total 2015 725,000 3.0 69,900
Total 6,940,000     2.0 445,700    

TABLE 17-1 PRODUCTION FIGURES FOR UNION REEFS PLANT SINCE RESUMPTION OF OPERATIONS IN. 2009

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Technical Report Newmarket Gold Inc.
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18 PROJECT INFRASTRUCTURE

18.1 INTRODUCTION

The following infrastructure review relates to all sites and requirements as outlined for each area. For all operations outlined below, the processing facility utilized would be the Union Reefs facility located on MLN1109. More details on this plant can be found in Section 17.

18.2 COSMO MINE

The Cosmo Mine has been operating for approximately four and half years and has all major infrastructure in place.

To mine the Cosmo mineral reserve, as identified in this technical report, the following changes or additions will be made to the current infrastructure:

There is no other infrastructure required for the execution of this schedule.

18.3 UNION REEFS UNDERGROUND – PROSPECT

The Prospect Underground mineral reserve is located on the Union Reefs mineral lease and is in close proximity to the processing facility and supporting infrastructure.

18.3.1      ONSITE INFRASTRUCTURE

At the present time there is limited mining infrastructure onsite in the immediate vicinity of the Prospect and Lady Alice pits. The Union Reefs mill site is within 2km of the Prospect deposit. At the Union Reefs site there is a large machinery workshop, which would be utilized for major and minor repairs on mining machinery.

Onsite infrastructure will be required to enable the supply of mine services including;

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18.3.2      OFFSITE INFRASTRUCTURE

All other required infrastructure will be situated at the Union Reefs mill site. The current administration offices, workshops, change facilities and ablutions will be utilized for the Prospect project. Accommodation for the work force is available at the Newmarket Gold Cosmo Village camp.

18.4 UNION REEFS OPEN PIT – ESMERALDA

The Esmeralda deposit is located 7km to the south of the Union Reefs processing facility. The mineral reserves included in this technical report are to be extracted using open pit mining techniques so the required infrastructure is outlined below.

18.4.1      POWER

Power can be accessed from the local Pine Creek grid, which is located close by or alternatively; power may be generated on site using generators. Power requirements will not be significant.

18.4.2      WORKSHOPS

It is envisaged that service facilities would be shared between all the Newmarket Gold mine sites. Mobile workshops will be erected for the minor servicing and refueling of equipment. There is a large workshop available for use on the Union Reefs mineral lease for major repairs as required.

18.4.3      OFFICE

Administration will be managed from the existing office facilities at Union Reefs, and will accommodate management, administration and technical staff. All telephone, data and office facilities exist at the Union Reefs offices.

18.4.4      CAMP FACILITIES

Employees and contractors conducting work in the open pit mining areas could be accommodated at the Cosmo Village camp.

Some local personnel, however, are expected to opt to live in private residences in Pine Creek, Adelaide River or Katherine and will commute to the mine sites.

18.5 PINE CREEK OPEN PITS

Proposed infrastructural developments for the open pit mining include the expansion of the new pits, ROM pads and waste rock dumps for each deposit.

Ancillary mine plan components include haul roads connecting the pits and mine workings with waste rock dumps, ROM pads, site office and access roads.

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18.5.1      POWER

Electrical power supply to the mining operations and associated infrastructure will be from NT Power Water via the Pine Creek Grid. Power lines are present on access roads from Pine Creek to the Newmarket Gold mining tenements. Where power is not available, portable diesel generators will be used to provide power for temporary site facilities.

18.5.2      WORKSHOPS

It is envisaged that service facilities would be shared between all the Company’s mine sites. Mobile workshops will be erected for the minor servicing and refueling of equipment.

18.5.3      OFFICE

Administration will be managed from the existing office facilities at Union Reefs, and will accommodate management, administration and technical staff. All telephone, data and office facilities exist at the Union Reefs offices.

18.5.4      CAMP FACILITIES

Employees and contractors conducting work in the open pit mining areas will be accommodated at the Cosmo mine village and commute to the mine by company bus.

Some local personnel, however, are expected to opt to live in private residences in Pine Creek, Adelaide River or Katherine and will commute to the mine sites.

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19 MARKET STUDIES AND CONTRACTS

19.1 MARKETS

Gold dore produced by Newmarket Gold is currently shipped to Perth Mint for smelting. On notification of produced amounts of gold dore, Newmarket Gold notifies Auramet Trading LLC (“Auramet”) of the upcoming gold shipment and deals for future delivery of gold to the mint. During this process a given number of ounces will be included in the sale amount for delivery directly after smelting has occurred. Generally, a period of one week is required for this process to occur.

Auramet is a New Jersey based company, which specializes in the sale of both base metals such as copper, nickel and zinc and precious metals such as gold, silver, platinum and palladium.

During the smelting process the mint can extract other minerals. The main economic mineral that is recovered is silver, which is sold to Perth Mint.

19.2 GOLD PRICE

To determine the Australian denominated gold price to use in the mineral resource and mineral reserve calculations, reference was made to publicly available price forecasts by industry analysts for both the gold price in US dollar terms and the A$/US$ foreign exchange rate.

This exercise was completed in December 2015, and yielded the following average gold forecast prices and corresponding average forecast US$:A$ FX rates.

For mineral reserve purposes, a US$1,100/oz gold price was used and an FX rate of $0.76 for an approximate Australian dollar gold price of A$1,450 per ounce.

For mineral resource purposes, a US$1,125/oz gold price was used and an FX rate of $0.75 for an approximate Australian dollar gold price of A$1,500 per ounce.

The average US$ gold price per ounce for the last three years was as follows:

  2013 - US$1,411
       
  2014 - US$1,266
       
  2015 - US$1,160

19.3 MATERIAL CONTRACTS

The following is a summary of the major contracts related to the Cosmo Mine and Northern Territories Operations area. Newmarket Gold has several relevant contracts in place to assist with mining and development of the Cosmo Mine and the operating of the Union Reefs processing facility.

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Technical Report Newmarket Gold Inc.
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19.3.1      POWER SUPPLY

Power in the Northern Territory is generated and distributed by three government owned corporations including Territory Generation (production), Power and Water Corporation (transmission and distribution) and Jacana Energy (retailer).

Territory Generations largest power station is the natural gas fired station situated at Channel Island in Darwin, which has an installed capacity of 310MW. A further 26.6MW power station exists at Pine Creek. The interconnected system is linked by a 132kV transmission line from Darwin to Katherine. A 66kV line connects the Union Reefs processing facility, Brocks Creek, Cosmo Mine and the Cosmo camp to the Pine Creek Township.

To maintain the supply of power a current Connection Agreement is required from Power and Water Corporation and a Sale of Electricity Agreement from Jacana Energy.

The current Sale of Electricity Agreement commenced July 1, 2015 for 12 months, with an option to extend for six months.

Power costs for commercial entities in the Northern Territory are significantly higher than for other localities around Australia. However, there is currently only one supplier of grid power in the Northern Territory, which limits the opportunity to gain discounts. Several private power generators are looking to establish markets in the Northern Territory, which will have an impact on current prices.

It should be noted that the Northern Territory is a large land area with limited population, therefore economies of scale for power production means a higher cost of production than for other areas of the country. This is reflected in the power costs supplied in the Northern Territory.

19.3.2      DOWNER EDI LIMITED

Downer has been awarded an underground mining contract for an initial term of two years, completing in March 2016, with the ability to be extended by one year at any stage of the term. The option to extend for a third year was exercised in June 2015.

Downer is a leading provider of engineering and infrastructure management services to customers operating in market sectors including Minerals & Metals, Oil & Gas, Power, Road & Rail Infrastructure, Telecommunications and Water. Downer is listed on the Australian Securities Exchange and employs more than 20,000 people in Australia, New Zealand and the Asia Pacific region.

It is known that the terms of the contract are within industry norms for similar types of contracts at other sites.

19.3.3      FAWCETT CATTLE COMPANY

Ore haulage from the Cosmo Mine to the Union Reefs mill is performed by the Fawcett Cattle Company. They currently supply heavy haulage road trains, which have a carrying capacity of around 100 tonnes each. These Quad Haulage Trucks transport the Ore around 67km from the Cosmo Mine to the Union Reefs Mill on a 24 hour continuous shift.

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The Fawcett Cattle Company is an Adelaide River based contractor specializing in livestock and bulk haulage, with over 30 years of experience.

The term of the haulage contract is for two years with a one year extension and is for the supply of Ore haulage from the Cosmo Mine.

This contract was placed out for tender in 2014. The Fawcett Cattle Company tendered for the contract and won on several fronts, which included the overall cost of haulage. It is therefore believed that the contract rates paid are within the norm for similar operations.

19.3.4      CATER CARE GROUP

The Cater Care Group provides camp and messing support at the Cosmo Village. They are one of Australia’s leading providers of contract catering, accommodation and facility management services. Cater Care were established in 1999 and employs around 1,200 staff both nationally and overseas.

The term of the Cater Care contract is for two years with a one year extension, the contract was signed in July 2014.

It is believed that the contract rates for the Cater Care Group are within the norm of other operations. Cater Care also supply camping and messing support to other mines in close proximity to Newmarket Gold operations; this will have an effect on the supply cost for this contract with economy of scale savings.

Mr. Edwards has reviewed the contracts as outlined above and the results support the assumptions contained within the technical report.

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Technical Report Newmarket Gold Inc.
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20 ENVIRONMENTAL STUDIES, PERMITTING AND SO CIAL OR COMMUNITY IMPACT

Newmarket Gold has an internal department specializing in environmental permitting and management. This department is responsible for submitting all relevant documents to the Northern Territory Government departments. The reports, such as the MMPs, are a statutory reporting requirement under the Mining Management Act.

Newmarket Gold also engages the local community to talk through plans or issues regarding mining and exploration. Regular meetings have been held allowing the local community an opportunity to speak with the Company’s management.

In 2014 and 2015 Crocodile Gold and later Newmarket Gold have been involved in consultations with Traditional Owners, facilitated by the Northern Land Council. The consultations, separately involved the Jawoyn Traditional Owners regarding finalizing the draft ILUA for the proposed Esmeralda Project agreement; and the Kamu Traditional Owners regarding finalizing the draft Burnside ILUA.

Newmarket Gold currently holds the following authorisations.

Northern Territory Operations
Project Area Authorisation Number
Maud Creek 0524-02
Moline 0525-02
Fountain Head / Tally Ho 0526-01
Brocks Creek 0528-01
North Point & Princess Louise 0530-01
South Burnside 0531-02
Pine Creek 0538-01
Union Reefs 0539-03
Cosmo-Howley 0536-03

TABLE20-1 LIST OF CURRENT MMP’S FOR NEWMARKET GOLD NT OPERATIONS

Any changes to the approved mining will need an amendment to the approved MMP, which is submitted to the DME to be assessed under the Mining Management Act. If the DME determines that the amendment triggers actions that may need further assessment, then the amendment is referred to the Northern Territory Environmental Protection Authority (EPA) for further assessment. Details of the assessment process are detailed below.

All new projects that do not hold an Authorization to Mine or have an Authorization but the area is in care and maintenance will need to submit the NOI to the DME and follow the approvals process as set out below (taken from (NTEPA 2015)

20.1 NOTICE OF INTENT (NOI)

An NOI is submitted to the Department of Mines and Energy for consideration. This document is reviewed and a request for further information may be requested. An NOI outlines what activities the Company plans to undertake and is detailed enough for the Department to assess the potential environmental and community impacts.

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The NOI is assessed and if the projects potential environmental impacts are not significant then approval may be granted at this point. If this is the case, the environmental responsibilities will be managed against a MMP regulated under the Mining Management Act. If the environmental impact is significant or triggers certain activities then it will be forwarded to the EPA for further assessment. The EPA will determine the next steps, as it could be subject to an Environment Impact Assessment (EIA); a Public Environment Report (PER); or no further assessment required and is sent back to the DME with environmental management recommendations to be included into the MMP.

20.2 ENVIRONMENTAL IMPACT ASSESSMENT (EIA)

If an EIA is determined to assist in assessing environmental impacts, this means that the project triggers certain aspects that are considered significant either for site-specific issues, offsite issues and conservation values and/or the nature of the proposal.

Key points on an EIA are:

The EIA approval process generally is longer than the PER process, up to four years, depending on the project complexities and the response to requests for further information and community concerns.

20.3 PUBLIC ENVIRONMENTAL REPORT (PER)

If a Public Environment Report (PER) is determined, this means the project’s environmental impacts are considered significant but limited in extent. It is not a precursor to an Environmental Impact Study (EIS) hence the decision on a PER or an EIS has to be made on receipt of the NOI.

Key points on a PER:

The PER approval process generally is shorter than the EIA process, up to two years, depending on the project complexities and the response to requests for further information and community concerns.

See Figure 20-1 flow chart from NT EPA (NTEPA 2015) showing approval process for a project.

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20.4 NAF/PAF ENVIRONMENTAL TEST WORK PROCEDURE: COSMO MINE

The management program is designed to minimize the environmental impact of stockpiled acid generating waste material during and after mining operations cease. Procedures have been formulated to ensure that waste material, which is (or potentially) acid forming and/or contains elevated arsenic values, is stabilized and contained.

Test work is done to obtain a representative sample of future mined material from multiple rock type domains in order to assess the acid mine drainage generation capacity. Rock material hauled to the surface is then be classified as either non-acid forming (NAF) or potential acid forming (PAF) material and treated accordingly.

20.4.1      SAMPLE PROCEDURE

Composite samples of diamond drill core are collected from within the same lithological unit/interval. Only half of the core is sent for sampling in waste rock whereas mineralization zone samples that have already been sampled for gold (half cored) must be cut further (quarter core) to retain some core for reference. For larger units of a continuous lithology (e.g. dolerite), approximately 5kg of core is sampled for every 20m downhole.

Hole ID Sample
ID
mFrom mTo Sample
Type
Date
Sampled
Sampled
By
Rock
Type
North East mRL
CW92008 H207701 4 14 QCORE 28/08/2012 E. Bew Psl 1609.6 4902.6 923.25
CW92008 H207702 45.5 50.5 HCORE 28/08/2012 E. Bew Pdz 1608.89 4863.34 928.63
CW92008 H207703 65.5 70.5 HCORE 28/08/2012 E. Bew Pdz 1609.27 4843.42 930.97

TABLE 20-2 EXAMPLE OF NAF/PAF SAMPLE COMPOSITE INFORMATION TO BE COLLECTED – HOLE CW92008

Samples were tested for the following properties.

Type Analytes
Type 2 NET ACID PRODUCING POTENTIAL (NAPP & APP kg H2SO4/t)
Type 3 Total Metals ARSENIC (mg/kg) & SULFUR (%)
Type 4 NET ACID GENERATION (NAGpH Units & kg H2SO4/t)
Type 5 Acid Neutralizing Capacity (ANC, kg H2SO4/t)

TABLE 20-3 TYPE AND ANALYTES TESTED

These tests determine the potential of the rock to react to create acid mine drainage also if they have the potential to neutralize acid mine drainage such as some calcium carbonate rock types (Dolomite).

The tested samples are treated as being located at the midpoint of the sample interval along the drill hole and are then plotted on a plan/section and attributed to a particular rock type domain (see Figure 20-2 and Figure 20-3).

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Sampling of a rock domain is deemed to be relevant approximately 50m vertically above and below a hole as the acid generating capacity of rock can change spatially.

In 2012 five holes were drilled in the Cosmo deposit that were tested for NAF/PAF classifications. These are holes are CE102501, CEGT97008, CGT0003, CE84037 and CW92008. Their spatial locations are shown in Figure 20-3 below. Since there have been no lithological changes in the underground operation, no further sampling has been required to date

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Test work to date has shown that the Dolerite and Dolomite rock units are NAF. All other rock types including the Greywacke/metasediments and Graphitic Mudstone have been shown to be a combination of NAF and PAF. As such all metasediments at this stage are treated as being PAF.

All waste PAF material is dumped in the Cosmo Open Pit or used as stope backfill material as Cosmo Deeps in currently a ‘waste negative’ operation, meaning all waste material will be returned to the mine as backfill material into open stopes, alternative material will be required for backfilling to be completed as there is insufficient waste material to use from operations.

NAF material is used at the surface for site works such as bunding, roads and pads.

20.5 ENVIRONMENTAL ISSUES & LIABILITIES

20.5.1      COSMO MINE

The Cosmo Mine Authorization required that an MMP be developed to manage the project’s environmental activities. This document broadly details: the project facts; management structure; current environmental features; commitments held by Newmarket Gold which were determined as part of the approval to mine; and reporting, monitoring and rehabilitation requirements. The MMP is updated every four years or as the project changes, the MMP will require updating reflecting the changes to the project. The DME audit the MMP every second year.

Each year the Cosmo Mine requires an Operational Performance Report submitted detailing Newmarket Gold’s performance in meeting the commitments. The Company must ensure it meets these commitments or risk penalties from the DME.

The types of commitments within the MMP are varied but a summary below outlines some of these (Jensen 2012).

Cosmo Mine operates with a Waste Discharge License (WDL 180-02) regulated under the Northern Territory Water Act and Northern Territory Waste Management Pollution Control Act, which require monitoring and management of site active and passive release of water during the wet season.

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The Cosmo Howley site has a positive water balance (more water enters site than leaves site naturally) and the challenge for water management at this site is ensuring the water on site is managed.

Newmarket Gold has a number of methods for reducing water inventories on site. These include wastewater treatment, discharge, and evaporation by irrigation technology.

Wastewater discharge and monitoring from the Cosmo Howley site has always been a logistical and technical challenge. To maximize wastewater discharge monitoring, a telemetered gauging station was installed at Creek six monitoring and reporting point. Data can now been accessed via the internet to ensure wastewater discharge meets license conditions and assist in planning for ongoing water releases.

The caustic soda (sodium hydroxide) treatment plant was used during the 2013/14 wet season (October to April) to treat operational and legacy wastewater prior to release from site. Approximately one gigalitre (1,000 million liters) of wastewater treated in 2013/14 wet season. Ongoing treatment will occur each year to support operations and manage the site wide water balance.

The Mine Water Dam constructed over 20 years ago by previous operators has an irrigation sprinkler system to augment evaporation of the dam’s wastewater. Extensions to the evaporation system are likely to progress in 2015-16.

Newmarket Gold is committed to continually improve wastewater management a part of the WDL requirements. Preliminary investigations have commenced with trailing other treatment technologies in conjunction with the caustic treatment to improve efficiencies.

GHD Consultants have developed final landform plans to rehabilitate the two sulphide waste rock dumps (WRD) at Howley, namely called Howley WRD and Mottrams WRD. Rehabilitation works are likely to commence in 2016, which will significantly reduce the outstanding liabilities for the Cosmo/Howley area.

20.5.2      UNION REEFS AREA

The area surrounding Union Reefs comprises a number of open cut pits and underground adits and shafts. Several small pits have been bunded and currently do not require significant rehabilitation. During 2011 and 2012 Crocodile Gold conducted significant surface exploration at Union Reefs, however, during the latter part of 2012 all drillhole sites were rehabilitated with all collars cut back below the surface, all diamond drill sumps backfilled and all tracks not required were ripped for seeding.

Historically there has been extensive drilling completed in the area, the legacy of which Crocodile Gold inherited when it took over the tenements in 2009. In addition to this, in 2011/2012 Crocodile Gold conducted a drilling program mainly focused around the historic Prospect, Crosscourse and Lady Alice open pits. A rehabilitation Environmental Management Plan was created and included in the URPA Mine Management Plan, in an effort to prevent or minimize adverse impacts on the environment. This entails procedures such as initially locating drill pads in a manner that minimizes disturbance to an area; the implementation of a clearing permit system; plugging the collars of drillholes to prevent erosion and mixing of ground and surface waters; backfilling any sumps to cover and contain any drilling sediments and prevent inadvertent trapping of fauna; removal of any rubbish from the area and finally reshaping and replacing topsoil before planting of endemic shrubs and grasses. At completion of rehabilitation activities, photographs were taken at each of the sites as part of the post rehabilitation monitoring program.

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Technical Report Newmarket Gold Inc.
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The old landfill site was decommissioned and a new site commissioned in late 2011/early 2012. With commencement of the new site, the old landfill was covered and seeded. Monitoring of this area will be conducted in the future to ensure that re-vegetation is adequate and that no additional maintenance works are required. During the reporting period, some seepage from the Eastern WRD was observed. Samples of the seepage were collected and sent to an external laboratory for analysis. Analysis of these samples indicated high metal loads (further details and sample data are provided in the Union Reefs WMP). A management plan to address this seepage will be developed during the next twelve months.

Water is regularly monitored and the site has a Water Discharge License (WDL 138-02) for passive surface discharge during the wet season. This is in conjunction with the Unions Reefs Gold Plant water monitoring program.

20.5.2.1  Union Reefs Gold Plant

The Union Reefs gold plant was re-started in December 2009 by Crocodile Gold after being placed into care and maintenance when the previous operators went into receivership. Rehabilitation of historical disturbances is substantially complete; recent and on-going disturbances will be required to be rehabilitated upon completion of mining and processing.

All mineralization is transported to the Union Reefs gold plant from other Newmarket Gold sites and stockpiled on a ROM pad. From there, mineralization is fed through a three-stage crushing plant before being screened and passing through the milling circuit. The mineralization is reduced to approximately 75μm by two ball mills before entering directly into the leach circuit for leaching and adsorption or passing through gravity gold circuit. The gravity circuit includes four Knelson gravity concentrators that receive underflow slurry from the ball mills. Higher specific gravity gold particles are separated from non-gold bearing particles and removed directly to the gold processing facility. All remaining slurry is pumped to the Carbon-In-Leach circuit. This consists of two leach tanks followed by seven adsorption tanks where gold solution is absorbed onto carbon granules. The loaded carbon then passes through the elution circuit before being sent to the gold processing facility for the production of gold ingots.

Since 2002 tailings from the processing plant are treated and then deposited in the former Crosscourse open pit, which is estimated to have capacity for tailings at an assumed rate of 2.5Mtpa for the next 20-30 years. It currently receives approximately 300t of tailings per hour during normal processing conditions. The water and tailings level in the pit is regularly monitored.

Crosscourse Pit is bunded and little to no runoff is received. Due to the high volume to surface area ratio, evaporation from the pit, when compared to its volume is relatively low. An evaporation pipeline that sprays pit water into the air above the pit to increase evaporation was installed around the perimeter of Crosscourse Pit in November 2011, contributing to the reduction of the pit water inventory.

The water in Crosscourse Pit is extracted and recycled for use by the mill. A second pipeline and pump was installed in September 2010 to increase the volume being returned to the mill. This system is designed to increase the reuse of water from Crosscourse Pit, minimizing the requirements for processing additives and consequently reducing the potential for environmental impacts, extending the life of Crosscourse Pit as a tailings storage facility and reducing operating costs for Newmarket Gold.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Renewed Waste Discharge License (WDL number 138-03) was granted by NT EPA (Commencing December 19, 2014 and expiring on the December 1, 2016) for the UR Project Area. This license authorizes passive discharges from the historical Decant Pond at the south of the site into the McKinley River and Dam A to the north of the site into the McKinley River via Wellington Creek. A study was conducted but no action has been taken to date.

20.5.3      PINE CREEK AREA

The former Pine Creek mining operation has been rehabilitated and is in compliance with NT Government rehabilitation completion guidelines. All open pits at the Pine Creek area, with the exception of the Enterprise and South Gandy’s open pits, were backfilled and rehabilitated during operations.

A monitoring program is maintained, particularly over the tailings dams, northwest and southern waste rock dumps, run-of-mine (“ROM”) pad (stockpile 3) and the heap leach operations that were capped/repaired/ rehabilitated and seeded during 2001/02. Work is planned to repair some remediation undertaken by the previous operators on a small section of one WRD.

A water monitoring program is undertaken as per the requirements of the WDL 166-03, which is a license for passive discharge. A currently project is being undertaken to determine the effects of a passive discharge on the downstream environment. This is according to the WDL 166-03.

A water monitoring program is undertaken as per the requirements of Renewed Waste Discharge License (WDL number 166-03), granted by NT EPA on the December 9, 2014, and expiring December 1, 2016 for the Pine Creek Area. This license authorizes active discharges from the historical Process Water Dam (PCPWD) to Copperfield Creek and passive discharges from the Enterprise and South Gandy’s Pits to the Pine Creek system.

There have been a number of reports previously submitted to the Northern Territory Government regarding the rehabilitation activities undertaken at the Pine Creek area. The rehabilitation work undertaken has been widely recognized, having won an award for environmental excellence in 2002. In addition, the site was also selected by the Australian Center for Minesite Environmental Research (ACMER) national study of Landscape Functional Analysis as an index of rehabilitation success on mines in 1997. The Pine Creek area has been regularly utilized as a field trip venue to demonstrate excellence in environmental management for conferences and seminars.

As part of the previous WDL 166-01, SKM consultants were engaged to assess the 2010-2012 monitoring data (water/sediment/biological) for the development of Site Specific Trigger Values (SSTV’s). The review and update of SSTV’s occurs annually with the engagement of GHD consultants. This work is ongoing for the current WDL. Additionally assessment of Safe Dilution (SD) factors is currently ongoing with advice and engagement of GHD consultants.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

20.5.3.1  Environmental Management – Pine Creek

The Pine Creek deposits lies within areas, which have been subject to significant historical mining and mineralization processing activities for over 100 years. This historical activity, like many mining areas worldwide, has left permanent evidence of this activity on the physical landscape and the natural environmental balance may also have affected.

Location of the Property lies within an environment characterized by low relief, abundant ephemeral and permanent drainage and, particularly closer to the coast, sizeable billabongs and wetlands and a monsoonal wet season with heavy rainfall requires careful management of water, particularly discharge water from mining and milling operations.

Acid rock drainage is an issue at several locations and various systems have been developed to carefully manage this issue.

Newmarket Gold has included environmental management as an integral part of its operations. All exploration activities and mining operations have been performed in compliance with all environmental regulations within a defined environmental management program. Past operators reported that environmental assessments and project reviews have been completed as required and were thoroughly scrutinized before commencement of operations.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Site rehabilitation and reclamation has also been completed in a number of locations. This is currently an active part of the mining operations with waste dump rehabilitation a part of the daily mining activities. Site rehabilitation is factored into the operation costs for the earth moving contractor and is therefore required to be completed as soon as areas become available.

All recent mining operations have operated in accordance to Mine Management Plans submitted to DME, with various environmental permits in place, particularly including Waste Discharge Permits.

Since the Company took over the responsibility of the tenements in November 2009 several steps have been taken to ensure the environmental sustainability of the project. Several historical issues have been noted and the Company is in the process on ensuring these legacy issues are managed. An example of this is the work Newmarket Gold is doing on trails on waste water management in conjunction with external consultants on ways to affectively treat waste water.

There are currently no investigations of breaches of any regulatory regime nor are there any current sanctions or restrictions imposed by Government Departments. The Northern Territory Government has a constant review process including site visits. On these visits they inspect current and past mining areas to ensure the Company is compliant to the MMP’s approval as well as to the relevant Mining legislation. To date no major issues have been identified or recorded against the Company.

20.5.4      BURNSIDE AREA

Within the Burnside area, the deposits which are most likely to have significant future environmental concerns are Brocks Creek and Cosmo Mines, and the Fountain Head/Tally Ho, Rising Tide, Howley Pit and Mottrams deposits.

20.5.4.1  BROCKS CREEK

The Zapopan (Brocks Creek) Underground Mine is no longer operational and the Zapopan Pit is now flooded; initially with water from Faded Lily Pit mid-2011, then from the initial dewatering of Rising Tide Pit via Zapopan Creek in January 2012.

The Brocks Creek tailings dam, which was inherited from the previous owners, is currently in a stable form and the historical partial capping is currently rehabilitating naturally with grass cover establishing itself around the perimeter of the tailings dam. Approximately 1,213kt of oxide waste rock from the Rising Tide Stage 1 open cut operation was extracted and used to continue the cover of the Old Tailings Dam as part of closure criteria for the site. An estimated 9.8ha of the tailings dam was covered with 4m to 8m of oxide material.

The historical Faded Lily and Alligator WRD’s were rehabilitated during the mid to late 90’s with a native tree, shrub and grass species mix. The landforms are in a stable state. Alligator was primarily an oxide waste dump, and AMD is not an issue. Faded Lily was a sulphide WRD, and exhibits signs of AMD. At the base of the waste dumps exists sediment traps and wetland filters to minimize downstream impacts.

The Brocks Creek area is included in an approved Waste Discharge License (WDL 180-03) that is combined with the Cosmo Howley WDL (commencing December 23, 2015 and expiring September 30, 2017) , regulating wastewater discharge in the wet season. A regular sampling regime is employed to monitor creek flows and creek water quality.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

The former Brocks Creek CIL processing facility covered a relatively small area. This area has largely been rehabilitated, with only the plant foundations, office buildings and some scrap metal remaining. The office complex is now used by the PNX metals exploration team.

20.5.5      BURNSIDE AREA MMP'S

The Burnside area currently has a series of authorizations registered under care and maintenance/exploration and mining MMP’s approved. These MMP’s allows Newmarket Gold to do limited drilling for exploration and also has monitoring commitments attached. Newmarket Gold must ensure it meets these commitments or risk penalties from the DME. The approved mining MMP’s exist at North Point/Princess Louise, Howley and at Rising Tide deposits where previous mining has occurred.

20.5.6      ENVIRONMENTAL MANAGEMENT PLAN

Under the terms of the Mining Management Act, existing mining operations in the Northern Territory are required to submit an annual MMP to DME. This plan covers key aspects of mine operation, Occupational Health and Safety, environmental management and mine closure. This plan is then assessed and audited by DME. Upon approval of the MMP, an Authorization to Operate is issued to the mining operation.

The Company has submitted annual MMPs for all of its operating and exploration activities, and provided required annual reports to DME and other relevant departments.

Newmarket Gold has MMP’s in place with the DME for the Cosmo Project and are under Authorization numbers 0546-03. More details of the other plans currently in place for Newmarket Gold can be found in sections 4-11 and 4-12 above.

20.5.7      ENVIRONMENTAL BONDS

20.5.7.1  NT Opeations

Unconditional performance bonds totaling $12,221,052 for the NT Operations have been lodged with the Northern Territory Government to cover the anticipated cost of the rehabilitation commitments associated with the mining project. This is included in a total of $6,375,921 currently held by the Northern Territory Government for the Cosmo Mine. This bonding and mining authority allows for Newmarket Gold to conduct mining operations at the Cosmo Mine.

More details on bonds held can be found in Section 4-12 above.

Project/Site Authorization
No
Tenements Bonds
Maud Creek 0524-02 EL25054; EL28902; ML30260; ML30293 $107,984
Moline 0525-02 MLN1059; ML24173; EL28616; MLN41 $288,438
Fountain Head 0526-01 MLN4; MLN206; MLN1020; MLN1034; MCN1172; MCN4785 $984,816
Brocks Creek 0528-01 MLN1139 $1,264,915

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Project/Site Authorization
No
Tenements Bonds
North Point &
Princess Louise
0530-01 MLN823; MLN824; MLN825; MLN826; MLN827; MLN828; MLN829; MLN830; MLN831; MLN832; MLN858; MLN859; MLN860; MLN861; MLN862; MLN863; MLN940; MLN1112; MCN46; MCN47; MCN49; MCN50; MCN624; MCN625; MCN898; MCN899; MCN4432; MCN4434 $1,148,871
Pine Creek 0538-01 MLN13; MLN1130; MCN523; MCN1054; MCN1055; MA416 $538,738
Union Reefs 0539-03 ELR130; MA398; MA399; MA400; MA401; MA402; MCN506; MCN507; MCN734; MCN735; MCN738; MLN135; MLN779; MLN780; MLN822; MLN833; MLN856; MLN1109 $1,511,369
Cosmo Howley 0546-03 MCN377; MCN378; MCN379; MCN380; MCN852; MCN853; MCN854; MCN855; MCN856; MCN857; MCN1014; MCN1015; MCN1035; MCN1231; MCN1232; MCN3099; MCN3100; MCN3101; MCN3102; MCN3103; MCN3104; MCN3105; MCN3106; MCN3107; MCN3108; MCN3109; MCN3110; MCN3111; MCN3112; MCN3113; MCN3114; MCN3115; MCN3117; ML30892; ML30887; MLN809; MLN884; MLN885; MLN886; MLN887; MLN888; MLN889; MLN890; MLN891; MLN892; MLN993; MLN1000; MLN1027; MLN1053; MLN1060; MLN1062; MLN1129; ML27938 $6,375,921
Total     $12,221,052

TABLE 20-4 LIST OF BONDING HELD BY NT OPERATIONS

20.6 COMMUNITY CONSULTATION

Newmarket Gold has very little impact on the local community throughout the NT Operations mine sites. The main interaction at this site is with the local landowner on Douglas Station. Consultation meetings are held quarterly with site staff and property managers to discuss issues such as water management and land management activities.

20.7 MINE CLOSURE REQUIREMENTS

The NT Government Department of Mines and Energy retains an environmental bond against all exploration, mining and extractive ground disturbances undertaken in the NT. The bond is calculated using a tool supplied by the DME, which breaks the nature of the disturbance into discreet aspects of an operation. The breakdown of these items is outlined below.

Removal of site infrastructure

This includes the removal of buildings and offices, concrete pads, disconnection of services, removal of plant, removal of contaminated material and the removal of any processing/mining facilities.

Rehabilitation of Extractive pits/quarries

This includes the stabilization of any pit walls, placement of abandonment bunds, re vegetation and signage for any existing pits or quarries on site.

Rehabilitation of underground workings

This includes the sealing of any underground accesses, ventilation shafts, barricading and signage.

Rehabilitation of water/tailing storage facilities

This includes ensuring any dam is safe, removing sediment for dams, shaping and leveling material, drainage, re-vegetation and fertilizing.

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Technical Report Newmarket Gold Inc.
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Rehabilitation of stockpiles/waste dumps.

This covers any shaping, spreading, contouring, cleaning of material around an existing waste dump. It also includes the re-vegetation of the waste dump to allow for stabilization through vegetation. Finally water management from the waste dump would also need to be considered.

Exploration

This includes the rehabilitation of any surface exploration activities including drill pads, sumps, access tracks, core farms and sample farms. The majority of this work is completed after drilling campaigns are completed and this is currently up to date on site.

Access and Haul Roads

This covers the ripping and removal of any non-required roads around site. Also included would be the reapplication of topsoil and seeding/fertilizing to allow for re-vegetation.

Decommissioning and Post closure costs

This would cover the on-going assessment, monitoring and management of the site. Monitoring would cover water, weed and fire monitoring.

Contingency costs

This is established as 15% of the total of the costs for each of the areas outlined above. This amount is incorporated into the assessment tool to address variation in unit of measure costs, and changes over time due to project location or inflation

20.7.1      COSMO MINE

Mine Closure costs are estimated within the bonding arrangements with the government. To calculate the required level of security bonding a range of mine closure requirements would be required for each operation. In the case of the Cosmo Mine this has been done with the final estimated amount required for bonding in excess of $6.4 million.

On October 1, 2013, the Northern Territory Government adopted an initiative, as part of the amendment to the Mining Management Act, to address the legacy mine issues in the Northern Territory by the introduction of a 1% securities levy. There was a two-pronged approach to this arrangement. The first being that all current operations with securities would receive a 10% refund of the total bond held. The second being that an annual 1% security levy would be required on the current bond held as of July 1 of each year and made payable to the DME. The Cosmo Howley security bond is broken down as follows:

Domains Calculated Cost
Site Infrastructure $397,282
Hard Rock Pits & Quarries $119,124 
Underground Workings  $11,250
Tailings Storage Facilities and Dams $1,227,402    
Stockpiles and Waste Rock Dumps $2,176,056    
Exploration $21,150

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Domains Calculated Cost
Access and Haul Roads $127,170
Decommissioning & Post Closure Management $1,454,058   
Contingency @ 15% $830,023
Total Cost $6,363,515     

TABLE 20-5 MINE CLOSURE REQUIREMENTS FOR COSMO MINE

20.7.2      UNION REEFS

In the case of the Union Reefs area the final estimated amount required for bonding is in excess of $1.6 million. This is broken down as follows:

Domains Calculated Cost
Site Infrastructure   $64,000
Hard Rock Pits & Quarries $135,000
Underground Workings   $25,000
Tailings Storage Facilities and Dams $220,000
Stockpiles and Waste Rock Dumps $138,000
Exploration $178,000
Access and Haul Roads $137,000
Decommissioning & Post Closure Management $551,000
Contingency @ 15% $213,000
Total Cost $1,661,000     

TABLE 20-6 MINE CLOSURE REQUIREMENTS FOR UNION REEFS OPERATION

20.7.3      PINE CREEK

To calculate the required level of security bonding a range of mine closure requirements would be required for each operation. In the case of the Pine Creek area this would be required before a mining approval was granted, the final estimated amount required for bonding is approximately $3.0 million. This is broken down as follows:

Domains Calculated Cost
Site Infrastructure   $16,000
Hard Rock Pits & Quarries   $12,000
Underground Workings           $0
Tailings Storage Facilities and Dams   $32,000
Stockpiles and Waste Rock Dumps $2,160,000   
Exploration   $23,000
Access and Haul Roads     $7,000
Decommissioning & Post Closure Management $329,000
Contingency @ 15% $387,000
Total Cost $2,966,000     

TABLE 20-7 MINE CLOSURE REQUIREMENTS FOR PINE CREEK SITE

20.7.4      BURNSIDE

Mine Closure costs are estimated within the bonding arrangements with the government, however it is planned that this work is completed while operations are in place to reduce the exposure to significant

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

costs at the end of mining. To calculate the required level of security bonding a range of mine closure requirements would be required for each operation. In the case of the North Point deposit the final estimated amount required for bonding is approximately $1.2 million. This is broken down as follows:

Domains Calculated Cost
Site Infrastructure   $28,000
Hard Rock Pits & Quarries $122,000
Underground Workings            $0
Tailings Storage Facilities and Dams           $0
Stockpiles and Waste Rock Dumps $307,000
Exploration   $50,000
Access and Haul Roads   $36,000
Decommissioning & Post Closure Management $506,000
Contingency @ 15% $166,000
Total Cost $1,215,000     

TABLE 20-8 MINE CLOSURE REQUIREMENTS FOR NORTH POINT

20.8 COMMENTS ON ENVIRONMENTAL ISSUES AND LIABILITIES

The Author is not an expert in the assessment of potential environmental liabilities associated with mineral properties. Information contained herein, subject to Section 3.0 of this report, is sourced from earlier reports and Newmarket Gold site representatives.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

21 CAPITAL AND OPERATING COSTS

21.1 COSMO MINE

Costs detailed in this section have been derived from the 2016 Cosmo Mine budget and include current contractor rates for underground mining. Additional capital costs have been included for the removal of tailings from the Cosmo open pit immediately above the underground mine. This will enable the removal of the crown pillar, which forms part of the mineral reserve estimate.

21.1.1      CAPITAL COSTS

A summary of the capital costs for the Cosmo Underground Mine are presented as follows:

Capital Costs (Mining) $/t (ore)
Mobile Plant & Equipment $ 1.41
Processing Plant $ 1.69
Infrastructure $ 0.44
Underground Development $ 22.39
Tailings Removal $ 10.19
mineral resource Definition $ 2.06
Total $ 38.18

TABLE 21-1 COSMO MINE CAPITAL COST SUMMARY

21.1.2      OPERATING COSTS

The operational costs estimated for the mineral reserve analysis are summarized in Table 21-2. The mining fixed costs include the fixed labor and overhead costs for the Downer mining contract and the ownership costs for Downer’s mining equipment.

Operating Costs (Mining) $/t (ore)
Mining Fixed Costs $ 12.29
Power & Fuel $ 7.52
Stope Production Costs $ 19.99
Technical Services $ 1.75
Geology $ 6.32
Ore Haulage $ 6.41
Camp Costs- Cosmo $ 4.59
Processing $ 28.90
Administration $ 8.12
Cosmo Maintenance $ 1.75
Mining Fixed Costs $ 97.62

TABLE 21-2 COSMO MINE OPERATING COST SUMMARY

The mining costs are based on the 2016 budget and Downer’s contract rates for Cosmo Mine.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Processing costs are based on the 2016 budget operating costs for the Union Reefs processing plant.

21.2 UNION REEFS UNDERGROUND – PROSPECT

Costs have been derived from the 2012 Prospect Underground Mine Prefeasibility Study and relevant current operating costs for the Cosmo Mine and mill facilities. The cost modeling is based upon the use of contract personnel and rates to undertake the mining. Ongoing monthly allowances for technical staff and power consumption have also been included in the cost models.

21.2.1      CAPITAL COSTS

A summary of the capital costs for the Prospect Underground Mine is presented in Table 21-3 below.

Capital Costs $/t
Mobilization 0.69
Site Set Up 1.53
Pit Dewatering 0.18
Portal Preparation 0.61
Decline Establishment 0.41
Capital Development 35.37
Diamond Drilling 0.98
Vertical Development 4.61
Cablebolt 0.88
Personnel 4.98
Electrical Equipment and Installation 2.41
Total 52.64

TABLE 21-3 CAPITAL COST SUMMARY

21.2.2      OPERATING COSTS

The operational costs estimated for the mineral reserve analysis are summarized in Table 21-4 below.

Operating Costs $/t
Mining Fixed Costs 17.7
Power and Fuel 7.6
Stope Production Costs 32.7
Technical Services 2.4
Geology 4.9
Ore Haulage 6.3
Camp Costs - Cosmo Village 4.8
Processing 28.9
Administration 9.1
Prospect Maintenance 1.6
Total 116

TABLE 21-4 OPERATING COST SUMMARY

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

The mining costs have been estimated from both the current Cosmo Mine actual costs and contract rates for Cosmo.

21.3 UNION REEFS OPEN PIT – ESMERLADA

Costs for mining activities at the Emeralda Mine have been derived using the costs outlined in the Pine Creek 2013 technical report, particularly the operating costs. This is due to the Authors opinion that these costs are still seen as suitable for mine planning as there has not be a significant shift in mining costs since that time. Additionally some costs have been determined using the actual costs incurred during operations by the company at sites such as North Point and the Howley open pits.

Operating costs for the Esmeralda would be similar to the Pine Creek operations as the planning and designs have been constructed using the same methodology and requirements.

21.3.1      CAPITAL COSTS

The capital costs outlined in this section of the report are taken from previously incurred costs from open pit mining completed by the Company at places like North Point and Howley.

The mobilization costs, in this case, have been assumed to be zero as the fleet proposed would be mobilized for the mining operations at Pine Creek prior to commencing work at Esmeralda. There could possibly be other cost savings by having this operation commence with Pine Creek operations but these savings have not be justified at this point in time. It would be recommended to understand these potential savings prior to commencing open pit mining in the NT Operations.

Item $’000 $/t
Mobilization 0 0
Pre-mining 70 0.29
Top Soil 50 0.21
Progressive Rehab 50 0.21
Temporary infrastructure 180 0.75
Total 350 1.46

TABLE 21-5 CAPITAL COSTS FOR ESMERALDA MINERAL RESERVES

Operation Costs

The operational costs estimated for the mineral reserve analysis are summarized in Table 21-6 below. The physical mining cost (excluding haulage and processing costs) is approximately $4.83/t for both mineralization and waste rock.

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Technical Report Newmarket Gold Inc.
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Item $/t
Drilling 0.32
Blasting 0.20
Loading 0.51
Haulage 1.04
Auxiliary 0.59
Grade Control 0.07
Labor 2.00
Fixed Costs 0.10
Haulage costs 4.00
Processing Cost 26.00
Total 34.83

TABLE 21-6 OPERATING COSTS FOR ESMERALDA OPERATIONS

21.4 PINE CREEK OPEN PITS

Operating Costs have been derived from the Pine Creek 2013 technical report. A detailed summary of the costs and requirements is outlined in Basile 2013 (Basile and Edwards 2013). The assumptions made in this section have not materially changed in the opinion of the author, however prior to operations commencing a more detailed review of both operating and capital costs should be determined.

The Capital Costs have been established using previously incurred costs from other open pit operations managed by the Company.

21.4.1      CAPITAL COSTS

The Mining Capital Costs were estimated to be $2.0 million over the proposed life of the open pit mining period. The capital costs are estimated to have an accuracy of ±25%. Greater variations in the estimated capital costs may occur if there are changes to the proposed mine plan.

The mining capital costs included mining equipment and site engineering. In addition, it also includes site closure and monitoring at the end of the open pit mining life.

A table outlining capital expenditure requirement is below in Table 21-7. All pre-stripping of waste material has been excluded from the capital expenditure and included in the operating costs as to fully reflect the amount of pre-strip needed to expose the mineralization in all pits.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

Item $’000 $/t
Mobilization 400 0.32
Pre-mining 350 0.28
Top Soil 150 0.12
Progressive Rehab 300 0.24
Temporary infrastructure 800 0.64
Total 2,000 1.61

TABLE 21-7 CAPITAL COSTS FOR PINE CREEK OPERATIONS

21.4.2      OPERATING COSTS

The operational costs estimated for the mineral reserve analysis are summarized in Table 21-8 below. The physical mining cost (excluding haulage and processing costs) is approximately $4.83/t for both mineralization and waste rock.

Item $/t
Drilling 0.32
Blasting 0.20
Loading 0.51
Haulage 1.04
Auxiliary 0.59
Grade Control 0.07
Labor 2.00
Fixed Costs 0.10
Haulage costs 4.34
Processing Cost 28.90
Total 38.07

TABLE 21-8 OPERATING COSTS FOR PINE CREEK OPERATIONS

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

22 ECONOMIC ANALYSIS

22.1 GENERAL COMMENTS

A combined mining and processing schedule for all NT Operations, inclusive of all Pine Creek open pits, Cosmo Mine underground, and Union Reefs (Prospect Underground and Esmeralda open pits) was developed so that an overall economic cash flow model could be developed. This section details the creation and underlying assumptions for this model and the predicted financial value of the combined mineral reserve.

22.2 SCHEDULING STRATEGY AND ASSUMED OPERATING PARAMETERS

The schedule was produced at Quarterly (i.e. 3 month) intervals allowing for adequate resolution of specific mill feeds and practical mining campaigns within dry season intervals.

The mining sequence proposed for NT Operations is as follows:

The schedule assumes no significant delays in obtaining the appropriate NT Government approval to mine at each separate project. The milling capacity at Union Reefs remains constant at 1.2 Mtpa feed for the life of the schedule.

Capital cost spends were assigned to each of the specific operations. An estimate of taxes to be paid was made based on net operating cash flows for each period taking into account depreciation of assets which was estimated on a straight line basis for each reserve asset.

The mining costs applied to each operation are presented in Table 22.1.

The royalties, processing recoveries and milling costs used in the financial model are presented in Table 22.2.

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Technical Report Newmarket Gold Inc.
December 2015 Northern Territory Operations

The total process costs for each deposit vary primarily due to two major factors, the type of material (i.e. Oxidized, Transitional or Fresh), which primarily affects direct comminution costs and the distance of each deposit from the Union Reefs mill, which affects road train haulage costs per tonne of feed material.

Mine Mineralized material
Mining Cost
(A$/ t)
Waste material
Mining Cost
(A$/t)
Underground    
     Cosmo Deeps (Development and Stoping) 97.6  
     Prospect (Development and Stoping) 116.0
Open Pit  
     International 4.83 4.83
     South Enterprise 4.83 4.83
     Kohinoor 4.83 4.83
     Cox 4.83 4.83
     Esmeralda A 4.83 4.83
     Esmeralda B 4.83 4.83

TABLE 22-1: MINING OPERATING COSTS

Mine Royalties Processing Recovery
%
Milling Cost
A$/ore t
Underground
     Cosmo 0.00 92.0% 28.90
     Prospect 0.00 93.0% 28.90
Open Pit      
     International $4/oz 85.0% 33.24
     South Enterprise $4/oz 90.0% 33.24
     Kohinoor $4/oz 90.0% 33.24
     Cox $4/oz 90.0% 33.24
     Esmeralda A 2% NSR 90.0% 30.00
     Esmeralda B 2% NSR 90.0% 30.00

TABLE 22-2 ROYALTIES AND PROCESSING PARAMETERS

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22.3 NET PRESENT VALUE AND INTERNAL RATE OF RETURN

Newmarket Gold modeled the cash flows of the combined operations using a base gold price of A$1450/troy oz. and a 10% p.a. discount rate.

Base case reserve economic model results are presented in Table 22.3.

Project KPIs Pre - Tax Post - Tax
Undiscounted Cash (A$ M) 3.6 1.8
Net Present Value (NPV) (A$ M) 2.0 0.7
Internal Rate of Return % 11.2% 4.7%
Payback (Quarters) Undiscounted Cash 9 10
Payback (Quarters) Discounted Cash 9 10
Max. Undiscounted Cash Draw (A$ M) 3.75 5.1
Max. Discounted Cash Draw 3.5 4.9
Operating Margin % (EBIT / Revenue) 14.9%  

TABLE 22-3 ECONOMIC ANALYSIS RESULTS

22.4 CASH FLOWS

Figure 22 4 presents the cumulative discounted and undiscounted pre-tax cash flow, and NPV of the project discounted at 10%. The graph shows that the Project has negative cash flows in many periods, which reflect periods of lower gold production, higher operating costs (e.g. Initial high strip ratios for open pits) and capital cost outlays (e.g. development of the Prospect underground decline). However, the cumulative pre-tax cash flow becomes positive in Quarter 9 and total cash flow is positive for the life of the combined project.

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The post-tax NPV for the project is calculated to be A$0.7 M with a payback period of 2.5 years. This is illustrated in Figure 22-5 below.

Table 22.4 presents the combined operational cash flows for the Project.

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22.5 SENSITIVITIES

The project’s greatest risk and sensitivities are in relation to the revenue generated from the project.

Economic model sensitivity analysis was completed on the gold price, as well as capital and operating cost estimates, with the results presented in Table 22.5. The results indicate that the Project is most sensitive to variations in gold price, mining operating costs, mill operating costs and capital costs in that order. Figure 22 6 presents a graphical representation of these findings.

Pre - Tax Sensitivities % Move (NPV $AM)
  -20% $8.8
Capital Cost -10% $5.6
  10% -$0.8
  20% -$4.0
  -20% $16.4
     
Mill Operating Cost -10% $9.4
  10% -$4.6
  20% -$11.6
  -20% $29.1
  -10% $15.8
Mining Operating Cost 10% -$10.9
  20% -$24.3
  -20% -$45.4
Gold Price -10% -$21.5
  10% $26.3
  20% $50.3

TABLE 22-5 PRE-TAX SENSITIVITIES

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23 ADJACENT PROPERTIES

The material in this section has been reviewed by Author Mark Edwards in terms of content. As the properties are not owned and operated by Newmarket Gold, no attempt has been made by the Company or the Authors to assess or verify the technical information related to the following section and such information is not necessarily indicative of the mineralization on the Property. Where possible, publicly stated data has been used to compile this section.

23.1 NON - NEWMARKET GOLD DEPOSITS

There are several other mines in the Northern Territory and numerous historical gold, base metal and uranium occurrences within the Pine Creek Origin and at any time some of these prospects may be subject to various stages of evaluation or development.

The following is a description of some of the more relevant properties, which are proximal to the Cosmo Mine; however, it is important to note that this list is not complete. Information contained herein is derived from public sources, including company websites and has not been verified by the Author. All references to mineral resources and/or mineral reserves are reported to have been prepared in accordance to JORC Code; however, the Authors have not verified this information.

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23.1.1      SPRING HILL GOLD PROJECT

Spring Hill was discovered in 1880 and produced about 20,000oz of high grade gold over the next 25 years (Ahmad, Wygralak and Ferenczi 2009).

The Spring Hill Gold deposit was held by Western Desert Resources Ltd but was under option to Thor Mining PLC.

In 2014 Thor Mining held equity rights to 80% of the Spring Hill mineral resource, including its current 51% project ownership through the issue to Western Desert Resources of 10 million plus shares to the value of A$500,000. On September 5, 2014, Western Desert Resources was placed into voluntary administration

In early 2016 Thor Mining announced that it had sold the property to PC Gold, a private Australian company. Thor had been advised that PC Gold had completed their technical due diligence, and had moved towards finalization of formal agreements.

PC Gold will pay a royalty of A$6.00 per ounce of gold produced from the Spring Hill tenements where the gold is sold for up to A$1,500 per ounce; and A$14 per ounce of gold produced from the Spring Hill tenements where the gold so produced is sold for amounts over A$1,500 per ounce.

The tenement package comprises a 1,100 hectare Mining Lease Application is surrounded by a 36km2 Exploration License.

The deposit lies in the historical Pine Creek Goldfield to the north of Newmarket Gold’s Union Reefs processing facility. The deposit is hosted within greywacke and siltstone units of the Mount Bonnie Formation, which occur with distinctive inter-beds of hematitic siltstones associated with layers of banded granular quartz and ironstone.

Gold at Spring Hill occurs mainly in quartz veins concentrated in fracture zones and the axial zones of anticlinal fold structures. Much of the gold is relatively coarse-grained, in the visible range, imparting significant ‘nugget effect’ to drill samples.

Four main zones of gold mineralization cover an area of approximately 1,000mx400m. They have been outlined during the early 1990s and mid-2000s by drilling conducted by previous owners of the project around historic workings. These zones have not been drill tested below 150m. Additionally, several subordinate occurrences have been identified in adjoining areas.

The Spring Hill deposit has a Historical Resource of 3.6Mt at 2.34g/t gold for approximately 274,000oz of contained gold (McKibben, Standing and Gerritsen 2008).

Thor Mining started a 10-hole drilling program in September 2011 designed to confirm and extend the known mineralization.

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A high-resolution helicopter based magnetic (helimag) survey was conducted over the Thor tenements in July 2012.

23.1.2      MT PORTER GOLD PROJECT

The Mt Porter Gold deposit, located 15km north of the Union Reefs gold processing plant, is a sediment-hosted stratiform gold deposit with quartz-sulphide-vein-hosted stratabound gold mineralization associated with cherty iron formation and carbonaceous mudstone mainly in the Middle Koolpin Formation in the north-northwesterly trending Mt Porter Anticline.

Stream sediment sampling and RC drilling of gold anomalies located by the earlier work of Homestake Gold of Australia in 1996-1999 by Arafura discovered the deposit. Arafura completed two drilling campaigns at Mt Porter. The 2006 program encountered a 13 meter intersection grading 2.28 g/t gold to the west of and beneath the identified resource. This represents a new and unexpected zone of mineralization.

The following information has been extracted from www.arafuraresources.com.au

“The Mt Porter Gold Project represents an advanced exploration opportunity within the highly prospective Pine Creek region of the Northern Territory, Australia. Features about the deposit include:

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Most of the mineralization intersected to date at Mt Porter occurs in a complex many-hinged fold zone, on and immediately to the west of the main axis of the Mt Porter Anticline. This zone is bounded by at least three major faults – a NE trending structure to the southeast, an ESE trending structure to the north at about 10500N (local grid) and a major NS trending fault and shear zone to the west on about 10100E.

In the Frances Creek area, some 5km to the northeast of Mt Porter, Arafura completed forty RC holes by the end of 2004 on a number of vein deposits including the Golden Slips and Golden Honcho deposits. This drilling encountered high gold grades within quartz veins hosted by sandstone. Grades as high as 47.5g/t Au over intervals of up to 4m were recorded. The Golden Honcho deposit is open along strike to both north and south and at depth.

In 2006, Arafura completed an RC drilling program (four holes for 321m) into the margins of the 10400 Zone where the westernmost hole encountered a previously unknown zone of gold mineralization (“248 Gold Zone”) over a 13 meter interval and grading 3.53g/t Au. This intersection is deeper and to the west of the known mineral resource.

In March 2013, Arafura executed an agreement over the Mt Porter-Frances Creek Project with Ark Mines Limited. In 2015t was announced that Ark was in compliance with its obligations and had met all expenditure requirements. In August 2015 Ark announced it had executed a processing agreement with Newmarket Gold to process the Mount Porter ore at the union Reefs mill with Ark receiving 55% of net cash from such processing after cost, expenses and royalties. The project has secured environmental approval and a Native Title agreement. Ark has announced it expects to start mining in July 2016.

23.1.3      FRANCIS CREEK IRON MINE

The Frances Creek Iron Mine, now privately owned by the Noble Group, is located 20km north of Pine Creek on the Stuart Highway, 190km south of Darwin and about 15km from the recently completed Adelaide-Darwin rail line. Singapore-listed Noble owns Territory Resources after completing a $A133 million takeover of the Australian company in 2011.

The Frances Creek Mine produced high-grade hematite mineralization for export. Mining commenced in May 2007 and made the first shipment from the port of Darwin in September 2007. The company exported its 500,000th tonne of high-grade iron mineralization in May 2008. The mine continued to operate until January 2015 when it was placed into care and maintenance. It was the last Iron Ore mine operating in the Northern Territory after the collapse of Western Desert Resources and Sherwin Iron in late 2014.

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Geologically, the area is underlain by lower Proterozoic folded sediments intruded by lower Proterozoic granitoids. The iron mineralization occurs in a strike fault breccia in the lower Wildman Siltstone and ranges in composition from hematite to goethite and limonite.

The main iron mineralization at Frances Creek is known over a distance of approximately 35km. Of the fifty named deposits, 13 located on the southern end of the mineralized belt were exploited from 1967 to 1974 for a total production of around 6Mt. As of February 2007, total mineral resources were 9.73Mt at 60.7% Fe with 4.81Mt in mineral reserves.

Territory Resources report that there are excellent prospects for iron mineralization within the project area. Massive hematite continues at depth below the floors in the abandoned pits and occurs in a number of non-mined deposits drill-evaluated at Frances Creek in preparation for mining.

Re-optimisation of the mine in August 2010 resulted in an iron ore mineral reserve increase of 50 percent thereby increasing the life of mine to beyond 2014.

   

New beneficiation plant installed in 2011.

   

Acquired by multi-national China based Noble Group (parent company to Territory Iron Limited) in 2011.

The Frances Creek Mine takes in an area covering 619km2 with the main iron mineralization covering a distance of approximately 30km. Between 1967 and 1974 there were 50 known deposits with 13 being mined at that time, mainly at the southern end of the mineralization zone from which 6Mt were produced. Mineral resources and mineral reserves at the end of 2010 had indicated and inferred mineral resources standing at 9.9Mt producing 58.1 percent iron and probable mineral reserves of 5.8Mt at 57.9 percent iron. The mine is capable of handling 2Mt of iron ore a year. The Frances Creek deposit consists of haematite with some limonite and goethite all of which is exported to Asia, predominately China.

In January 2011 the then owner of the Frances Creek Mine, Territory Resources Limited, commissioned the new $4.3 million iron ore beneficiation plant. It was an immediate success when tested with an 800,000t stockpile with a grading of 51% iron. This was beneficiated to 60% iron to be sold as fines. The new plant had a feed rate of 130t per hour with a designed 55 percent mass recovery. The plant was soon treating 80,000t of scalp feed producing 44,000t of high grade fines a month.

The following mineral resource and Ore Reserve in the table below are a compilation of JORC Code compliant Resource and Reserve reports, depleted for production up to the end of 31 March 2011 and are a further update to that last published in the 2010 Territory annual report.

As reported by Territory, Frances Creek project contained 5.2 Mt of Probably Ore Reserves within 9.3 Mt of mineral resources (including 1.27 Mt of Manganiferous Goethite mineral resources)

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TABLE 23-1 FRANCIS CREEK M INERAL RESERVES AND MINERAL RESOURCES, 2011

In 2015 with the significant drop in Iron Ore prices the Fracnis Creek mine was placed on care and maintenance with all mining activities on hold.

23.1.4      PNX METALS

PNX Metals (formerly named Phoenix Copper) acquired the Iron Blow and Mt Bonnie ML’s from Crocodile Gold in 2014. Both properties hosted massive sulphide deposits that had been explored at various times in the past. The Iron Blow deposit had previous mineral resource estimates completed, while the Mount Bonnie deposit had a significant amount of drilling completed on it.

PNX issued a mineral resource statement based on a new interpretation and inclusion of drill holes completed by Crocodile Gold in 2011.

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Depth AuEq cut-off (g/t) Tonnes AuEq
(g/t)
Au (g/t) Ag (g/t) Cu
(%)
Pb (%) Zn (%) ZnEq %
>-90
mRL
0.7 2.2Mt 6.7 2.4 140 0.3 1 4.9 11.8
<-90
mRL
3 0.4Mt 5.6 2.7 71 0.4 0.4 4.1 10
Total Inferred mineral resource 2.6Mt 6.5 2.4 130 0.3 0.9 4.8 11.5
Total Contained Metal 543,000
oz
203,000
oz
10,700,000
oz
7,000
t
23,000
t
125,000
t
300,000
t

TABLE 23-2 IRON BLOW DEPOSIT MINERAL RESOURCE ESTIMATE

They subsequently drilled 2 diamond drill holes and completed some metallurgical test work. Drill hole IBDH023 intersected 50.39m @ 10.12% Zn, 2.66g/t Au, 283g/t Ag, 0.57% Cu, 1.39% Pb from 155.72m, In addition a Western Lode was intersected between 229.0m and 248.0m down-hole that returned 19.0m @ 3.45% Zn, 1.33g/t Au, 18.2g/t Ag, 0.38% Cu, 0.1% Pb from 229m. Immediately below this Western Lode intercept is a gold bearing breccia, which contained 26.5m @ 1.5g/t Au from 248m.

Drill Hole IBDH024 intersected disseminated low grade sulphide mineralization consistent with the margin of the deposit, including 12.85m @ 1.22% Zn, 0.56g/t Au, 27g/t Ag from 265.15m

At Mount Bonnie PNX drilled 12RC holes for 1,114m. Highlights included:

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PNX Metals is also looking at a geophysical target 2.5km southeast of the Iron Blow deposit that displays a similar signature to that of the VMS deposits in the region.

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24 OTHER RELEVANT DATA AND INFORMATION

In previous Technical Reports on the NT Operations, the Maud Creek Deposit was included in the mineral resource and Reserve Statements. In 2015 Newmarket Gold commissioned SRK (Australia) Pty. Ltd. to complete a staged Feasibility Study on the deposit The first stage of this study was to complete a Preliminary Economic Assessment (PEA). This PEA study was designed for the Maud Creek deposit to be a standalone operation relying on its own infrastructureusing a combination of open pit and underground mining. As part of this study an optional scenario to utilize Newmarket Gold’s infrastructure at Union Reefs will be included to assess the economic benefits to the Maud Creek operation. This information is currently not available but will be considered when the PEA technical report is completed. This report is due to be completed in the near future and will be lodged on SEDAR.

In 2015 Newmarket Gold entered into a toll treating agreement with Ark Mines Ltd (Ark) with respect to their Mt Porter deposit located to the northeast of the Company’s Union Reefs processing facility. Ark continue to advance their mine approvals at the time of writing this report. If they are successful in gaining approval to mine the deposit then some additional mill feed will be supplied to the Union Reefs processing facility. This toll treating has not been factored into the capital or operating costs for the NT Operations. It may have an effect on some cost per tonne estimates within this report, as the base tonnes processed through the mill will be increased. The Authors do not believe these changes will be material to the mineral reserves reported in this report.

No other relevant information is required to make the technical report understandable and not misleading.

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25 INTERPRETATION AND CONCLUSIONS

The NT Operations includes 29 deposit mineral resources (excluding the Maud Creek deposit) and 7 deposit mineral reserves over 3 separate areas (Cosmo, Union Reefs and Pine Creek). These mineral resources have been generated over several years using several different modeling techniques.

Operations continue at the Cosmo Mine with the processing facility at Union Reefs continuing to perform as expected. The cost structures are well understood through local experience in both open pit and underground mining since operations commenced in 2009.

       The Authors have made the following interpretations and conclusions:

There is some uncertainty regarding the permiting process for the Pine Creek deposits, however, the Newmarket team understand the obligations to obtaining these approvals and will work within the requirements of the mining department to gain approvals. There is some risk that this could delay the commencement of the operations but it is not of the opinion of the author that this approval will not be granted.

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There are no other significant risks or uncertainties (outside a significant reduction in gold price assumptions) that could be reasonably expected to affect the reliability or confidence in the contained information relating to mineral resources or mineral reserves as outlined in this technical report. The risk of the gold price assumptions has been mitigated by the use of an industry standard approach to estimating the price to be used in all estimations.

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26 RECOMMENDATIONS

26.1 COSMO MINE

Advancement in the geological and mineralization understanding for the Cosmo Mine over the past 12 months has resulted in the definition of exploration targets within the mine. It is recommended that this understanding continues to be developed and advanced through exploration drilling campaigns. Table 26-1 covers the proposed exploration programs to continue the development of the Cosmo Mine.

Target Current
Exploration
Status
Potential
outcome
Description Diamond
Drill Meters
Total
Cost
Western Lodes Advanced Scoping Inferred Status Plunging mineralization system close to the 640 exploration drill drive 3,000 $350,000
Lantern 700 lode Project Scoping Inferred Status Detailed drilling of 700 lode material within Lantern lode close to current development 3,500 $350,000
Lantern Central Exploration Investigative Longer drilling testing the Central zone of the lantern target area 3,000 $400,000
Hinge Footwall Project Scoping Inferred Status Drill testing the hinge zone below the F1 fault, currently intersected with Sliver drilling 2,000 $250,000
Cosmo Deeps Project Scoping Investigative Investigative drilling of the 100-300 lodes down plunge of current mineral resources 3,500 $420,000
Sliver Project Scoping Inferred Status Continue the development of the Sliver target down plunge of current mineral resources 5,800 $700,000
Cosmo Surface
2300mN
Exploration Investigative Test the down plunge extensions of the Sliver and Cosmo Deeps target 3,600 $900,000
Total Exploration       24,400 $3,370,000

TABLE 26-1 EXPLORATION PLANS FOR COSMO

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Figure 26-1 shows the approximate location of the Western Lode, Sliver Lode, Lantern Lode and the Footwall Hinge Lode. These areas have the potential to add additional ounces to the mineral resource and mineral reserve estimates over the next 12 months. The 640 Drill Drive extension is currently underway and will form a good platform for underground drilling of most of these targets.

Infill drilling is also planned to continue in 2016 with the aim being to maintain at least 12-18 months of mineral reserves drilled out to 20m x 10m spacing. The possible status for the end of 2016 (Table 26-1) reflects the opinion of the Author of this technical report. This drilling is seen as a critical path to replacing mineral reserves mined each year. This is the highest priority drilling for the Cosmo Mine and is the focus for the geological team based at the mine. Each program is reviewed regularly and altered to provide the required outcomes for mine planning purposes.

26.2 UNION REEFS

Drilling completed at the Esmeralda deposit has demonstrated the potential for future mining activities. While the overall reduction in mineral resources (combining Indicated and Inferred material) has resulted, it has been recognized that one diamond hole performed badly with lower than expected core recoveries. This has resulted in the reduction of tonnes in the core of the Esmeralda A deposit. It is recommended that a second RC hole be twinned with this diamond hole that returned questionable assay results to confirm the width of the mineralization. It would be estimated that an additional 100m of RC drilling be completed at a cost of $10,000 (excluding any potential mobilization costs).

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Work also needs to continue on developing the Mine Management Plan for the Esmeralda deposit. This work advanced significantly over the past 12 months with base line studies and a Notice of Intent supplied to the Northern Territory Environmental Protection Agency (NTEPA). This work will continue at an estimated cost of $80,000 to finalize the approval to mine.

With the potential to commence mining operations at the Esmeralda deposit located to the south of the Union Reefs processing facility, it is recommended that all mineral resources around the Union Reefs processing facility be reviewed for potential mining. Some drilling was completed in 2011 around deposits, such as Millars and Lady Alice. These drilling results should be used with the new understanding gained at Esmeralda and Prospect deposits and new mineral resource estimations should be completed. These could then be optimized to identify the potential for open pit mining. The costs of this work would be captured within the current NT Operations staff budget.

26.3 PINE CREEK

In the Pine Creek area there are currently four different mineral reserve deposits reported, including International, Kohinoor, Cox and South Enterprise deposits. There is the potential to add one year’s additional processing material for the Union Reefs facility. Some work is required to further understand the permitting process for these operations. While the deposits are located on an active Mineral Lease, work is required on the development of a Mine Management Plan for operations. This will require $150,000 of test-work and reporting to be completed.

There is also the potential to identify additional mineral resources at Pine Creek, particularly around the Enterprise South and Gandy’s North deposits. It would be estimated that approximately 2,000m of RC drilling for Enterprise South would be required at a cost of $200,000. At Gandy’s North a diamond drilling program of 1,500m is recommended at a cost of $400,000 in order to test the higher grade plunging structure. This would potentially be an underground target, but due its proximity to the surface there would also be some open pit mineral resource potential.

Target Current
Exploration
Status
Potential End
of 2016 Status
Description Diamond
Drill Meters
Total Cost
Gandy's North Project Scoping Inferred Status Drilling the down plunge mineralization to inferred status 1,500 $400,000
Enterprise South Project Scoping Inferred Status Drilling the extensions of the Enterprise South mineral resource 2,000 $200,000
Total
Exploration
         3,500 $600,000

TABLE 26-2 EXPLORATION PLANS FOR PINE CREEK

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26.4 BURNSIDE AREA

Newmarket Gold has been active in the past 2 years in rationalizing land holdings and mineral resources within the Burnside area. During this period the Iron Blow deposit has been divested to PNX Metals Ltd, the Bridge Creek deposit has been divested to a local quarry operation and the Glencoe deposit has been divested to Ark Mines Ltd. Also during this period a series of smaller, non-core Mineral Leases have been sold to third parties. It is recommended that this divestment of non-core assets continue to rationalize holdings within the Company’s NT Operations.

The Western Arm, Kazi and Bon’s Rush deposits are located proximal to each other. These deposits contain mineral resources in the Inferred category. None of these three deposits have been previously mined It is interpreted that they contain significant amounts of oxide mineralization. The mineral resource estimates with respect to these deposits were completed in the early 2000’s and will require an update. Investigations are underway to understand the amount and quality of diamond drilling that was completed and the remaining drill core that is available for additional study and test work. This drill core could be analyzed for required QAQC purposes. It would be estimated that this work would cost in the order of $10,000.

Drilling would also be required to convert these Inferred mineral resources to a high category to be used in future mineral reserve estimations. Below is an estimation of the drilling requirements for each deposit to convert them into a suitable mine plan.

Target Current
Exploration
Status
Potential End of
2016 Status
Description Diamond
Drill Meters
Total Cost
Western Arm Project Scoping Indicated Status In-fill drilling to expand the current Inferred mineral resource 6,500 $900,000
Bon's Rush Project Scoping Indicated Status In-fill drilling to expand the current Inferred mineral resource 6,500 $1,075,000
Kazi Project Scoping Indicated Status In-fill drilling to expand the current Inferred mineral resource 3,500 $575,000
Total
Exploration
         16,500 $2,550,000

TABLE 26-3 EXPLORATION PLANS FOR BURNISDE AREA

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27 REFERENCES

Ahmad, M., Wygralak, A. S., Ferenczi, P. A. (2001) GOLD DEPOSITS of the Northern Territory. Report 11, Northern Territory Geological Survey, Department of Mines And Energy, Northern Territory.

Ahmad, M., Wygralak, A. S., Ferenczi, P. A. (2009) GOLD DEPOSITS of the Northern Territory. Report 11, Northern Territory Geological Survey, Department of Mines and Energy, Northern Territory.

Second edition updated by Wygralak, A.S. and Scrimgeour, I. R. (2011)

Alexander, K. R., M. E. Kavanagh, and G. L. Rolfe. 1990. "Cosmo Howley Gold Deposit." Geology of the mineral deposits of Australia and Papua New Guinea (Australasian Institute of Mining and Metallurgy) Monograph 14: 751-753.

AMC. 2014. "Geotechnical Study and report on the Cosmo Deeps Mine". Internal Report for Crocodile Gold.

Ashley, P.M. (2013) “Petrographic report on twelve drill core samples from the International deposit, Pine Creek area area, Northern Territory” For Crocodile Gold Australia Operations Limited – unpublished

Australian Bureau of Statistics. 2014. Australian Demographic Statistics . June. http://www.abs.gov.au/ausstats/abs@.nsf/mf/3101.0.

Australian Government Bureau of Meteorology. n.d. Northern Territory Weather and Warnings Summary. http://www.bom.gov.au/nt/.

Bajwah, Z. (2007a) Annual Exploration Report on MLN’s 4, 206, 1020, 1034 MCN’s 1172, 4785 For period ending 28th February 2007, Fountain Head Burnside Project NT. Annual Technical Report for the DME

Bajwah, Z. (2007b) Annual Exploration Report “Yam Creek/North Point Group”, For the period ending 31 December 2006. Annual Technical Report for the DME

Bajwah, Z. (2007c) Annual Exploration Report, Mount Bonnie Group MLN’s 342, 346, 405, 459, 811, 1033, 1039 MCN’s 504, 505, 3161. For the period ending 31 December 2006. Annual Technical Report for the DME

Bajwah, Z. (2008). Annual Exploration Report MCN’s 3705-37017; MLN1103 For Period Ending 30 April 2008 Woolwonga Group Burnside Project NT. Annual Technical Report for the DME

Balfour, I. S. 1981. Administration of the Northern Territory, Chief Warden (Mining) of the Northern Territory, Extracts from annual reports 1905-1925.

Blanchard, R., and G. Hall. 1937. Interim Report on Howley Propoerties North Australia, May 1937. May.

Biddlecombe, R.M., (1985). The Elizabeth Gold Mine, Pine Creek, N.T., Australia

Butler, Ian (1991) “Report on Exploration Carried Out on Mineral Leases N51-N57 Caroline Leases, N.T.” Eupene Exploration Enterprises Pty Ltd for Nicron Resources Limited

Card, D. (2012). Burnside VTEM 2011 Interpretation Memo. Memo from Southern Geoscience Consultants regarding VTEM prioritization work

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Chapman, W. (2013) “Prospect Underground Mine 2012 Prefeasibility Study, Union Reefs, Northern Territory” Study completed for Crocodile Gold

Chow, K. C. (2002) mineral resource Statement Union Reefs Gold Mine at 31 December 2001. AngloGold Australia Limited

Dann, R.N., Delaney, W.R., (1984), Geology of the Enterprise Mine, Pine Creek, Northern Territory, The AusIMM Annual Conference, Darwin, N.T. 1984, pages 407-413

Dunn, M., Hall, D., Payne, N. and Tangney, G., (1998); Report on the Zapopan prospect, Internal report to Acacia Resources Ltd, September 1998.

Eupene, G. S. (1989) Review of Pine Creek Deposits, Prepared by Eupene Exploration Enterprises Pty Ltd for James Askew Associates.

Farrelly, C.; (1996): Bridge Creek, MLN766; 1060, ore resource block model, June 1996. Unpublished Northern Gold NL report.

Fawcett, C. (1993) “Report on Work Undertaken on the Gandy’s Hill and International Leases from 1980 to 1993, Prior to the commencement of Mining” Open File report from Northern Territory Geological Survey (NTGS)

Floresca, A. (2012) Report on the Resources and Reserves of the Pine Creek North (MLN1130). Internal unpublished report for Crocodile Gold

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Groves, D.I., Goldfarb, R.J., Gebre-Mariam, M., Hagemann, S.G., and Robert, F., (1998), Orogenic gold deposits: A proposed classification in the context of their crustal distribution and relationship to their gold deposit types: Ore Geology Reviews, v. 13, p. 7–27.

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Makar, B. (2005) “Burnside Joint Venture Esmeralda -July – 2005 Review” Burnside Operations Pty Ltd

Makar, B. and Chow, K. C. (2003) mineral resource Statement Union Reefs Gold Mine at 30 September 2003. AngloGold Australia Limited

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Makar, B. and Muller, F. (2006) “Burnside Joint Venture South Enterprise Resource Report MLN 13” Internal mineral resource report for Burnside Joint Venture group

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Mathews K.E., Hoek E., Wyllie, D.C., and Steward, S.B.V. 1980. Prediction of stable excavation spans for mining at depths below 1000m in hard rock. Centre for Mining and Excavation Technology (CANMET).

Matthaei, S.K., Henley, R.W. and Heinrich, C.A., (1995a) Gold precipitation by fluid mixing in bedding parallel fractures near carbonaceous slates at the Cosmopolitan Howley gold deposit, northern Australia. Economic Geology 90, 2123 – 2142.

Matthaei, S.K., Henley, R.W., Bacigalupo-Rose, S.B., Binns, R.A., Andrew, A.S., Carr, G.R., French, D.H., McAndrew, J. and Kananagh, M.E., (1995b) Intrusion-related high temperature gold quartz veining in the Cosmopolitan Howley metasedimentary rock-hosted gold deposit, Northern Territory, Australia. Economic Geology 90, 1012 – 1045.

McEnhill, I. (2013) “Crocodile Gold Australia Prospect Underground Mining Geotechnical Review” Report GCPL-Prospect-310113

McGuire, T., Muller, F., Sheehan, E. (2007) “Resource Update Czarina Gold Project” NI43-101 compliant report created for GBS Gold

McKibben, J. A., Standing, C. A., Gerritsen, H. (2008) GBS Gold International: Northern Territory Gold Projects, Australia. NI43-101 Technical Report. Snowden Mining Industry Consultants Pty Ltd and HG Mine Design Pty Ltd.

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Milligan, I. M. (1990) Report on Exploration for The year to 19th November 1990. Mineral Claims N20-N43, N1303-1313 & Exploration License 4810. Report for Magnum Resources

Milligan, I.M (1997) Report on Activities Mineral Claims N20-43, N1303-1313 & N3578 October 1992-July 1997. Northern Territory. Scanned Internal report for Magnum Resources

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Muller, F.L., Gerritsen, H., Edwards, M., (2011), Report on the Resources & Reserves of the Northern Territory Gold and Base Metals Properties, ForCrocodile Gold Corp, April 4th 2011.

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Zerovitch, C. (1994). Contrasting styles of gold mineralization at the Western Arm and Bridge Creek deposits, Adelaide River, NT. BSc Honors, University of Western Australia, Perth. Unpublished.

Websites used
 
http://www.kitco.com/
 
http://www.vistagold.com/
 
http://thormining.com
 
http://www.abs.gov.au
 
https://www.sedar.com
 
http://web.cim.org/UserFiles/File/CIM_DEFINITON_STANDARDS_Nov_2010.pdf
 
http://www.jorc.org/docs/jorc_code2012(4).pdf
 
Northern Territory and Federal Government Acts
 
Mineral Title Act 2010
 
         http://www.nt.gov.au/d/Minerals_Energy/Content/File/MineralTitlesAct/MineralTitlesAct.pdf
 
Mining Act
 
         http://notes.nt.gov.au/dcm/legislat/legislat.nsf/linkreference/MINING%20MANAGEMENT%20REGULATIONS
 
Mine Management Act
 
         http://notes.nt.gov.au/dcm/legislat/legislat.nsf/linkreference/MINING%20MANAGEMENT%20ACT
 
Mineral Royalty Act 2011
 
         http://notes.nt.gov.au/dcm/legislat/legislat.nsf/linkreference/MINERAL%20ROYALTY%20ACT
 
Native Title Act 1993
 
         http://www.comlaw.gov.au/Details/C2007C00406
 
Aboriginal Land use Act 1976
 
         http://www.comlaw.gov.au/Details/C2007C00360

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28 SIGNATURE PAGE

CERTIFICATE OF QUALIFIED PERSON

I, Murray Smith, BEng (Mining), MAusIMM (CP), do hereby certify that:

1.

I am a Principal Mining Consultant employed by Mining Plus Pty Ltd, with offices at:

 

 

Level 27, 459 Collins Street
Melbourne,
Victoria, 3000
Australia

 

 

Telephone: 61-3-9670-3410

Email: murray.smith@miningplus.com.au

 

 

2.

I graduated with a Bachelor of Engineering (Mining) degree from the Western Australian School of Mines (WASM), a regional campus of Curtin University, in 1993.

 

 

3.

I am a Member and Chartered Engineer of the Australasian Institute of Mining & Metallurgy (MAusIMM (CP)) – Membership No. 111064.

 

 

4.

I have worked for more than 20 years in underground mining including more than 10 years in gold mining operations. I am familiar with and have visited a variety of styles of mineral deposits worldwide, with a particular emphasis on precious metals.

 

 

5.

I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

 

 

6.

I am responsible for Sections 1-3, 15.1, 15.2, 15.5, 16.1, 16.2, 18.1, 18.2, 18.3, 21.1, 21.2, 22, 24, 25, 26 and 27 of this technical report titled “REPORT ON THE MINERAL RESOURCES AND MINERAL RESERVES OF THE NORTHERN TERRITORY OPERATIONS IN THE NORTHERN TERRITORY, AUSTRALIA FOR NEWMARKET GOLD INC.”, effective December 31, 2015 and dated 21 March, 2016 (the “Report”).

 

 

7.

I visited the Property in March 2014 for a duration of 3 days.

 

 

8.

As of the effective date of this Report, to the best of my knowledge, information and belief, this Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

 

 

9.

Mining Plus Pty Ltd and myself are independent of Newmarket Gold Inc. and its subsidiaries, applying the test set out in Section 1.5 of NI 43-101.

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December 2015 Northern Territory Operations

10.

I have read NI 43-101 and Form 43-101F1, and this Report has been prepared in compliance with that instrument and form.

 

Dated this 21 of March, 2016.


Murray Smith, B.Eng.(Mining), MAusIMM (CP)

PRINCIPAL MINING CONSULTANT

MINING PLUS Pty. Ltd.

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CERTIFICATE OF QUALIFIED PERSON

I, Mark Edwards, BSC., MAusIMM (CP), MAIG do hereby certify that:

1.

I am the General Manager of Exploration for Newmarket Gold Inc. (Newmarket Gold), with offices at:

   

1/48 Smith Street
Darwin,
Northern Territory, 0800
Australia

   

Telephone: 61-8-8982-4444

Email: medwards@newmarketgoldinc.com

   
2.

I graduated with a Bachelor of Science (Geology) degree from Flinders University in 1997.

   
3.

I graduated with a Bachelor of Science (Geology) Honors Degree from the University of Tasmania in 1998.

   
4.

I am a Member of the Australasian Institute of Mining & Metallurgy (MAusIMM - CP) – Membership No. 220787.

   
5.

I am a Member of the Australian Institute of Geoscientists (MAIG) – Membership No. 3655

   
6.

I have worked as a geologist since graduation and have over 18 years’ experience in the exploration, development and mining of mineral properties in, Australia and Botswana and am familiar with and have visited a variety of styles of mineral deposits worldwide, with particular emphasis on precious metals.

   
7.

I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for the purposes of NI 43-101.

   
8.

I am responsible for Sections 1-14, 15.3, 15.4, 15.5, 16.3, 16.4, 17, 18.1, 18.4, 18.5, 19, 20, 21.3, 21.4 and 22-27 of this technical report titled “REPORT ON THE MINERAL RESOURCES AND MINERAL RESERVES OF THE NORTHERN TERRITORY OPERATIONS IN THE NORTHERN TERRITORY, AUSTRALIA FOR NEWMARKET GOLD INC.”, effective December 31, 2015 and dated 21 March, 2016 (the “Report”).

   
9.

I work directly for Newmarket Gold in the region and have regular visits to site as required in my role as General Manager of Exploration. I regularly work on site and I last visited the site on March 2, 2016.

   
10.

I am not independent of Newmarket Gold applying the test set out in Section 1.5 of NI 43-101.

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December 2015 Northern Territory Operations

11.

As of the effective date of this Report, to the best of my knowledge, information and belief, this Report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.

   
12.

I have read NI 43-101 and Form 43-101F1, and this Report has been prepared in compliance with that instrument and form.

 

Dated this 21 of March, 2016.


Mark Edwards, B.Sc., MAusIMM (CP), MAIG

GENERAL MANAGER

NEWMARKET GOLD INC.

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