Costerfield Operation, Victoria, Australia NI 43-101 ... · PDF fileDanny Kentwell, MSc Mathematics & Planning (Geostatistics), FAusIMM, Principal Consultant . ... Project Manager:

Costerfield Operation, Victoria, Australia NI 43-101 Technical Report

Prepared for:

Mandalay Resources Corporation 76 Richmond Street Toronto ONTARIO MSC1P1 Canada

Prepared by:

SRK Consulting (Australia) Pty Ltd

ABN 56 074 271 720 Level 1, 10 Richardson Street West Perth Western Australia 6005

SRK Project Number: PLI022

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

Simon Walsh, BSc (Extractive Metallurgy & Chemistry), MBA (Hons), MAusIMM (CP), GAICD, Associate Principal Consultant (Metallurgy)

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

Date of Report: 17 March 2017

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FAIR/EBBE/swee PLI022_Costerfield_Operations_NI_43 101_Dec 2016_Rev1 March 2017

Date and Signature Page

SRK Project Number: PLI022

SRK Consulting (Australasia) Pty Ltd Level 1, 10 Richardson Street West Perth Western Australia 6005

Mandalay Resources Corporation Costerfield Operation Victoria, Australia NI 43-101 Technical Report

Project Manager: Peter Fairfield

Date of Report: 17 March 2017

Effective Date: 31 December 2016

Signature Qualified Persons:

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

Simon Walsh, BSc (Extractive Metallurgy & Chemistry), MBA, MAusIMM(CP), GAICD, Associate Principal Consultant Peter Fairfield, BEng (Mining), FAusIMM (No: 106754), CP(Min), Principal Consultant

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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 Mandalay Resources Corporation. 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 Mandalay Resources Corporation’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 Mandalay Resources Corporation’s production, development and exploration activities; the timelines for Mandalay Resources Corporation’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. Mandalay Resources Corporation 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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NON-IFRS MEASURES

This technical report contains certain non – International Financial Reporting Standards measures. Such measures have non standardised meaning under International Financial Reporting Standards (IFRS) and may not be comparable to similar measures used by other issuers.

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Table of Authors and Qualified Persons (QP)

Section Description Nominated QP Contributing Authors

Executive Summary Peter Fairfield April Westcott Chris Davis

2 Introduction Peter Fairfield Peter Fairfield

3 Reliance on Experts Peter Fairfield

4 Property Description and Location Peter Fairfield April Westcott

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography Peter Fairfield April Westcott

6 History Danny Kentwell April Westcott

7 Geological Setting and Mineralisation Danny Kentwell April Westcott Cael Gniel

8 Deposit Types Danny Kentwell April Westcott Cael Gniel

9 Exploration Danny Kentwell April Westcott Cael Gniel

10 Drilling Danny Kentwell April Westcott

11 Sampling Preparation, Analyses and Security Danny Kentwell Craig Winter

April Westcott Cael Gniel

12 Data Verification Danny Kentwell Chris Davis

13 Mineral Processing and Metallurgical Testing Simon Walsh Damon Buchanan

14 Mineral Resource Estimates Danny Kentwell Kirsty Sheerin Chris Davis Cael Gniel

15 Mineral Reserve Estimates Peter Fairfield Chloe Cavill Ross Laity

16 Mining Methods Peter Fairfield Chloe Cavill

17 Recovery Methods Simon Walsh Damon Buchanan

18 Project Infrastructure Peter Fairfield Chloe Cavill

19 Market Studies and Contracts Peter Fairfield Andre Booyzen Manfred Ruff

20 Environmental Studies, Permitting and Social, or Community Impact Peter Fairfield Adam Place

April Westcott

21 Capital and Operating Costs Peter Fairfield Chloe Cavill

22 Economic Analysis Peter Fairfield Chloe Cavill

23 Adjacent Properties Peter Fairfield April Westcott

24 Other Relevant Data and Information Peter Fairfield Peter Fairfield

25 Interpretation and Conclusions Peter Fairfield Danny Kentwell

26 Recommendations Peter Fairfield Danny Kentwell

27 References Peter Fairfield

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List of Abbreviations Abbreviation Meaning

2D two dimensional 3D three dimensional AAS atomic absorption spectroscopy AGD AGD Mining Pty Ltd 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 AuEq 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 CRF cemented rock fill dmt dry metric tonne DEDJTR Department of Economic Development, Jobs, Transport and Resources DELWP Department of Environment, Land, Water and Planning DTM digital terrain model EM electromagnetic EPA Environmental Protection Agency EVCs Ecological Vegetation Classes FAR fresh air rise FAusIMM Fellow of The Australasian Institute of Mining and Metallurgy Federation Federation Resources NL 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

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Abbreviation Meaning

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 kg kilogram kL kilolitre km kilometre 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 life of mine L/s litres per second LRGM LRGM Consultants M million Ma million years Mandalay Mandalay Resources Corporation Pty Ltd

MAusIMM(CP) Chartered Professional Member of The Australasian Institute of Mining & Metallurgy

Metcon Metcon Laboratories MEM Mid-East Minerals NL mg/kg milligrams per kilogram mg/L milligrams per litre mH metres high ML million litres mm millimetres MMI mobile metal ion MODA McArthur Ore Deposit Assessments Pty Ltd Moz million ounces mRL metres reduced level MRSD Act Mineral Resources (Sustainable Development) Act 1990

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Abbreviation Meaning

Mtpa million tonnes per annum m3 cubic metres m3/s cubic metre per second m3/s/KW cubic metre per second per kilowatt MVA megawatt ampere mW metres 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 RBA Silurian Regional Basement Aquifer RC reverse circulation RO reverse osmosis

ROM run-of-mine RPD relative paired difference SAA recent shallow alluvial aquifer Sb antimony 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 WPV Work Plan Variation

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Table of Contents

Important Notice .......................................................................................................................................... iii

List of Abbreviations .................................................................................................................................... vi

1 Summary ....................................................................................................................... 1

1.1 Introduction ......................................................................................................................................... 1

1.2 Property Description and Location ...................................................................................................... 1 1.3 Accessibility, Climate, Local Resources, Infrastructure and Physiography ........................................ 1

1.4 History ................................................................................................................................................. 2

1.5 Geological Setting and Mineralisation................................................................................................. 2

1.6 Deposit Types ..................................................................................................................................... 3

1.7 Mineral Resource Estimates ............................................................................................................... 3

1.8 Mining Methods ................................................................................................................................... 4 1.9 Mineral Reserve Estimate ................................................................................................................... 4

1.10 Metallurgy and Recovery Methods ..................................................................................................... 5

1.11 Project Infrastructure ........................................................................................................................... 6

1.12 Underground Infrastructure ................................................................................................................. 6

1.13 Services .............................................................................................................................................. 6 1.14 Contracts ............................................................................................................................................. 6

1.15 Environmental Studies, Permitting and Social Impacts ...................................................................... 7

1.16 Capital and Operating Costs ............................................................................................................... 7

1.17 Offsite Costs ........................................................................................................................................ 8

1.18 Economic Analysis .............................................................................................................................. 8

1.19 Adjacent Properties ............................................................................................................................. 9 1.20 Other Relevant Data and Information ................................................................................................. 9

1.21 Remnant Mining .................................................................................................................................. 9

1.22 Interpretation and Conclusions ........................................................................................................... 9

1.23 Recommendations .............................................................................................................................. 9

2 Introduction ................................................................................................................ 11

2.1 Scope of Work ................................................................................................................................... 11

2.2 Work Program ................................................................................................................................... 12 2.3 Basis of Technical Report ................................................................................................................. 12

2.4 Qualifications of SRK and SRK Team .............................................................................................. 12

2.5 Acknowledgement ............................................................................................................................. 13

2.6 Declaration ........................................................................................................................................ 13

3 Reliance on Other Experts ......................................................................................... 15

3.1 Marketing .......................................................................................................................................... 15

4 Property Description and Location ........................................................................... 16

4.1 Property Location .............................................................................................................................. 16

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4.2 Land Tenure ...................................................................................................................................... 17

4.3 Underlying Agreements .................................................................................................................... 20 4.4 Environmental Liability ...................................................................................................................... 20

4.5 Royalties ........................................................................................................................................... 20

4.6 Taxes ................................................................................................................................................ 21

4.7 Legislation and Permitting ................................................................................................................. 21

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ......... 23

5.1 Accessibility ....................................................................................................................................... 23

5.2 Land Use ........................................................................................................................................... 23 5.3 Topography ....................................................................................................................................... 23

5.4 Climate .............................................................................................................................................. 23

5.5 Infrastructure and Local Resources .................................................................................................. 24

6 History ......................................................................................................................... 26

6.1 Introduction ....................................................................................................................................... 26

6.2 Ownership and Exploration Work ..................................................................................................... 26

Mid-East Minerals (1968 - 1971) ........................................................................................... 26 Metals Investment Holdings (1971) ....................................................................................... 27

Victorian Mines Department (1975 - 1981) ........................................................................... 27

Federation Resources NL (1983 - 2000)............................................................................... 27

Australian Gold Development NL / Planet Resources JV (1987 - 1988) .............................. 27

Australian Gold Development NL (1987 - 1997) ................................................................... 27 AGD Operations Pty Ltd (2001 - 2009) ................................................................................. 27

6.3 Historical Resource and Reserve Estimates .................................................................................... 32

6.4 Historical Production ......................................................................................................................... 34

7 Geological Setting and Mineralisation ...................................................................... 35

7.1 Regional Geology.............................................................................................................................. 35

7.2 Property Geology .............................................................................................................................. 36

7.3 Stratigraphy of the Costerfield Formation ......................................................................................... 39 7.4 Structural Geology of the South Costerfield area ............................................................................. 41

8 Deposit Types ............................................................................................................. 42

8.1 Property Mineralisation ..................................................................................................................... 42

8.2 Deposit Mineralisation ....................................................................................................................... 45

9 Exploration .................................................................................................................. 52

9.1 Costean/ Trenching ........................................................................................................................... 52

9.2 Petrophysical Analysis ...................................................................................................................... 52

9.3 Geophysics ....................................................................................................................................... 52 Ground Geophysics ............................................................................................................... 52

Airborne Geophysics ............................................................................................................. 53

9.4 Geochemistry .................................................................................................................................... 53

Mobile Metal Ion .................................................................................................................... 53

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Bedrock Geochemistry .......................................................................................................... 53

9.5 Aerial Photogrammetry Survey ......................................................................................................... 58

10 Drilling ......................................................................................................................... 59

10.1 Mandalay Resources (2009 - Present) ............................................................................................. 59

10.2 2009/ 2010 ........................................................................................................................................ 59

10.3 2010/ 2011 ........................................................................................................................................ 59

10.4 2011/ 2012 ........................................................................................................................................ 59

10.5 2012/ 2013 Cuffley Lode Drilling ....................................................................................................... 60

10.6 2014 Cuffley/ N Lode Drilling ............................................................................................................ 61 10.7 2015 Cuffley/ N Lode/ Cuffley Deeps/ Sub King Cobra Drilling ........................................................ 61

10.8 2016 Cuffley Deeps/ Cuffley South/ M and New Lode/ Sub King Cobra Drilling/ Margaret/ Brunswick 61

10.9 Drilling Methods ................................................................................................................................ 65

10.10 Collar Surveys ................................................................................................................................... 67

10.11 Downhole Surveys ............................................................................................................................ 67 10.12 Data management............................................................................................................................. 67

10.13 Logging Procedures .......................................................................................................................... 67

10.14 Drilling Pattern and Quality ............................................................................................................... 68

Augusta ............................................................................................................................. 68

Cuffley ............................................................................................................................... 68

10.15 Interpretation of Drilling Results ........................................................................................................ 68 10.16 Factors That Could Materially Impact Accuracy of Results .............................................................. 69

11 Sample Preparation, Analyses, and Security ........................................................... 71

11.1 Sampling Techniques........................................................................................................................ 71

Diamond Core Sampling ....................................................................................................... 71

Underground Face Sampling ................................................................................................ 71

11.2 Data Spacing and Distribution .......................................................................................................... 72 11.3 Testing Laboratories ......................................................................................................................... 72

11.4 Sample Preparation .......................................................................................................................... 72

11.5 Sample Analysis ................................................................................................................................ 73

11.6 Laboratory Reviews .......................................................................................................................... 73

11.7 Assay Quality Assurance and Quality Control .................................................................................. 74

Standard Reference Material ................................................................................................ 74 Blank Material ........................................................................................................................ 77

Duplicate Assay Statistics ..................................................................................................... 78

Check Assay Program – Sample Pulps ................................................................................ 79

11.8 Sample Transport and Security ........................................................................................................ 82

12 Data Verification ......................................................................................................... 83

13 Mineral Processing and Metallurgical Testing ......................................................... 85

13.1 Metallurgical Testing ......................................................................................................................... 85

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Brunswick Testwork .............................................................................................................. 85

13.2 Ore Blend Effect on Throughput and Recovery Forecasts ............................................................... 86 13.3 Throughput ........................................................................................................................................ 86

13.4 Recovery ........................................................................................................................................... 88

Grade vs Recovery ................................................................................................................ 88

Antimony Recovery ............................................................................................................... 89

Gold Recovery ....................................................................................................................... 90

14 Mineral Resource Estimates ...................................................................................... 94

14.1 Introduction ....................................................................................................................................... 94 14.2 Diamond Drill Hole and Underground Face Sample Statistics ......................................................... 95

14.3 Data Interpretation and Domaining ................................................................................................... 99

14.4 Grade Capping ................................................................................................................................ 100

14.5 Estimation Domain Boundaries ....................................................................................................... 104

14.6 Vein Orientation Domains ............................................................................................................... 106

14.7 Bulk Density Determinations ........................................................................................................... 108 14.8 Variography ..................................................................................................................................... 108

14.9 Estimation Parameters .................................................................................................................... 115

14.10 Block Model Estimation ................................................................................................................... 126

14.11 Block Model Validation .................................................................................................................... 127

14.12 Mineral Resource Classification ...................................................................................................... 133

14.13 Mineral Resources .......................................................................................................................... 134 14.14 Cut-off Grade Calculations .............................................................................................................. 147

14.15 Reconciliation .................................................................................................................................. 147

14.16 Other Material Factors .................................................................................................................... 148

15 Mineral Reserve Estimate ........................................................................................ 149

15.1 Modifying Factors ............................................................................................................................ 149

Mining Dilution and Recovery ............................................................................................. 149 Mine Design and Planning Process .................................................................................... 152

Cut-off Grade ....................................................................................................................... 152

16 Mining Methods ........................................................................................................ 153

16.1 Introduction ..................................................................................................................................... 153

16.2 Geotechnical ................................................................................................................................... 155

16.1.1 Rock Properties ................................................................................................................... 155

16.1.2 Mine Design Parameters ..................................................................................................... 156 16.1.3 Ground Support ................................................................................................................... 158

16.3 Mine Design .................................................................................................................................... 158

16.1.4 Method Selection ................................................................................................................. 158

16.1.5 Method Description ............................................................................................................. 159

16.1.6 Materials Handling ............................................................................................................... 160

16.4 Mine Design Guidelines .................................................................................................................. 160

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16.1.7 Level Development .............................................................................................................. 160

16.1.8 Vertical Development .......................................................................................................... 160 16.1.9 Stoping ................................................................................................................................ 160

16.1.10 Mine Design Inventory .................................................................................................... 160

16.5 Ventilation ....................................................................................................................................... 164

16.1.11 Ventilation Circuit ............................................................................................................ 164

16.6 Backfill 166

16.7 Mineral Reserve Schedule .............................................................................................................. 167 16.1.12 Development Schedule ................................................................................................... 167

16.1.13 Equipment Requirements ............................................................................................... 167

16.1.14 Personnel ........................................................................................................................ 168

16.8 Schedule Summary ......................................................................................................................... 168

17 Recovery Methods .................................................................................................... 170

17.1 Brunswick Processing Plant ............................................................................................................ 170

Crushing and Screening Circuit .......................................................................................... 170 Milling Circuit ....................................................................................................................... 170

Flotation Circuit ................................................................................................................... 170

Concentrate Thickening and Filtration ................................................................................ 171

Tailings Circuit ..................................................................................................................... 171

Throughput .......................................................................................................................... 171

Recovery ............................................................................................................................. 171 Concentrate Grade .............................................................................................................. 171

17.2 Services .......................................................................................................................................... 172

Water ................................................................................................................................... 172

Air ........................................................................................................................................ 172

Power .................................................................................................................................. 172 17.3 Plant Upgrades ............................................................................................................................... 174

Crushing and Screening Circuit .......................................................................................... 174

Milling Circuit ....................................................................................................................... 174

Flotation Circuit ................................................................................................................... 174

Concentrate Thickening and Filtration ................................................................................ 174

Tailings Circuit ..................................................................................................................... 174 Recovery Projects ............................................................................................................... 175

Reagent Mixing and Storage ............................................................................................... 175

18 Project Infrastructure ............................................................................................... 176

18.1 Surface Infrastructure...................................................................................................................... 176

18.2 Underground Infrastructure ............................................................................................................. 178

Secondary Means of Egress ............................................................................................... 178

Refuge Chambers ............................................................................................................... 178 Compressed Air ................................................................................................................... 178

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Ventilation System ............................................................................................................... 179

Dewatering System ............................................................................................................. 179 Infrastructure ....................................................................................................................... 179

18.3 Tailings Storage .............................................................................................................................. 179

18.4 Power Supply .................................................................................................................................. 180

Power Reticulation .............................................................................................................. 180

18.5 Water Supply ................................................................................................................................... 181

18.6 Water Management......................................................................................................................... 181 18.7 Waste Rock Storage ....................................................................................................................... 181

18.8 Augusta to Brunswick ROM Pad Transport .................................................................................... 182

18.9 Diesel Storage ................................................................................................................................. 182

18.10 Explosives Storage ......................................................................................................................... 182

18.11 Maintenance Facilities .................................................................................................................... 182

18.12 Housing and Land ........................................................................................................................... 182

19 Market Studies and Contracts ................................................................................. 183

19.1 Concentrate Transport .................................................................................................................... 183

19.2 Marketing ........................................................................................................................................ 183

Critical/Strategic Supply Reviews ....................................................................................... 185

19.3 Contracts ......................................................................................................................................... 186

20 Environmental Studies, Permitting, and Social or Community Impact ............... 187

20.1 Environment and Social Aspects .................................................................................................... 187 Mine Ventilation ................................................................................................................... 187

Water Disposal .................................................................................................................... 187

Waste Rock ......................................................................................................................... 188

Tailings Disposal ................................................................................................................. 188

Air Quality ............................................................................................................................ 188

Groundwater ........................................................................................................................ 189 Noise ................................................................................................................................... 190

Blasting and Vibration ......................................................................................................... 190

Native Vegetation ................................................................................................................ 190

Visual Amenity ................................................................................................................ 191

Heritage ........................................................................................................................... 191

Community ...................................................................................................................... 191 Mine Closure and Revegetation ..................................................................................... 192

20.2 Regulatory Approvals ...................................................................................................................... 192

Work Plan Variation ............................................................................................................. 192

Other Permitting .................................................................................................................. 192

21 Capital and Operating Costs ................................................................................... 194

21.1 Capital Costs ................................................................................................................................... 194

Processing Plant ................................................................................................................. 194

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Administration ...................................................................................................................... 195

Environmental ..................................................................................................................... 195 Mining .................................................................................................................................. 195

Capital Development ........................................................................................................... 195

Closure ................................................................................................................................ 195

21.2 Operating Costs .............................................................................................................................. 195

Lateral Development ........................................................................................................... 196

Production Stoping .............................................................................................................. 197 Mining Administration .......................................................................................................... 197

Geology ............................................................................................................................... 197

ROM Haulage ...................................................................................................................... 197

21.3 Processing Plant ............................................................................................................................. 197

21.4 Site Services ................................................................................................................................... 197

21.5 General and Administration ............................................................................................................ 198 21.6 Selling Expenses ............................................................................................................................. 198

22 Economic Analysis ................................................................................................... 199

23 Adjacent Properties .................................................................................................. 200

23.1 General Statement about Adjacent Properties ............................................................................... 200

24 Other Relevant Data and Information ..................................................................... 202

24.1 Remnant Mining .............................................................................................................................. 202

25 Interpretation and Conclusions............................................................................... 203

26 Recommendations ................................................................................................... 204

27 References ................................................................................................................ 205

List of Tables Table 2-1: List of Qualified Persons ............................................................................................................ 14

Table 4-1: Granted Tenement Details ......................................................................................................... 17

Table 6-1: Historical Drilling Statistics for the Costerfield Property ............................................................. 26

Table 6-2: Historical Mineral Resources – Mandalay Resources – Costerfield Project .............................. 33

Table 6-3: Historical Mineral Reserves – Mandalay Resources – Costerfield Project ................................ 33 Table 6-4: Historical Mine Production: Mandalay Resources – Costerfield Project .................................... 34

Table 9-1: Auger drill hole totals by prospect .............................................................................................. 57

Table 10-1: Drill hole summary ..................................................................................................................... 59

Table 10-2: Significant Cuffley Main, Cuffley Deeps, Sub King Cobra, Brunswick, Margaret, N Lode and New Vein Lode drill hole intercepts: 2016 .......................................................................... 63

Table 11-1: Summary of Onsite gold duplicate statistics .............................................................................. 78 Table 11-2: Summary of Onsite original vs Onsite duplicate, ALS, Bureau Veritas gold duplicate

statistics ..................................................................................................................................... 80

Table 11-3: Summary of Onsite original vs Onsite duplicate, ALS, Bureau Veritas antimony duplicate statistics ..................................................................................................................................... 80

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Table 13-1: Brunswick Processing Plant key processing results – January 2014 to December 2016 ......... 92

Table 14-1: Changes made to models at year-end 2016 .............................................................................. 94 Table 14-2: Face and diamond drilling sample statistics .............................................................................. 95

Table 14-3: Sample statistics of the data selected for analysis .................................................................... 99

Table 14-4: CM lode Domain 3 composite file with new AuACC_BC and AuACC_NC variables created, and grade capping (200 g/t) applied to AuACC ......................................................... 100

Table 14-5: Accumulated gold statistics before and after top cuts ............................................................. 102

Table 14-6: Variogram Model Parameters .................................................................................................. 117 Table 14-7: Search Parameters for top cut estimate .................................................................................. 120

Table 14-8: Search Parameters for influence limitation estimate ............................................................... 125

Table 14-9: Block Model Dimensions .......................................................................................................... 126

Table 14-10: Mine Planning Regularised Block Model Dimensions ............................................................. 126

Table 14-11: Declustered thickness weighted, composited sample grades and widths compared to estimated model grades and true thickness values ................................................................. 128

Table 14-12: Mineral Resources at Costerfield, Inclusive of Mineral Reserves, as of 31 December 2016 .. 134

Table 14-13: Summary of the Augusta, Cuffley and Brunswick Mineral Resource, inclusive of Mineral Reserve .................................................................................................................................... 135

Table 15-1: Mineral Reserves at Costerfield, as of 31 December 2016 ..................................................... 149

Table 15-2: Costerfield Mine recovery and dilution assumptions ............................................................... 151

Table 16-1: Stope and Development Inventory by lode .............................................................................. 161 Table 16-2: Development and Stoping Inventory by level for major lodes (CE, CS, CSE, NSP2-4,

NV44, NW, P2 and W lodes not included) ............................................................................... 161

Table 16-3: Development Rates .................................................................................................................. 167

Table 16-4: Underground Mobile Equipment Fleet ..................................................................................... 168

Table 16-5: Summary of design physicals .................................................................................................. 168 Table 18-1: Current Augusta Licence - Maximum quantities and types of explosives ............................... 182

Table 20-1: Permit requirements ................................................................................................................. 193

Table 21-1: Costerfield Operation – Capital Cost Estimate ........................................................................ 194

Table 21-2: Costerfield Operation – Operating Cost Estimate .................................................................... 196

Table 21-3: Operating Cost Inputs .............................................................................................................. 196

Table 21-4: Summary of Development requirements ................................................................................. 196 Table 23-1: Ownership of Augusta Mine adjacent properties ..................................................................... 200

Table 23-2: Distance from the Augusta Mine site to significant mining projects ......................................... 201

List of Figures Figure 4-1: Costerfield Operation Location .................................................................................................. 16

Figure 4-2: Plan of Area – Mining Licence No. 4644 .................................................................................... 18

Figure 4-3: Current Mandalay Resources Mining and Exploration Lease boundaries ................................. 19

Figure 5-1: Monthly average temperature and rainfall ................................................................................. 24

Figure 5-2: Augusta box-cut, portal and workshop ....................................................................................... 24 Figure 5-3: Aerial view of Brunswick Processing Plant and Brunswick Open Pit ........................................ 25

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Figure 7-1: Geological map of the Heathcote – Colbinabbin – Nagambie Region ...................................... 36

Figure 7-2: Costerfield property geology and old workings .......................................................................... 38 Figure 7-3: Regional stratigraphic chart of the Costerfield region illustrating age relationships of

defined sedimentary succession of the Darraweit Guim Province ............................................ 39

Figure 7-4: Stratigraphic column of the Costerfield Formation illustrating the relative positions of newly defined (informal) stratigraphic units................................................................................ 40

Figure 8-1: Paragenetic history and vein genesis of the Costerfield region ................................................. 43

Figure 8-2: Two views of E Lode at 1,070 Level South, looking south, with annotation on right-hand side 44 Figure 8-3: Close-up of mineralisation in west-dipping E Lode at 1070 Level, looking south ...................... 44

Figure 8-4: Schematic cross section 5050N northing, illustrating the ‘apparent offset’ between the Cuffley Deeps lode and Cuffley lodes across the west-dipping series of faults .................. 48

Figure 8-5: Brunswick cross section 5880N ................................................................................................. 49

Figure 9-1: Augusta South aircore program results with lode locations ....................................................... 54

Figure 9-2: Augusta South (Margaret Zone) – Auger drill 2011 bedrock geochemistry: antimony contours ..................................................................................................................................... 55

Figure 9-3: Auger geochemistry results displayed as antimony contours .................................................... 56

Figure 10-1: Known drill hole collar locations for Augusta and Cuffley including 2016 drilling ...................... 66

Figure 10-2: Example of cross section at 4,300 mN post drilling geological interpretation of the Augusta deposit ........................................................................................................................................ 69

Figure 11-1: MR30-20 Gold Standard Reference Material – Assay Results for 2016 ................................... 75 Figure 11-2: G310-6 Gold Standard Reference Material – Assay Results for 2016 ...................................... 75

Figure 11-3: MR04/6 Gold Standard Reference Material – Assay Results for 2016 ..................................... 76

Figure 11-4: MR30/20 Sb Standard Reference Material – Assay Results for 2016 ....................................... 76

Figure 11-5: MR04/6 Antimony Standard Reference Material – Assay Results for 2016 .............................. 77

Figure 11-6: Au Blank assay results for 2016 ................................................................................................ 77 Figure 11-7: Sb Blank assay results for 2016 ................................................................................................ 78

Figure 11-8: Scatter plot for Onsite gold duplicates (g/t) for 2016 ................................................................. 79

Figure 11-9: Relative paired difference plot for Onsite gold duplicates (g/t) in 2016 ..................................... 79

Figure 11-10: Relative pair difference plot for Onsite Original vs Onsite duplicate, ALS, BV gold duplicates (g/t) ............................................................................................................................................. 81

Figure 11-11: Relative pair difference plot for Onsite original vs Onsite duplicate, ALS, BV antimony duplicates (%) with BV outliers .................................................................................................. 81

Figure 11-12: Scatter plot for Onsite original vs Onsite duplicate, ALS, BV antimony duplicates (%) ............. 82

Figure 11-13: Scatter plot for Onsite original vs Onsite duplicate, ALS, BV gold duplicates (g/t) .................... 82

Figure 12-1: Cuffley Main lower drive south end showing mineralised structures ......................................... 84

Figure 13-1: Historic Brunswick Processing Plant throughput: Apr 2007 to Dec 2016 .................................. 87

Figure 13-2: Historic Brunswick Processing Plant throughput: Jan 2012 to Dec 2016 .................................. 87 Figure 13-3: Historic Brunswick Processing Plant throughput: Jan 2014 to Dec 2016 .................................. 88

Figure 13-4: Cuffley/ Augusta blend feed grade vs recoveries: 2014 - 2016 ................................................. 89

Figure 13-5: Antimony yield to concentrate vs upgrade for Cuffley/ Augusta ore blend: 2015 ...................... 89

Figure 13-6: Mill recoveries and throughput for Cuffley/ Augusta ore blend: 2014 – 2016 ............................ 91

Figure 13-7: Mill recoveries and throughput 2016 .......................................................................................... 93 Figure 13-8: Feed grade and recoveries 2016 ............................................................................................... 93

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Figure 14-1: Long section of data collected for sample comparison analysis showing face sample (red) and drill hole (blue) data............................................................................................................. 98

Figure 14-2: Long section of NM Lode showing fault intercepts identified as controls on mineralisation ...... 99

Figure 14-3: Estimate produced using the AuACC_BC variable compared to composites ......................... 100

Figure 14-4: AuACC (left) and AuACC_NC (right) estimates compared to composites .............................. 101

Figure 14-5: Final AuACC estimate; cells delineated by the AuACC_BC estimate have their AuACC value replaced with the AuACC_NC value (circled) ................................................................ 101

Figure 14-6: CM Lode Estimation domain boundaries and composite samples .......................................... 104 Figure 14-7: NM Lode Estimation domain boundaries and composite samples .......................................... 105

Figure 14-8: CD (220) Lode Estimation domain boundaries and composite samples ................................. 105

Figure 14-9: Brunswick Lode Estimation domain boundaries and composite samples ............................... 105

Figure 14-10: NE Lode Estimation domain boundaries and composite samples .......................................... 106

Figure 14-11: CE Lode Estimation domain boundaries and composite samples .......................................... 106

Figure 14-12: CM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor ..................................................................................................................... 107

Figure 14-13: NM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor ..................................................................................................................... 107

Figure 14-14: Bulk density determinations ..................................................................................................... 108

Figure 14-15: Major direction experimental variogram for gold accumulation in Domain 3 of CM Lode ....... 109

Figure 14-16: Semi major experimental variogram for gold accumulation in Domain 3 of CM Lode ............. 110 Figure 14-17: Back transformed final variogram model for gold accumulation in Domain 3 of CM Lode ...... 110

Figure 14-18: Long section view of CM Lode Domain 3 composites coloured by gold accumulated grade and search ellipse used ................................................................................................. 111

Figure 14-19: Major direction experimental variogram for antimony accumulation in Domain 3 of CM Lode ......................................................................................................................................... 111

Figure 14-20: Semi major experimental variogram for antimony accumulation in Domain 3 of CM Lode ..... 112

Figure 14-21: Back-transformed final variogram model for antimony accumulation in Domain 3 of CM Lode ......................................................................................................................................... 112

Figure 14-22: Long section view of CM Lode Domain 3 composites coloured by antimony accumulated grade and search ellipse used............................................................................ 113

Figure 14-23: Major direction experimental variogram for True Thickness in Domain 3 of CM Lode ............ 113 Figure 14-24: Semi major experimental variogram for True Thickness in Domain 3 of CM Lode ................. 114

Figure 14-25: Back transformed final variogram model for antimony accumulation in Domain 3 of CM Lode ......................................................................................................................................... 114

Figure 14-26: Long section view of CM Lode Domain 3 composites coloured by True Thickness grade and search ellipse used ........................................................................................................... 115

Figure 14-27: Cross section 5818 mN of Brunswick lode, composites coloured by true thickness ............... 116 Figure 14-28: NM Lode Domain 3 Gold Accumulation Swath Plot by Northing ............................................. 130

Figure 14-29: NM Lode Domain 3 Gold Accumulation Swath Plot by Elevation ............................................ 130

Figure 14-30: NM Lode Domain 2 Antimony Accumulation Swath Plot by Northing ..................................... 131

Figure 14-31: NM Lode Domain 2 Antimony Accumulation Swath Plot by Elevation .................................... 131

Figure 14-32: CM Lode Domain 1 Gold Accumulation Swath Plot by Northing ............................................. 132 Figure 14-33: CM Lode Domain 1 Gold Accumulation Swath Plot by Elevation ............................................ 132

Figure 14-34: CM Lode Domain 1 Vein Width Swath Plot by Northing .......................................................... 133

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Figure 14-35: CM Lode Domain 1 Vein Width Swath Plot by Elevation ......................................................... 133

Figure 14-36: NM Lode Resource Block Model with Resource Category Boundaries .................................. 137 Figure 14-37: CM Main Lode Block Model with Resource Category Boundaries .......................................... 137

Figure 14-38: CE Lode Block Model with Resource Category Boundaries.................................................... 138

Figure 14-39: CD (220) Lode Block Model with Resource Category Boundaries .......................................... 138



Figure 14-42: CS (212) Lode Block Model with Resource Category Boundaries .......................................... 140 Figure 14-43: CS (213) Lode Block Model with Resource Category Boundaries .......................................... 140

Figure 14-44: NV (43) Lode Block Model with Resource Category Boundaries ............................................ 141

Figure 14-45: NV (44) Lode Block Model with Resource Category Boundaries ............................................ 141

Figure 14-46: K Lode Block Model with Resource Category Boundaries ...................................................... 142

Figure 14-47: C Lode Block Model with Resource Category Boundaries ...................................................... 142

Figure 14-48: BSP Lode Block Model with Resource Category Boundaries ................................................. 143 Figure 14-49: B Lode Block Model with Resource Category Boundaries ...................................................... 143

Figure 14-50: NE Lode Block Model with Resource Category Boundaries.................................................... 144

Figure 14-51: Brunswick Lode Block Model with Resource Category Boundaries ........................................ 144

Figure 14-52: Brunswick PK Lode Block Model with Resource Category Boundaries .................................. 145

Figure 14-53: NSP1 Lode Block Model with Resource Category Boundaries ............................................... 145

Figure 14-54: NSP2 Lode Block Model with Resource Category Boundaries ............................................... 146 Figure 14-55: NSP3 Lode Block Model with Resource Category Boundaries ............................................... 146



Figure 16-1: Long section of proposed Underground Mine Design along Cuffley (base) and N Lode (top) 154

Figure 16-2: Schematic cross section of the Augusta and Cuffley systems ................................................ 155 Figure 16-3: Longhole CRF stoping method ................................................................................................ 159

Figure 16-4: Augusta Ventilation Circuit ....................................................................................................... 165

Figure 16-5: Cuffley Ventilation Circuit ......................................................................................................... 166

Figure 17-1: Brunswick Processing Plant summary flowsheet .................................................................... 173

Figure 18-1: Augusta Mine Site .................................................................................................................... 176

Figure 18-2: Brunswick Site Area ................................................................................................................. 177 Figure 19-1: Estimate of global antimony demand by end-use segment ..................................................... 184

Figure 19-2: Antimony metal prices 2010 - 2015 ......................................................................................... 184

Figure 19-3: World antimony production and consumption, 2000 - 2020 (t antimony) ................................ 185

Figure 20-1: Groundwater elevation contour map of the areas surrounding the Augusta Mine as at December 2016 ........................................................................................................................ 190

Figure 23-1: Augusta Mine adjacent properties ........................................................................................... 200

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1 Summary 1.1 Introduction

SRK Consulting (Australasia) Pty Ltd (SRK) in conjunction with Mandalay Resources Costerfield Operations Pty Ltd has prepared the Costerfield Operation NI 43-101 Technical Report for the continuing operations at Augusta Mine in connection with the addition of mineralisation sourced from the Cuffley Lode. This report updates the finding of the Mineral Resources and Mineral Reserves previously filed NI 43-101 compliant Technical Report of 3 June 2016 (SRK 2015).

This Technical Report is based on Mineral Resources that conform to Canadian Securities Administrators’ National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101).

Mandalay Resources Corporation (Mandalay) is a publicly traded company listed on the Toronto Stock Exchange and trading under the symbol MND. On 1 December 2009, Mandalay completed the acquisition of AGD Mining Pty Ltd (AGD), the sole owner of the Costerfield Operation, resulting in AGD becoming a wholly-owned subsidiary of Mandalay.

The Costerfield Operation is located within the Costerfield mining district, approximately 10 km northeast of the town of Heathcote, Victoria. The Augusta Mine has been operational since 2006 and was the sole ore source for the Brunswick Processing Plant until the commencement of mining of the Cuffley Lode in December 2013.

1.2 Property Description and Location The Costerfield Operation encompasses the underground Augusta Mine (Augusta) and the Cuffley deposit, the Brunswick Processing Plant, and associated infrastructure, all of which are located within mining licence MIN 4644. The mining licence is located within exploration lease EL 3310 which is 100% held by Mandalay Resources Costerfield Operations Pty Ltd.

The Augusta Mine is located at latitude of 36° 52’ 27” south and longitude 144° 47’ 38” east. The Brunswick Processing Plant is located approximately 2 km northwest of Augusta. The Cuffley Lode is located approximately 500 m north-northwest of the Augusta workings and is accessed by an underground decline from Augusta.

1.3 Accessibility, Climate, Local Resources, Infrastructure and Physiography Access to the Costerfield Operation is via the sealed Heathcote–Nagambie Road, which is accessed off the Northern Highway to the south of Heathcote. The Northern Highway links Bendigo with Melbourne.

The nearest significant population to Costerfield is Bendigo, with a population of approximately 100,000 people, located 50 km to the west-northwest. The Costerfield Operation is a residential operation with personnel residing throughout central Victoria and Melbourne. Local infrastructure and services are available in Heathcote, the largest town within the vicinity of the Costerfield Operation.

The Augusta mine site is located on privately held land, while the Brunswick Processing Plant is located on unrestricted Crown land. The surrounding land is largely rocky, rugged hill country administered by the Department of Environment, Land, Water and Planning (DELWP) as State Forest. The Puckapunyal Military Area is located on the eastern boundary of the Project Area.

The area has a Mediterranean climate with temperatures ranging from -2°C in winter (May to August) to +40°C in summer (November to February). Annual rainfall in the area is approximately 500 mm to



600 mm, with the majority occurring between April and October. The annual pan evaporation is between 1,300 mm and 1,400 mm.

The weather is amendable to year-round mining operations; however, construction activity is restricted to the summer months as high winter rainfall can lead to saturated ground conditions that can affect surface activities.

1.4 History Exploration for antimony gold deposits in the Costerfield area of Central Victoria started in the early 1850s and resulted in the discovery of the main Costerfield Reef in 1860. At around the same time, the Kelburn (Alison) Reef and Tait’s Reef were discovered at South Costerfield.

The Alison Mine ceased operations in 1923, while the South Costerfield / Tait’s Mine operated sporadically from the 1860s until 1978; it was the last shaft mine to operate on the field.

In 1970, Mid-East Minerals NL identified a large bedrock geochemistry anomaly south of Tait’s Shaft, which they called Tait-Margaret. This was subsequently drilled by the Mines Department in 1977 and mineralised veins were intersected.

In 2001 AGD drilled the Tait-Margaret anomaly, which was re-named Augusta. AGD commenced underground mining of the Augusta resource (N, C, W and E Lodes) in 2006. Brownfield exploration core drilling by Mandalay in 2011 located a faulted offset of the Alison Lode beneath the old Alison Mine and New Alison Mine workings. The deeper offset mineralisation was renamed the Cuffley Lode. Subsequent definition drilling throughout 2011 and 2012 resulted in an initial Inferred Resource for the Cuffley Lode being established in January 2012. Further infill and extension drilling has continued and converted more of this Inferred Resource to Measured and Indicated in 2013.

1.5 Geological Setting and Mineralisation The Costerfield Gold-Antimony field is located within the Costerfield Dome, in the Melbourne Zone, which consists of a very thick sequence of Siluro-Devonian marine sedimentary rocks.

The western boundary of the Costerfield Dome is demarcated by the Cambrian Heathcote Volcanic Belt and north-trending Mt William Fault. The Mt William Fault is a major structural terrain boundary that separates the Bendigo Zone from the Melbourne Zone. The Dome is bounded to the east by the Moormbool Fault, which has truncated the eastern limb of the Costerfield Anticline, resulting in an asymmetric dome structure.

The quartz-stibnite lodes are controlled by north-northwest trending faults and fractures located predominantly near the crest, on the western flank of the Costerfield Dome.

The Augusta Deposit currently comprises 19 lodes: E Lode, W Lode, C lode, N Main Lode, NV1 Lode, NV2 Lode, NW Lode, NE Lode, NSW Lode, NSP1 to NSP5 Lodes, P1 Lode, P2 Lode, K Lode, B Lode, and Bob Splay Lode. The Cuffley Deposit is currently comprised of eight lode structures: Cuffley Main (CM Lode), Cuffley East (CE Lode), Cuffley South (CS Lode), Cuffley South East (CSE Lode), Cuffley Deeps (CD Lode), Cuffley Deeps West (CDW lode), Cuffley Deeps Lower (CDL lode) and Alison South (AS Lode). The Brunswick deposit comprises the main lode structure (Brunswick shear) and takes in a portion of mineralisation between the Penguin and Kiwi faults that has an offset at depth to the west.

Mining has demonstrated that mineralised splay veins and oblique, cross-cutting fault structures influence grade in the north-northwest trending lodes. The lodes can vary from massive stibnite with microscopic gold to quartz-stibnite with minor visible gold, pyrite and arsenopyrite. In some cases, gold occurs in quartz veins with little or no stibnite.



1.6 Deposit Types The Costerfield Gold-Antimony Field is part of a broad gold-antimony province mainly confined within the Siluro-Devonian Melbourne Zone. Although antimony commonly occurs in an epithermal setting (in association with silver, bismuth, tellurium and molybdenum), the quartz-stibnite-gold narrow veins of the Melbourne Zone are mesothermal-orogenic and are part of a 380 - 370 Ma tectonic event. The quartz-stibnite-gold veins contain only accessory amounts of pyrite and arsenopyrite and trace amounts of galena, sphalerite and chalcopyrite. Pyrite and arsenopyrite also occur in the wall rocks in narrow alteration halos around the lodes; traces of gold are also found in the Brunswick Reef wall rocks. Gold in Central Victoria is believed to have been derived from underlying Cambrian greenstones. The origin of the antimony is less certain.

The mineralisation occurs as narrow veins or lodes, typically ≤500 mm wide, hosted within low-grade (anchizone) mudstone and siltstone of the Lower Silurian Costerfield formation. Mineralised shoots in the Costerfield Property are structurally controlled by the intersection of the lodes with major crosscutting, puggy, sheared fault structures.

1.7 Mineral Resource Estimates The Mineral Resources are stated here for the Augusta, Cuffley and Brunswick deposits with an effective date of 31 December 2016. The Mineral Resource is depleted for mining up to 31 December 2016. The estimate excludes remnant mineralisation that remains in pillars.

The Mineral Resources are reported at a cut-off grade of 3.5 g/t gold equivalent (AuEq), with a minimum mining width of 1.2 m. The gold equivalence formula used is calculated using typical recoveries at the Costerfield processing plant and using a gold price of USD1,400 per troy ounce and an antimony price of USD10,000 per tonne as follows:

• AuEq = Au (g/t) + 1.76 x Sb (%)

All relevant diamond drill hole and underground face samples in the Costerfield Property, available as of 30 November 2016 for the Augusta, Cuffley and Brunswick deposits, were used to inform the estimate. The estimation methodology was amended slightly (changes were made to the influence limitation grade capping estimation) from the approach used by SRK in the December 2015 Resource Estimate reported in SRK’s NI 43-101 Technical Report on the Property (SRK, 2015).

Details of the Augusta, Cuffley and Brunswick Mineral Resources are given in Table ES-1.

Table ES-1: Mineral Resources at Costerfield, Inclusive of Mineral Reserves as of 31 December 2016

Category Inventory (kt)

Gold Grade (g/t)

Antimony Grade

(%)

Contained Gold (koz)

Contained Antimony

(kt)

Measured 286 9.5 4.0 88 11.4

Indicated 812 5.9 2.5 155 20.6

Measured + Indicated 1,098 6.9 2.9 242 32.0

Inferred 611 5.5 1.5 108 9.0

Notes:

1) Mineral Resources estimated as of December 31, 2016, and depleted for production through December 31, 2016. 2) Mineral Resources stated according to CIM guidelines and include Mineral Reserves. 3) Tonnes and contained gold (oz) rounded to the nearest thousand; contained antimony (t) rounded to nearest hundred. 4) Totals may appear different from the sum of their components due to rounding. 5) A cut-off grade of 3.5 g/t AuEq over a minimum mining width of 1.2 m is applied where AuEq is calculated at a gold price

of USD1,400 /oz, antimony price of USD$10,000 /t and exchange rate USD:AUD of 0.75.



6) The AuEq value is calculated using the formula: AuEq = Au g/t + 1.76 * Sb %. 7) Geological modelling and sample compositing was performed by Cael Gniel, who is a full-time employee of Mandalay

Resources and Chris Davis, MAusIMM, who is a full-time employee of Mandalay Resources, and was independently verified by Danny Kentwell, FAusIMM, who is a full-time employee of SRK Consulting.

8) The Mineral Resource estimation was performed by Kirsty Sheerin, MAusIMM, who is a full-time employee of SRK Consulting, Cael Gniel, who is a full-time employee of Mandalay Resources and Chris Davis, MAusIMM, who is a full-time employee of Mandalay Resources. The resource models were verified by Danny Kentwell, FAusIMM, who is a full-time employee of SRK Consulting. Danny Kentwell, FAusIMM, full-time employee of SRK Consulting, is the Qualified Person under NI 43-101 and the Competent Person for the Resource.

1.8 Mining Methods The Augusta Mine incorporating the Augusta Lodes and Cuffley Lode is an underground antimony-gold mine currently producing approximately 150 ktpa of mill feed from a variety of mining methods to a depth of over 300 m below surface.

The Augusta Mine employs longitudinal uphole retreat mining methods working a bottom up sequence that has been utilised throughout 2010 to 2015. Cemented Rock Fill (CRF) is placed into stoping voids to maximise extraction and assist with mine stability.

Mandalay has developed the Cuffley Lode of which the southern boundary is located approximately 250 m to the northwest of the current Augusta Mine workings. Access to the Cuffley Lode is via a single decline that connects to the existing Augusta decline at the 1030 mRL. The Augusta Mine is scheduled to produce mill feed from three different mining methods: full face development, longhole CRF stoping and uphole stoping. All mined material is transported to the Augusta box-cut before being hauled to either the Brunswick run of mine (ROM) Pad or Augusta Waste Rock Storage Facility.

1.9 Mineral Reserve Estimate From the Mineral Resource, a mine plan was designed based on Measured and Indicated Resource blocks, using predominantly the CRF blast hole stoping method. A cut-off grade of 4.0 g/t AuEq and minimum mining widths of 1.8 m for development and 1.2 m for stoping were used, with planned and unplanned dilution at zero grade. Financial viability of Proven and Probable Mineral Reserves was demonstrated at metal prices of USD1,200 /oz Au and USD8,000 /t Sb and a USD:AUD exchange rate of 0.75. The Mineral Reserve estimate presented in Table ES-2 does not include any portion of the 611,000 t of Inferred Resources.

Table ES-2: Mineral Reserves at Costerfield, as of 31 December 2016

Category Reserve (kt)

Gold Grade (g/t)

Antimony Grade

(%)


Contained Antimony

(kt)

Proven 184 8.1 3.5 48 6.4

Probable 434 5.7 2.6 80 11.1

Proven + Probable 619 6.5 2.8 128 17.5

Notes:

1) Mineral Reserve estimated as of December 31, 2016, and depleted for production through to December 31, 2016. 2) Tonnes rounded to the nearest thousand; contained gold (oz) rounded to the nearest thousand and contained

antimony (t) rounded to nearest hundred. 3) Totals are subject to rounding error. 4) Lodes have been diluted to a minimum mining width of 1.2 m for stoping and 1.8 m for ore development. 5) A 4.0 g/t AuEq cut-off grade is applied. 6) Commodity prices applied are: gold price of USD1,200 /oz, antimony price of USD8,000 /t and USD:AUD exchange rate

of 0.75. 7) The AuEq value s calculated using the formula: AuEq = Au g/t + 1.64 * Sb %. 8) The Mineral Reserve is a subset, a Measured and Indicated only schedule, of a Life of Mine Plan that includes mining of

Measured, Indicated and Inferred Resources.



9) The Mineral Reserve estimate was prepared by Chloe Cavill, BSc, who is a full-time employee of Mandalay Resources, and was independently verified by Peter Fairfield, FAusIMM, CP (Mining), who is a full-time employee of SRK Consulting and a Qualified Person under NI 43-101.

1.10 Metallurgy and Recovery Methods All mill feed is processed at the existing Brunswick Processing Plant at Costerfield, near the town of Heathcote in central Victoria. This operation started as a pilot plant and has been converted into a full-time processing plant that is capable of processing sulphide gold-antimony containing material to produce gold-antimony concentrate. The plant consists of a two-stage crushing circuit, two ball mills in series with classification, and gravity concentration in closed circuit. The flotation circuit consists of a rougher, scavenger and single stage cleaner circuit for the production of antimony-gold flotation concentrate. The gravity gold concentrates can be either blended with the final flotation concentrate and bagged for shipment to customers in China or further refined via tabling to make a gravity concentrate which is sent to a refinery. The current flotation tailings are sent to a tailings storage facility to the east of the Brunswick Processing Plant – the Brunswick tailings storage facility (TSF).

Ore from the Augusta underground mine, including the Cuffley Deposit, has been the sole feed for the Brunswick Processing Plant since mining began at Augusta in 2006. The metallurgical performance of the Augusta and Cuffley ore has been demonstrated over this period of operation and has delivered stable and consistent recoveries.

Throughput began to plateau through 2014 to 2016 as shown in Figure ES-1; there was still a modest increase from an average 12,445 t/mth in 2014 to 12,822 t/mth in 2015 and then to 12,867 t/mth in 2016. The 2016 monthly throughput was regularly over 13,000 t and as high as 13,762 t. It is expected that ongoing process optimisation and minor, low capital cost debottlenecking projects will continue the ongoing production increase.

The 2016 historical gold recovery data used to forecast the life of mine (LOM) recovery is gold recovery of 90.1% and gravity recovery of 35.7% on 10.3 g/t head grade (resultant tailings grade of 1.08 g/t, and flotation recovery of 54.3%).

Figure ES-1 shows the grade versus recovery relationship from January 2014 to December 2016.

Figure ES-1: Cuffley / Augusta Blend Feed Grade vs Recoveries 2014 to 2016



1.11 Project Infrastructure The Costerfield Operations surface facilities are representative of a modern gold-antimony mining operation. The Augusta Mine site comprises the office and administration complex, underground workshops and surface infrastructure to support the underground operations. The Brunswick site comprises the gold-antimony processing plant and associated facilities, surface workshop, tailings storage facilities; Reverse Osmosis (RO) Plant and Core farm and core processing facility.

The Splitters Creek Evaporation Facility is situated on a 30 Ha parcel of land that is located approximately 3 km from the Augusta site. The facility provides an additional 120 ML per year of net evaporation. The purpose of the facility is to evaporate groundwater extracted from the Costerfield Operations and thereby maintain dewatering rates from the underground workings.

1.12 Underground Infrastructure The Augusta Decline has been developed from a portal within a box-cut with the majority dimensions of 4.8 m high by 4.5 m wide at a gradient of 1:7 down. The majority of the decline development has been completed with a twin boom jumbo; however, development of the decline from the portal to 2 Level was completed with a road-header; this section of decline has dimensions of 4.0 m high by 4.0 m wide. The decline provides primary access for personnel, equipment and materials to the underground workings.

Access to the Cuffley Lode is via a single decline that connects to the existing Augusta decline at the 1030 mRL. The Cuffley Decline currently extends down to approximately 895 mRL. At the 935 mRL, the Cuffley Incline extends off the Cuffley Decline and accesses mineral resources from the 945 mRL to the 1020 mRL. A second decline within Cuffley known as the 4800 Decline was commenced during 2014 to access the southern part of the Cuffley Lode which is positioned south of the East Fault. This Decline commences at the 960 mRL and extends to the 824 mRL.

1.13 Services The Costerfield Operation purchases electricity directly from the main national electricity grid and has connections at the Brunswick Processing Plant, Cuffley Mine and Augusta Mine.

The Costerfield Operation has an existing arrangement for High Voltage electrical supply of 2,222 kVA shared between the Augusta Mine and Cuffley Mine and a 1,000 kVA supply at the Brunswick Processing Plant.

Powercor owns and manages the central Victorian electrical distribution infrastructure. ERM Business Energy is the current electrical retailer.

A 1 MVA 4.5 V/ 11 kV step-up transformer is located on the Cuffley site to allow emergency generator backup.

1.14 Contracts The antimony-gold concentrate produced from the Costerfield mine is sold directly to smelter(s) capable of recovering both the gold and antimony from the concentrates, such that Mandalay receives payment based on the concentration of the antimony and gold within the concentrate. The terms and conditions of commercial sale are not disclosed pursuant to confidentiality requirements. The marketing of the concentrate is conducted through the agency of Penfold Marketing Pty Ltd.



1.15 Environmental Studies, Permitting and Social Impacts The environmental and community impacts associated with any proposed expansion of the current Costerfield Operation are assessed with the aim of defining permitting and approvals required and evaluating whether risks to the project can be appropriately managed.

Establishing a sustainable means for disposing of groundwater from the mine is a critical aspect that requires significant investment and needs to be to be carefully managed. Existing controls in relation to noise, air quality, blast vibration, waste rock and groundwater are expected to be appropriate but will require ongoing focus.

Mandalay has implemented a Community Engagement Plan which describes processes and strategies to manage community expectations and provide transparent information to keep stakeholders informed. This plan is considered an appropriate framework to manage any community concerns associated with the mine’s expansion and to foster ongoing support for the operation.

The Department of Economic Development, Jobs, Transport and Resources (DEDJTR) facilitates this approval process and will engage with relevant referral authorities, as required. The DEDJTR may prescribe certain conditions on the approval, which may include amendments to the environmental monitoring program. The Work Plan approval process involves a thorough consultation process with the regulatory authorities, and any conditions or proposed amendments requested to the Work Plan Variation (WPV) are generally negotiated to the satisfaction of both parties.

1.16 Capital and Operating Costs The costs for the project, as described in the following section, have been derived from a variety of sources, including:

• Historical production from the Costerfield Operation

• Development costs, which have dropped due to a change in mining method to take place in 2017. Costs have been estimated from a mining method trial that took place in 2016, allowing for reduced labour costs

• Manufacturers and suppliers

• First principle calculations (based on historical production values)

• Costs include allowances for power, consumables and maintenance.

All cost estimates are provided in 2017 Australian dollars (AUD) and are to a level of accuracy of ±10%. Escalation, taxes, import duties and customs fees have been excluded from the cost estimates.

For reporting purposes, summary tables provide estimates in AUD and US dollars (USD). The USD values have been estimated using an AUD:USD exchange rate of 0.75. The total capital required for the Costerfield Operation, by calendar year (CY) is summarised in Table ES-3.



Table ES-3: Costerfield Operation – Capital Cost Estimate

Item Total (AUD M)

CY17 (AUD M)

CY18 (AUD M)

CY19 (AUD M)

CY20 (AUD M)

CY21 (AUD M)

Plant 7.3 1.8 1.2 2.5 1 0.7

Admin 0.3 0.1 0.1 0.0 0.1 0.0

Environmental 6.0 2.7 1.0 1.0 0.8 0.4

OH&S 0.2 0.1 0.0 0.0 0.1 0.0

Exploration 0.1 0.1 0.0 0.0 0.0 0.0

Mining 7.4 3.9 1.1 2.4 0.0 0.0

Total - Plant and Equipment 21.2 8.7 3.5 5.9 1.9 1.2

Capital Development 2.3 2.3 0.0 0.0 0.0 0.0

Total 23.5 11 3.5 5.9 1.9 1.2

Note: Totals may not add up due to rounding.

The Mineral Reserve operating costs include both direct and indirect costs. Direct operating costs include ore drive development and stope production activities. All costs not directly related to mine construction, development and production activities have been included in the indirect operating costs. The operating cost inputs for the Costerfield Operation are summarised in Table ES-4.

Table ES-4: Costerfield Operation – Operating Cost Estimate

Description Operating Cost (per t milled)

AUD M AUD/t USD M USD/t

Mining 108 171 81 129

Processing 30 48 23 36

Site Services, General and Administration 47 74 35 56

Total 185 293 139 221

Note: Million dollars rounded to nearest million; “/t” rounded to nearest dollar.

1.17 Offsite Costs Mandalay utilises a third party company to arrange the sale and transport of concentrate from the Brunswick Processing Plant to the smelter in China. Mandalay’s portion of the selling expenses is calculated from historical costs and comprises road transport from the Brunswick Processing Plant to the Port of Melbourne, ship transportation from Melbourne to China, shipment documentation, freight administration and assay exchange / returns.

Mandalay pays royalties to the State Government of Victoria for antimony production as well as compensation agreement liabilities.

Royalties payable include a 2.75% royalty on antimony production less any selling expenses, which is dependent on metal prices and exchange rates at the time of sale.

1.18 Economic Analysis SRK has verified the economic viability of the Mineral Resource via cashflow modelling using the inputs discussed in this report.



1.19 Adjacent Properties The Costerfield Operation Mining Lease (MIN 4644) is completely enveloped by exploration leases held by Mandalay Resources Costerfield Operations Pty Ltd. In the immediate area of the Augusta Mine there are no advanced projects, and there are no other Augusta-style antimony-gold operations in production in the Costerfield district. Exploration on adjacent prospects (EL 3316, EL 5490 and EL 5406) is at an early stage and is not relevant for further discussion in relation to this Technical Report. Three other tenements are under application and waiting approval (EL 5546, EL 06280 and EL 5548).

1.20 Other Relevant Data and Information Additional information that is deemed relevant to ensure this Technical Report is understandable and not misleading is as follows.

1.21 Remnant Mining Remnant mineralisation exists above the 1,049 mRL, primarily in E Lode and to a lesser extent in W Lode. There is an opportunity to recover this mineralisation and realise its value; however, at this stage of assessment, sound methodologies to identify the hazards associated with remnant mining to ensure higher levels of safety and improved extraction efficiency have not yet been determined. For this reason, only areas of high confidence, to be extracted with our current mining methods, have been included in this report.

1.22 Interpretation and Conclusions It is the Qualifies Person’s (QP’s) opinion that the positive results of the financial analysis demonstrate Mandalay’s ability to maintain and manage the economics of the project. The work has been undertaken at an appropriate level to support the release of the Mineral Resource and Ore Reserve estimates.

Reconciliation results show good precision and reasonable accuracy between the resource block model data and the processing plant data. Unquantified errors such as stockpiling, ore-waste misallocation, and unplanned dilution influenced the reconciliation data. Over the period, the ounces of gold predicted by the model were 9% higher than produced by the plant. The tonnes of antimony predicted by the model were 1% lower than produced by the plant. Minor adjustments to the Resource estimation process for gold grades have been made, to attempt to improve the reconciliation for 2017. SRK makes the following observations:

• Inferred resources have not been included in the economic evaluation.

• There has been a history of conversion of Inferred to Indicated Resources resulting in additional Resources from outside the Mineral Reserve being included into the LOM plans, having the potential to improve the project economics.

• Mandalay has demonstrated an ability to improve the mining method and productivity, and thus mine designs and planning, based on improved geological information.

1.23 Recommendations The Costerfield Property is an advanced property and Mandalay has a history of successful exploration and mining on the Property. SRK has observed that the degree of technical competency evident in the work performed by Mandalay geologists is high, particularly in the structural analysis of the local geology. Therefore, there is no requirement for additional work programs over and above the existing operational plans.



SRK recommends that ongoing exploration and resource definition drilling at Costerfield and Brunswick, as well as near mine drilling, continue as planned to target ongoing Resource extension. SRK recommends that Mandalay continually review the cut-off grade strategy to balance cashflow, net present value (NPV) and mine life, and that the assessment of the Brunswick deposit continues.


FAIR/EBBE/swee PLI022_Costerfield_Operations_NI_43 101_Dec 2016_Rev1 17 March 2017

2 Introduction SRK Consulting (Australasia) Pty Ltd (SRK) has worked in conjunction with Mandalay Resources Costerfield Operations Pty Ltd (Mandalay) to prepare the Costerfield Operation NI 43-101 Technical Report. The report validates the viability of mining and processing mineralisation from continuing operations at the Costerfield Operation. The Costerfield Operation is located within the Costerfield mining district, approximately 10 km northeast of the town of Heathcote, Victoria. The Augusta Mine has been operational since 2006 and was the sole ore source for the Brunswick Processing Plant until December 2013, when ore production started from the Cuffley Deposit. The Cuffley deposit is located approximately 500 m to the north of the Augusta mine workings, and is being mined in conjunction with ore from the Augusta deposit to extend the current mine life of the Costerfield Operation.

The Costerfield Operation is contained within Mining Lease MIN 4644 and comprises the following:

• Underground mine (Augusta) with production from the Augusta and Cuffley Lodes

• Conventional flotation processing plant (Brunswick) with a current name plate capacity of approximately 130,000 t/year of feed

• Mine and mill infrastructure including office buildings, workshops, core shed and equipment.

Mandalay is a publicly traded company listed on the Toronto Stock Exchange (TSX) and trading under the symbol MND, with its head office at 76 Richmond Street East, Suite 330, Toronto, Ontario, Canada M5C 1P1. On 1 December 2009, Mandalay completed the acquisition of AGD Mining Pty Ltd (AGD) from Cambrian Mining Limited (Cambrian), a wholly-owned subsidiary of Western Canadian Coal Corporation (WCC), resulting in AGD becoming a wholly-owned subsidiary of Mandalay.

Units of measurement used in this report conform to the SI (metric) system as illustrated in the List of Abbreviations.

2.1 Scope of Work The Scope of Work, as defined in a letter of engagement executed during November 2016 and subsequent discussions between Mandalay and SRK, includes the review of documents provided by Mandalay and preparing documentation as required for the NI 43-101 Technical Report on the Costerfield Operation, in compliance with NI 43-101 and Form 43-101 F1 guidelines.

This work involved the review of the following aspects of this project:

• Supervise and review the Mineral Resource Estimate (MRE) and update for Augusta and Cuffley deposits prepared by Mandalay

• Review the Brunswick, Sub King Cobra and Cuffley Deeps MREs

• Prepare and review the mine design and mining schedules for Augusta and Cuffley Lodes

• Review and comment on the project infrastructure aspects

• Review and comment on the metallurgical and processing aspects

• Review the environmental considerations

• Review the capital and operating cost estimates

• Review the financial modelling

• Make recommendations for additional work.



2.2 Work Program The Costerfield NI 43-101 Technical Report herein is a collaborative effort between Mandalay and SRK. SRK has independently reviewed the work completed by Mandalay.

The Mineral Resource and Mineral Reserve Statement reported herein was prepared in conformity with generally accepted Canadian Institute of Mining (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.

The Technical Report was assembled in Melbourne during the months of January to March 2017.

2.3 Basis of Technical Report This report is based on information provided by Mandalay to SRK and verified during site visits conducted between 2012 and 2016, and additional information provided by Mandalay throughout the course of SRK’s investigations. SRK has no reason to doubt the reliability of the information provided by Mandalay. The most recent site visit was undertaken on 18 November 2016.

2.4 Qualifications of SRK and SRK Team The SRK Group comprises over 1,500 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 its ownership rests solely with its staff. This 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 the lead author Peter Fairfield, Principal Consultant (Project Evaluation), BEng (Mining), FAusIMM (No 106754), CP (Mining). By virtue of his education, membership of a recognised professional association and relevant work experience, Peter Fairfield is an independent Qualified Person (QP) as this term is defined by NI 43-101.

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

Simon Walsh, SRK Associate Principal Metallurgist, BSc (Extractive Metallurgy & Chemistry), MBA Hons, CP, GAICD, undertook a review of the metallurgical, mineral processing and infrastructure aspects of the project. By virtue of his education, membership of a recognised professional association and relevant work experience, Simon Walsh is an independent QP as this term is defined by NI 43-101.

Anne-Marie Ebbels, Principal Consultant (Mining), BEng (Mining), MAusIMM (No 111006), CP (Mining) conducted a peer review of the non-geological aspects of this Technical Report. Anne-Marie Ebbels is, by virtue of her education, membership of a recognised professional association and relevant work experience, an independent QP as this term is defined by NI 43-101.



Table 2-1 lists the individuals who, by virtue of their education, experience and professional association, are considered QPs as defined in NI 43-101 for this report. The table defines the areas of responsibility for the QPs, who all meet the requirements of independence as defined in NI 43-101.

2.5 Acknowledgement SRK would like to acknowledge the support and collaboration provided by Mandalay personnel for this assignment. Their collaboration was greatly appreciated and instrumental to the success of this project. The majority of the resource estimation and mine planning work was completed by Mandalay personnel with supervision and review by SRK. The Mandalay personnel with significant roles were: Chris Davies and Cael Gniel, who undertook the geological modelling and resource estimation work; Ross Laity and Chloe Cavill, who were involved in the reserves, scheduling and mine design work; Damon Buchanan, who undertook the mineral processing and metallurgical testwork; and April Westcott for the geological technical report writing and compilation of the report.

2.6 Declaration SRK’s opinion contained herein, effective 31 December 2016, 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 errors occur, SRK does not consider them to be material.

SRK is not an insider, associate or affiliate of Mandalay, and neither SRK nor any affiliate has acted as advisor to Mandalay, 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.



Table 2-1: List of Qualified Persons

QP Position Employer Last site visit date Professional designations Area of responsibility and Report sections

Peter Fairfield Principal Consultant (Project Evaluations) SRK 1 September 2016

BEng (Mining), FAusIMM, CP (Mining)

Sections 1 - 6, Sections 15 - 16, Sections 18 - 27

Anne-Marie Ebbels Principal Consultant (Mining) SRK 30 - 31 August 2012

BEng (Mining), GDip (Computer Studies), MAusIMM(CP)

SRK Peer Review

Danny Kentwell Principal Consultant (Resource Evaluation) SRK 29 November 2016

MSc Mathematics & Planning (Geostatistics), FAusIMM

Sections 7 - 12, Section 14

Simon Walsh Principal Metallurgist and Director

Simulus Engineers (SRK Associate) 1 September 2015

BSc (Extractive Metallurgy), MBA Hons, CPAusIMM, GAICD

Metallurgy, Processing and Infrastructure: Sections 13, 17 and 18

.



3 Reliance on Other Experts 3.1 Marketing

Marketing information for this report, specifically Section 19, relies entirely on information provided by Roskill Information Services Ltd.

A specific marketing study was not completed for this report.



4 Property Description and Location 4.1 Property Location

The Costerfield Operation is located within the Costerfield mining district of Central Victoria, approximately 10 km northeast of the town of Heathcote and 50 km east of the City of Bendigo, as shown in Figure 4-1.

The Costerfield Operation encompasses the underground Augusta Mine (Augusta) including the Cuffley Deposit, the Brunswick Processing Plant, Splitters Creek Evaporation Facility, Bombay Tailings Storage Facility (TSF) and associated infrastructure.

The Augusta Mine is located at latitude 36° 52’ 27” south and longitude 144° 47’ 38” east. The Brunswick Processing Plant is located approximately 2 km northwest of Augusta. The Cuffley Deposit is located approximately 500 m north-northwest of the Augusta workings and is accessed via decline from Augusta.

Figure 4-1: Costerfield Operation Location Source: Encom Discover Mapinfo Pro.



4.2 Land Tenure Tenure information for the two Mining Licences and five Exploration Licences is shown in Table 4-1.

Table 4-1: Granted Tenement Details

Tenement Name Status Company Area Grant date Expiry date

MIN4644 Costerfield Granted AGD Operations P/L 1219.3 Ha 25/02/1986

30/06/2016 Renewal lodged

EL3310 Costerfield Granted AGD Operations P/L

59.0 GRATS 17/09/1993 17/09/2017

EL5519 Antimony Creek Application

Mandalay Resources Costerfield

Operations Pty Ltd

11.0 GRATS 28/05/2015 27/05/2018

EL5432 Peels Track Granted AGD Operations P/L

40.0 GRATS 23/08/2012 22/08/2017

MIN5567 Splitters Creek Granted

Mandalay Resources Costerfield

Operations Pty Ltd

30 Ha 20/02/2013 20/02/2023

Note: 1 GRATS is equivalent to 1 km2.

Mandalay manages the Costerfield Operation and holds a 100% interest in tenements MIN 4644, MIN 5567, EL 3310, EL 5432, and EL5519 as shown in Figure 4-2 and Figure 4-3.



Figure 4-2: Plan of Area – Mining Licence No. 4644 Source: DEDJTR, Victorian State Government, 2009.



Figure 4-3: Current Mandalay Resources Mining and Exploration Lease boundaries Source: Mandalay Resources, 2016.



4.3 Underlying Agreements The sustainable and responsible development of Mineral Resources in Victoria is regulated by the State Government of Victoria through the Mineral Resources (Sustainable Development) Act 1990 (MRSD Act).

The MRSD Act, which is administered by the Department of Economic Development, Jobs, Transport and Resources (DEDJTR), requires that negotiation of access and/or compensation agreements with landowners affected by the work plans is undertaken between the mining tenement applicant and the relevant landowner prior to a Mining Licence being granted or renewed. In accordance with this obligation, Mandalay has compensation agreements in place for land allotments that are situated within the boundaries of MIN 4644 and owned by third parties.

Mandalay owns the land that contains MIN 5567 and, as such, no compensation agreements are required.

4.4 Environmental Liability The rehabilitation bond for MIN 4644 is currently AUD 3.106 M. The latest review was at the end of 2014, which was followed by a consultation period with community members in January 2015. The rehabilitation bond is reviewed by DEDJTR every two years or when a substantive Variation to the Work Plan is approved. The bond review for MIN 4644 has not commenced, but there have been no material changes to the Augusta mine site and processing plant in the last two years.

There is a further AUD10,000 bond paid to the DEDJTR for tenements EL 3310 and EL 5432 and Vic Roads licenses for pipelines that cross roads.

The rehabilitation bond for MIN 5567, the lease on which the Splitters Creek Evaporation Facility has been constructed, was calculated in October 2014 as AUD748,000.

The total bond for MIN 4464, the lease where the Augusta mine site and Brunswick Processing Plant are situated, is AUD3.106 M.

Rehabilitation is undertaken progressively at the Costerfield Operation, with the environmental bond only being reduced when rehabilitation of an area or site has been deemed successful by the DEDJTR. This rehabilitation bond is based on the assumption that all rehabilitation is undertaken by an independent third party. Therefore, various project management and equipment mobilisation costs are incorporated into the rehabilitation bond liability calculation. In practice, rehabilitation costs may be less if Mandalay chooses to utilise internal resources to complete rehabilitation.

Other than the rehabilitation bond, the project is not subject to any other environmental liabilities.

4.5 Royalties Royalties apply to the production of antimony and are payable to the Victorian State Government through the DEDJTR. This royalty is applied at 2.75% of the revenue realised from the sale of antimony produced, less the selling costs.

No royalty is payable on gold production, nor is there any royalty agreement in place with previous owners.

Royalties are payable to the Victorian State Government through the DEDJTR if waste rock or tailings is sold (or provided to) to third parties, because these are deemed to be ‘quarry products’. The royalty rate is AUD0.87/t.



4.6 Taxes Mandalay reports that, as at December 2016, there are no tax losses to be carried forward.

Income Tax on Australian company profits is set at 30%.

4.7 Legislation and Permitting Mining Licence MIN 4644 has a series of licence conditions that must be met and are the controlling conditions upon which all associated Work Plan Variations (WPVs) are filed with the regulatory authority. Apart from the primary mining legislation, which consists of the MRSD Act, operations on MIN 4644 are subject to the additional following legislation and regulations for which all appropriate permits and approvals have been obtained.

Legislation:

• Environment Protection Act 1970

• Planning and Environment Act 1987

• Environmental Protection and Biodiversity Conservation Act 1999

• 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

• Drugs, Poisons and Controlled Substances Act 1981

• Public Health and Wellbeing Act 2008

• Water Act 1989

• Crown Land (Reserves) Act 1978

• Radiation Act 2005

• Conservation, Forests and Lands Act 1987

• Wildlife Act 1975.

Regulations:

• Dangerous Goods (Explosives) Regulations 2011

• Dangerous Goods (Storage and Handling) Regulations 2000

• Dangerous Goods (HCDG) Regulations 2005

• Drugs, Poisons and Controlled Substances (Commonwealth Standard) Regulations 2011

• Mineral Resources Development Regulations 2002.

Mandalay is currently operating under an approved Work Plan in accordance with Section 39 of the MRSD Act. WPVs are required when significant changes from the Work Plan exist and it is deemed that the works will have a material impact on the environment and/or community. Various WPVs have been approved by the DEDJTR and are registered against the tenement.

In the process of renewing MIN 4644, Mandalay has submitted an application to renew Mining Licence MIN 4644 for a period of 10 years as opposed to the current renewal tenure of 2 years. In accordance with the Native Title Act 1993 (NTA), native title requirements for the renewal application for MIN 4644 have now been determined. A future act assessment has been completed and the Right to Negotiate



(RTN) process will be undertaken to address the native title requirements. The RTN process involves notification followed by negotiation with the registered native title claimants (if they exist) to reach a Section 31 agreement (a pro forma template exists to assist resource companies and Indigenous communities in negotiations. The template outlines terms and conditions that have already been agreed to by the Victorian Minerals and Energy Council and Native Title services Victoria.)

To commence the process, the DEDJTR will contact the Native Title Tribunal to search relevant registers to identify any existing native title claimants. Notification advertising for potential claimants runs for three months (until 9 June 2017) to become native title parties in relation to the renewal of the licence.

To the best of the author’s knowledge, there is no other significant factor or risk that may affect access, title, or the right or ability to perform work on the property.



5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Accessibility Access to the Costerfield Operation is via the sealed Heathcote–Nagambie Road, which is accessed off the Northern Highway to the south of Heathcote. The Northern Highway links central and north-central Victoria with Melbourne.

The Augusta Mine site is accessed off Heathcote–Nagambie Road via McNicols Lane, which is a sealed/gravel road that continues for approximately 1.5 km to the Augusta site offices.

The Brunswick Processing Plant is located on the western side of Heathcote–Nagambie Road, approximately one kilometre north of the McNicols Lane turnoff. The Brunswick site offices are accessed by a gravel road that is approximately 600 m long.

5.2 Land Use Land use surrounding the site is mainly small scale farming consisting of grazing on cleared land, surrounded by areas of lightly timbered Box-Ironbark forest. The majority of the undulating land and alluvial flats is privately held freehold land.

The surrounding forest is largely rocky, rugged hill country administered by DEDJTR as State Forest. The Puckapunyal Military Area is located on the eastern boundary of the Project Area.

The Augusta Mine site is located on privately held land, while the Brunswick Processing Plant is located on unrestricted Crown land.

The Cuffley Deposit, accessed via the Augusta Mine, is located beneath unrestricted Crown land that consists of sparse woodland, with numerous abandoned shafts and workings along the Historical Alison and New Alison mineralised zone.

5.3 Topography The topography of the Costerfield area consists of relatively flat to undulating terrain with elevated areas to the south and west sloping down to a relatively flat plain to the north and east. The low lying areas of the plain are a floodplain. The area ranges in elevation from about 160 metres above sea level (mASL) in the east along Wappentake Creek, to 288 mASL in the northwest.

5.4 Climate The climate of central Victoria in which the Costerfield Operation is located is ‘Mediterranean’ in nature and consists of hot, dry summers followed by cool and wet winters. Annual rainfall in the area is approximately 500 to 600 mm, with the majority occurring between April and October. The annual pan evaporation is between 1,300 and 1,400 mm.

The temperature ranges from -2°C in winter (May to August) to +40°C in summer (November to February). Monthly average temperature and rainfall data from Redesdale (the nearest weather recording station to Costerfield, which is some 19 km southwest of Heathcote) is shown in Figure 5-1.

The weather is amendable to year-round mining operations; however, occasional significant high rainfall events may restrict surface construction activity for a small number of days.



Figure 5-1: Monthly average temperature and rainfall Source: Bureau of Meteorology.

5.5 Infrastructure and Local Resources The nearest significant population to Costerfield is Bendigo, which is located 50 km to the west-northwest and has a population of approximately 100,000. The Costerfield Operation is a residential operation, with personnel residing throughout central Victoria as well as Melbourne. Local infrastructure and services are available in Heathcote.

The Augusta Mine site consists of a bunded area that includes site offices, an underground portal, workshop facilities, a waste rock storage area, settling ponds, a mine dam, change house facilities and a laydown area. Augusta has operated as an underground mine since the commencement of operations in 2006. The Cuffley operations use the infrastructure associated with the current Augusta operations. The Augusta Mine box-cut, portal and workshop are shown in Figure 5-2.

Figure 5-2: Augusta box-cut, portal and workshop



The Brunswick complex consists of a processing plant, run-of-mine (ROM) pad, site offices and Brunswick Open Pit, as shown in Figure 5-3. The Brunswick Processing Plant consists of a 150,000 tonnes per annum (tpa) gravity-flotation gold-antimony processing plant, with workshop facilities, site offices, TSFs, a core shed and a core farm. It produces an antimony-gold concentrate which is trucked to the Port of Melbourne, 130 km to the south, and transferred onto ships for export to foreign smelters.

Figure 5-3: Aerial view of Brunswick Processing Plant and Brunswick Open Pit

Electrical power for both operations is primarily sourced from the main national electricity grid, with a small component of diesel generation positioned at the Brunswick site primarily for operation of the Reverse Osmosis (RO) Plant.

The Costerfield Operation has an existing arrangement for supply of 2,000 kW 22 kV at the Augusta Mine and approximately 1,000 kVA 240/415 V at the Brunswick Processing Plant.

Process water for the Brunswick Processing Plant is drawn from the brine stream of the RO Plant and is supplemented by brine currently in storage when the RO Plant is not in operation. The Augusta Mine reuses groundwater that has been dewatered from the underground workings. Potable water is trucked in from Heathcote, while grey water is stored in tanks and sewage is captured in sewage tanks before being trucked offsite by a local contractor.

The Splitters Creek Evaporation Facility evaporates groundwater extracted from the Operations, thereby maintaining dewatering rates from the underground workings. Additional detail is provided in Section 20.

Construction of an additional 2.5 m raise for the Brunswick TSF commenced in November 2016. This raise will provide an additional 150,000 m3 of storage capacity and is expected to be completed in May 2017. Additional detail is provided in Section 20.



6 History 6.1 Introduction

From the 1860s, beginning with the initial discovery of the Costerfield Reef, until 1953, several companies developed and mined antimony deposits within the Costerfield area. Some underground diamond drilling is known to have occurred between 1934 and 1939, when Gold Exploration and Finance Company of Australia operated the Costerfield Mine, but details of these holes are scarce and poorly recorded.

Significant exploration of the Costerfield area using modern exploration techniques did not occur until 1966.

6.2 Ownership and Exploration Work This section describes the work carried out by different owners over time. Table 6-1 presents a summary of the historical drilling statistics for each company at Costerfield since 1966.

Table 6-1: Historical Drilling Statistics for the Costerfield Property

Company Year No. of Holes

Metres (Diamond)

Metres (Percussion / Auger)

Surface Drilling

Mid-East Minerals 1966 - 1971 33 3,676.2

Metals Investment Holdings 1971 12 1,760.8

Victoria Mines Department 1975 - 1981 32 3,213.0

Federation Resources NL 1983 - 2000 27 2,398.3

AGD / Planet Resources JV 1987 - 1988 23 1,349.2

AGD NL

1987 - 1988 14 1,680.8

1994 - 1995 142 1,368.5 5,536.0

1996 59 195.5 2,310.0

1997 23 725.0

AGD Operations

2001 27 3,361.1

2002 7 907.5

2003 30 1,522.0

2004 27 3,159.9

2005 31 4,793.4

2006 - 2007 67 4,763.4

2007 - 2008 11 2,207.2

2008 - 2009 8 1,785.8

Subtotal Surface 573 32,714.4 13,999.3

Underground Drilling

AGD Operations 2008 - 2009 11 799.8

Mid-East Minerals (1968 - 1971) From 1968 to 1969, the price of antimony rapidly rose from USD0.45/lb to USD1.70/lb. This encouraged Mid-East Minerals (MEM) to acquire large amounts of ground around Costerfield.



Between 1969 and 1971, MEM conducted large-scale geochemical, geophysical and diamond drilling programs. These were conducted across the south Costerfield area encompassing the Alison Mine and south towards Margaret’s, encompassing both the Cuffley Lode and the Augusta Mine areas. Diamond drilling for MEM was most successful at the Brunswick Mine. Dropping antimony prices in 1971 caused MEM to abandon its projects.

Metals Investment Holdings (1971) A series of diamond drill holes was completed by Metals Investment Holdings in 1971. Most drilling occurred to the north of the Alison Mine, with the exact locations of the holes unknown. Two drill holes were situated to the north of the Tait’s Mine (north of Augusta), but minimal detail on the Tait’s Mine remains.

Victorian Mines Department (1975 - 1981) A series of diamond drill holes was completed by the Victorian Mines Department in the late 1970s. Most drilling occurred to the south of the Brunswick Mine. However, two holes (M31 and M32), drilled approximately 150 m to the south of the South Costerfield Shaft (in the Augusta mine area), intersected a high-grade reef. This reef was interpreted as the East Reef, which was mined as part of the South Costerfield Mine.

Federation Resources NL (1983 - 2000) Federation Resources NL undertook several campaigns of exploration in the Costerfield area, but focused on the Browns-Robinsons prospects to the east of the Alison Mine. The exploration identified a gold target with no evidence of antimony. This target has yet to be followed up by Mandalay because it is viewed as low-priority target.

Federation Resources NL conducted desktop studies on the area above the Augusta Mine, noting the anomalous results of the soil geochemistry programs conducted by the Victorian Mines Department and MEM, but did not conduct drilling at this location.

Australian Gold Development NL / Planet Resources JV (1987 - 1988) Australian Gold Development NL conducted a short reverse circulation (RC) drilling program in conjunction with joint venture (JV) partner, Planet Resources in 1987. This drilling consisted of a total of 21 holes for 1,235 m across the broader Costerfield area. Gold was assayed via atomic absorption spectrometry (AAS), which compromised antimony grades. Drilling was also done with a tri-cone bit, which could have led to a high degree of contamination.

Australian Gold Development NL (1987 - 1997) From 1987 to 1997, Australian Gold Development NL undertook several programs of exploration and mining activities predominantly focused around the Brunswick Mine. A series of RC holes was drilled during 1997 to test for shallow oxide gold potential to the north of the Alison Mine. Several occurrences of yellow antimony sulphides were noted, but were not followed up.

AGD Operations Pty Ltd (2001 - 2009) In 2001, AGD Operations Pty Ltd (AGD – formerly Australian Gold Development) and Deepgreen Minerals Corporation Ltd entered into an agreement to form a JV to explore the Costerfield tenements. The agreed starting target was the MH Zone, now known as the Augusta Mine.



2001 AGD’s drilling in the MH Zone commenced on 5 April 2001. In total, 27 holes were completed for 3,301.1 m. All holes were drilled with an initial PQ or HQ collar to approximately 25 m depth and then finished with an NQ-sized drill bit, to maximise core recoveries. Triple-tube drilling was also employed in some areas to maximise recoveries. Cobar Drilling Company Pty Ltd, based in Rushworth, was contracted for the drilling program. Less competent rock adjacent to the mineralisation was successfully recovered during this program, but core loss was still estimated to be up to 15% within the mineralised zones. All holes were downhole surveyed and orientated during drilling. Collar locations were surveyed by Cummins & Associates from Bendigo.

This drilling was confined to an area 180 m south of the South Costerfield Shaft and over approximately 400 m of strike.

Because of prolonged mining and exploration undertaken in the Costerfield area, up to three metric grids were in use. The drilling undertaken in 2001 at Augusta was based on the mine grid established in the late 1950s. This grid set-up remains in use in present-day mining and exploration activities.

2002 In 2002, AGD completed a further five holes in the MH Zone for a total 732.3 m, including 41.7 m blade, 309.3 m of RC hammer and 381.3 m of HQ diamond drilling. Drill hole MH034 intersected an unmineralised zone at 55 m downhole. This is hypothesised to represent the Alison line of lode towards the south.

Towards the east of the MH Zone, AGD completed two lines of soil sampling comprising 400.5 m of aircore drilling in 88 holes. The known MH lodes were highly anomalous and a weak, gold-only trend was outlined 180 m east of the MH Zone. This zone was drilled by diamond drill hole MH028; it comprised a large siliceous lode zone with low-grade gold values.

To the south of the MH Zone, AGD sampled two soil lines in 42 holes. It was later recognised that these holes probably did not sample basement siltstones. A further line of 21 soil holes confirmed this prognosis. These holes picked up widespread anomalous gold geochemistry with a central strong anomaly. A total of 218 m of aircore drilling was completed.

2003 In 2003, the MH Zone was renamed the Augusta Deposit. In total, 30 diamond drill holes, for 1,514 m, were drilled by AGD as part of an infill and extension program for the Augusta Deposit. The main purpose of this drilling was to prove continuity of the deposit to near surface in preparation for open pit mining, and to extend the mineralised system both north and south. Mineralisation was shown to extend north to the South Costerfield Shaft and upwards to the surface. To the south, drilling confirmed that the lode system, although present, was not economic.

Each hole was logged in detail and geological lode thickness and recovered thicknesses were recorded. Core loss was estimated to be less in this drill program than previous drilling programs, even though most of the drill holes were drilled in the weathered zone.

In addition to the infill and extension program, 14 RC drill holes were drilled as part of a metallurgical testwork program. These holes were drilled at low angles to the lodes, specifically to obtain the required sample mass for the metallurgical testwork.

2004 / 2005 Between October 2004 and April 2005, AGD completed a 26 hole diamond drilling program at the Augusta Deposit. Apart from 5 m percussion pre-collars and four RC geotechnical holes, the holes were drilled by HQ triple-tube diamond drilling.



The objectives of the diamond drilling program were:

• Improvement in mineralisation definition by increasing drillhole density

• Extension of the mineralisation model by drilling around the deposit periphery

• Increasing the Mineral Resource and Mineral Reserve.

2006 / 2007 AGD’s drilling activities throughout 2006 and 2007 comprised grid drilling of the Brunswick Deposit and drilling of the periphery of the Augusta Deposit for a total of 7,562 m of diamond drilling. This comprised the following drill holes:

• 31 holes, totalling 4,994 m, drilled under the old Brunswick open pit for resource estimation

• 17 holes, totalling 755 m, drilled into the upper northern end of W Lode

• 20 holes, totalling 1,813 m, drilled north of the Augusta Mine to test the northern extent of the E Lode.

The Brunswick resource definition drilling was drilled using HQ triple-tube with a modified Longyear LM75 drill rig by Boart Longyear drilling. The area under the pit was drilled on a 40 x 40 m pattern.

Due to initial difficulty in following the W Lode underground, a bobcat-mounted Longyear LM30 diamond drilling rig was used to infill drill the near-surface portion of W Lode. This drilling was done using a narrow-kerf LTK60-sized core barrel, with a total of 17 holes, totalling 755 m, drilled adjacent to the Augusta box-cut.

On completion of the Brunswick and W Lode drilling, both the LM75 and LM30 rigs were used to drill north of the Augusta Mine, tracing the northern extent of E Lode towards the old South Costerfield workings. A total of 20 holes for 1,813 m were drilled north of the Augusta Mine.

Development of the Augusta decline commenced during the first quarter of 2006. By the end of the second quarter, all the surface infrastructure had been completed, together with open cut mining of the E and C lodes. Decline development commenced during June 2006 and underground production began by the end of the third quarter of 2006.

2007 / 2008 AGD’s drilling activities throughout 2007 and 2008 comprised reconnaissance drilling of the Tin Pot Gully Prospect and drilling along strike and down-dip of the existing Augusta Deposit. A total of 3,395.6 m of diamond drilling was carried out during the year. This comprised the following:

• 13 holes, totalling 1,188 m, drilled under the Tin Pot Gully Prospect

• 11 holes, totalling 2,207 m, drilled into the Augusta Deposit, particularly to test W and E lodes.

Encouraging results highlighted down-dip and strike extensions in terms of vein widths and grades, as described below:

• W Lode: 8 out of the 11 drill holes confirmed W Lode continuity down-dip, with true thicknesses ranging from 0.254 m to 0.814 m at 22.50 - 89.26 g/t gold and 16.19% - 47.20% antimony

• E Lode: 3 out of the 8 drill holes confirmed E Lode continuity down-dip, with true thicknesses ranging from 0.074 m to 0.215 m at 4.24 - 35.1 g/t gold and 3.25% - 32.2% antimony

• N Lode: 6 holes out of the 11 holes intercepted N Lode or a similar structure in the hanging wall of W lode, showing true thicknesses from 0.09 m to 0.293 m at 6.82 - 46.9 g/t gold and 6.81% to 27% antimony.



Based on these results, AGD commissioned AMC to undertake a resource estimate for the Augusta Deposit in January 2008.

Between February and June 2008, Silver City Drilling Company completed 11 drill holes, totalling 2,207.2 m, which were drilled on the northern section of the Augusta Deposit, from 4,411 mN to 4,602 mN.

The 11 surface drill holes covered an area of approximately 18,740 m2, delineating a 120 m down-dip continuation, below 4 level, of the three dominant Augusta Lodes: W Lode, E Lode and N Lode.

Holes ranged in size from HQ to NQ and LTK46.

By June 2008, capital development reached 7 level (1,081 mRL). Development was completed on E Lode on 5 level and was halfway through completion on 6 level. W Lode development was completed down to 4 level and development on 5 level was just beginning. Handheld / airleg rise mining had begun.

2008 / 2009 AGD’s drilling activities throughout 2008 and 2009 comprised drilling along strike and down-dip from the existing Augusta resource. A total of 2,585.95 m of diamond drilling was completed.

Drilling during 2008 and 2009 was concentrated on the definition of the W Lode resource. Five drill holes tested the depth extent of the W Lode. Another 13 holes were designed as infill holes to test ore shoots and gather geotechnical data. Holes ranged in size from HQ to NQ and LTK46.

By June 2009, capital development reached 9 level (1,070 mRL). Ore drive development was constrained to levels along E and W lodes. Stoping along W Lode was conducted and additional development along E Lode used three mining methods: floor benching, cut and fill and longhole stoping.

2009 / 2010

On 1 December 2009, AGD Mining Pty Ltd / AGD Operations Pty Ltd was acquired by Mandalay Resources Corporation.

Drilling from July 2009 to June 2010 consisted mainly of drilling along strike and down dip from the existing Augusta resource (MIN 4644). In total, 332.5 m of diamond coring was undertaken to target the Augusta resource.

In addition, 547 m of bedrock geochemistry (Augusta South) aircore drilling was completed within MIN 4644.

Outside of the main field, 120.5 m of diamond drilling was completed at the True Blue Reef prospect within EL 3310, and 122.8 m of diamond drilling was at Hirds Reef prospect within EL 4848.

Drilling during this reporting period concentrated on the definition of the W Lode resource. Four drill holes tested the depth extent of W Lode. Another six holes were designed as infill holes to test ore shoots and gather geotechnical data. The significant results of the Augusta Deposit drilling through this period have been tabulated below.

From July 2009 to June 2010, capital development reached 1,020 mRL (155 m below surface). Ore drive development was carried out on the E and W lodes.

2010 / 2011

Drilling was undertaken on two projects from July 2010 to June 2011: the Augusta Deeps drilling project and the Brownfields Exploration project. The Augusta Deeps project was undertaken with the view to extending the current Augusta resource to depth (MIN 4644). The objective of the Brownfields



Exploration project was to find additional ore sources within Mandalay’s tenements, to supplement the Augusta Deposit ore. The initial emphasis of the Brownfields Explortaion project was to identify sources of ore within 1 km of the Augusta Decline. In total, 9,890.7 m of diamond coring and 732 m of auger drilling were undertaken as part of the two projects.

Capital development reached 976 mRL (200 m below surface). Ore drive development was carried out on the E and W lodes.

2011 / 2012

Drilling was undertaken on four projects: the Augusta Deeps drilling project, the Alison / Cuffley drilling project, the Brownfields/ Target Testing drilling project and the Target Generation - Bedrock Geochemistry auger drilling project.

The Augusta Deeps project was undertaken with the view to extending the current Augusta resource to depth and along strike (MIN 4644). The Alison/ Cuffley Project was undertaken to outline the recently discovered Cuffley Lode and to define an initial Inferred Resource.

The objective of the Brownfields Target Exploration project was to find additional ore sources within Mandalay’s tenements, to supplement the Augusta Deposit ore. The initial emphasis of the Brownfields Project was to identify sources of ore within 1 km of the Augusta Decline. The drilling program is now more regional.

The Bedrock Geochemistry auger drilling project revealed anomalous zones under shallow alluvial/ colluvial cover throughout the tenements.

In total, 18,581.4 m of diamond coring and 7,295.6 m of auger drilling were undertaken for the four projects from July 2011 to June 2012.

On 17 June 2011, MB007 intersected the Cuffley Lode, just below a flat fault that had stopped production at 5 level in the Alison mine in 1922. Resource drilling commenced in July 2011.

The Cuffley Lode is 500 m north-northwest of the Augusta Deposit workings and scoping studies commenced in 2011 to access the deposit from the Augusta Decline.

From July 2011 to June 2012, capital development reached 926 mRL (252 m below surface) and ore drive development was carried out on the E and W lodes. 2012 / 2013

On 7 February 2013, AGD Operations Pty Ltd underwent a name change, becoming Mandalay Resources Costerfield Operations Pty Ltd.

Drilling was undertaken on two primary projects: Cuffley Resource Drilling and Augusta Resource Drilling. The goals of these projects were to:

• Expand the existing Inferred Resource of the Cuffley Lode, both along strike and at depth

• Increase confidence of the central part of the Cuffley Lode to aid mine planning and support a corporate decision to develop the Cuffley Lode

• Expand the existing Inferred Resource of the Augusta deposit, specifically targeting N Lode along strike from existing N Lode development

• Increase confidence in areas of N Lode that were easily accessible from the current mine development and those areas that could potentially be accessed from development to access the Cuffley resource.

In total, 25,774.8 m of diamond drilling and 3,838 m of auger drilling were undertaken on Mandalay’s tenements at Costerfield from July 2012 to June 2013. During this period, capital development



reached 878 mRL (300 m below surface) and ore drive development was carried out on the E, W and N lodes.

2013 / 2014

In 2013/ 2014, the focus was on finalising the Cuffley and Augusta Resource Drilling. The goals achieved included:

• Expanding the existing Inferred Resource of the Cuffley Lode, both along strike and at depth

• Increasing the confidence of the central part of the Cuffley Lode to aid mine development and stoping the Cuffley Lode

• Expanding the existing Inferred Resource of the Augusta deposit, specifically targeting N Lode along strike from the existing N Lode development

• Infilling and extending the Cuffley resource to the north and south, along with Cuffley Shallows in between the flat fault and the Adder Fault.

In total, 20,817 m of diamond drilling and 3,906 m of auger drilling were undertaken on Mandalay’s tenements at Costerfield during 2013/ 2014.

During 2014, mining took place along the Augusta and Cuffley deposits. Development on C1 and C2 lodes within the Cuffley deposit began in January 2014. Both deposits were accessed through the Augusta portal, with Cuffley capital infrastructure exiting the Augusta Decline at the 1030 level.

2014 / 2015

Exploration in 2015 was focused on extending the Cuffley and Augusta resources, both along strike and at depth. The expansion of the Cuffley resource included the commencement of drilling in the Cuffley Deeps and Sub King Cobra regions. The goals achieved included:

• Expanding the existing Inferred Resource of the Cuffley Lode along strike and defining a resource below the Cuffley Lode at depth

• Commencing drilling at depth below the Cuffley Deposit into the Cuffley Deeps and Sub King Cobra areas

• Increasing the confidence of the central part of the Cuffley Lode to aid mine development, and stoping the Cuffley Lode

• Expanding the existing Inferred Resource of the Augusta deposit, specifically targeting N lode along strike from the existing N Lode development

• Infilling and extending the Cuffley resource to the north and south, along with Cuffley Shallows in between the flat fault and the Adder Fault.

In total, 18,439 m of diamond drilling and 2,732 m of RC drilling were undertaken on Mandalay’s tenements at Costerfield during 2015.

During 2015, mining took place along the Augusta and Cuffley deposits. Development on C1 and C2 lodes within the Cuffley deposit began in January 2014. Both deposits were accessed through the Augusta portal, with Cuffley capital infrastructure exiting the Augusta Decline at the 1030 level.

6.3 Historical Resource and Reserve Estimates Mandalay reported Mineral Resources and Reserves from 2010 - 2016. The Mineral Resources are presented in Table 6-2 and Mineral Reserves are presented in Table 6-3.

These estimates have been superseded by the current resource and reserve estimates in this report.



Table 6-2: Historical Mineral Resources – Mandalay Resources – Costerfield Project

Effective Date

Au Price USD/oz

Sb Price USD/oz

AuEq Cut-off Grade (g/t)

Measured Resource Indicated Resource Inferred Resource

Tonnes (‘000)

Au (g/t)

Sb (%)

Au ounces (‘000)

Sb tonnes

Tonnes (‘000)

Au (g/t)

Sb (%)

Au ounces (‘000)

Sb tonnes

Tonnes (‘000)

Au (g/t)

Sb (%)

Au Ounces (‘000)

Sb tonnes

1/03/2010 1,000 6,000 67.2 16.9 10.0 36.4 6,749 189.6 9.6 4.6 58.4 8,683 245.7 7.8 4.2 61.5 10,202

31/12/2011 1,100 9,850 4.6 158.4 12.9 7.8 65.5 12,291 202.4 7.3 3.7 47.7 7,502 375.0 12.7 5.6 152.9 21,183

31/12/2012 1,600 12,500 4.7 167.0 8.0 4.9 42.7 8,202 367.0 10.0 3.5 117.9 12,912 610.0 7.1 3.2 139.8 19,490

31/12/2013 1,400 12,000 3.9 191.4 8.4 4.3 51.5 8,157 606.0 9.6 4.0 186.4 24,237 570.0 7.4 3.7 135.3 21,342

31/12/2014 1,400 12,000 3.8 213 9.8 4.5 67 9,600 786 6.9 3.3 175 26,300 519.0 5.3 2.6 89.0 13,700

31/12/2015 1,400 11,000 3.8 247 12.1 4.6 96 11,000 798 7.6 3.4 194 27,000 491 4.3 2.0 68.0 9,700

Table 6-3: Historical Mineral Reserves – Mandalay Resources – Costerfield Project

Effective Date

Au Price USD/oz

Sb Price USD/oz

AuEq Cut-off Grade (g/t)

Proven Reserves Probable Reserves Total Reserves

Tonnes (‘000)

Au (g/t)

Sb (%)

Au (koz)

Sb (t)

Tonnes (‘000)

Au (g/t)

Sb (%)

Au (koz)

Sb (t)

Tonnes (‘000)

Au (g/t)

Sb (%)

Au (koz)

Sb (t)

1/03/2010 1,000 6,000 20.1 16.9 9.7 10.9 1,953 45.4 11.4 5.8 16.7 2,636 65.6 13.1 7.0 27.6 4,588

31/12/2011 1,600 12,000 4.6 41.9 13.2 7.9 17.7 3,300 46.5 6.4 4.0 9.6 1,860 88.4 9.6 5.8 27.3 5,160

31/12/2012 1,600 12,500 4.7 48.1 11.0 6.5 17.0 3,128 130.0 8.1 3.2 33.9 4,161 178.2 8.9 4.1 50.9 7,289

31/12/2013 1,200 10,000 5.0 71.0 8.3 4.4 18.9 3,124 350.0 9.4 3.4 106.0 11,900 421.0 9.2 3.6 124.9 15,024

31/12/2014 1,200 10,000 5.0 98.0 10.4 4.5 32.0 4,400 333.0 7.4 3.3 80.0 11,200 431.0 8.1 3.6 112.0 15,600

31/12/2015 1,200 9,000 4.0 125 12.0 4.4 48.0 5,500 366 8.2 3.7 97.0 13,400 491.0 9.2 3.9 145.0 18,900



6.4 Historical Production The operation of the Augusta Mine was taken over by Mandalay Resources in December 2009. At this time the mine had been operating since early 2006 with a 3-month closure in 2008 - 2009. During this time, approximately 95,000 t were extracted producing 25,000 oz of gold and 4,200 t of antimony. The production record from the start of 2010 to end of 2016 is shown in Table 6-4.

Table 6-4: Historical Mine Production: Mandalay Resources – Costerfield Project

Year Inventory (kt)

Gold Grade (g/t)

Antimony Grade

(%)

Gold Metal Ounces (‘000)

Antimony Metal

(tonnes)

2010 50.7 7.4 4.2 12.0 2,140

2011 72.0 7.3 3.7 16.8 2,637

2012 96.3 8.3 4.3 25.6 4,166

2013 129.6 9.1 4.2 37.7 5,418

2014 167.1 9.1 3.8 48.8 6,345

2015 153.6 11.2 4.2 55.6 6,484

2016 158.4 9.6 3.4 49.0 5,407



7 Geological Setting and Mineralisation 7.1 Regional Geology

The Costerfield gold-antimony mineralisation zone is located at the northern end of the Darraweit Guim province, in the western portion of the Melbourne Zone. In the Heathcote area of the Melbourne Zone, the Murrindindi Supergroup within the Darraweit Guim Province encompasses a very thick sequence of Siluro-Devonian marine sediments. These consist predominantly of siltstone, mudstone, and turbidite sequences.

The western boundary of the Darraweit Guim Province is demarcated by the Cambrian Heathcote Volcanic Belt and north-trending Mt William Fault. The Mt William Fault is a major structural terrain boundary that separates the Bendigo Zone from the Melbourne Zone.

The lower Silurian Costerfield Siltstone is the oldest unit in the Heathcote area and is conformably overlain by the Wappentake Formation (sandstone/ siltstone), the Dargile Formation (mudstone), the McIvor Sandstone, and the Mount Ida Formation (sandstone/ mudstone).

The Melbourne Zone sedimentary sequence has been deformed into a series of large-scale domal folds. These major north-trending, sub-parallel folds in the Darraweit Guim Province include, from west to east: the Mount Ida Syncline; the Costerfield Dome/ Anticline; the Black Cat and Graytown anticlines; and the Rifle Range Syncline. These folds tend to be upright, open, large wavelength curvilinear structures.

The folds have been truncated by significant movements along two major north-trending faults, the Moormbool and Black Cat faults.

The Moormbool Fault has truncated the eastern limb of the Costerfield Anticline, resulting in an asymmetric dome structure. The Moormbool Fault is a major structural boundary separating two structural subdomains in the Melbourne Zone. West of the Moormbool Fault is the Siluro-Devonian sedimentary sequence, hosting the gold-antimony lodes. The thick, predominantly Devonian Broadford Formation sequence occurs to the east of the fault and contains minor gold-dominant mineralisation.



Figure 7-1: Geological map of the Heathcote – Colbinabbin – Nagambie Region Source: Vandenberg et al., 2000.

7.2 Property Geology The Costerfield gold-antimony mineralisation is located on the Costerfield Dome, which contains poorly-exposed lower Silurian Costerfield Siltstone at its core. Within the Costerfield area, four north–northwest (NNW) trending zones of mineralisation have been identified.



They are, from the west:

• Antimony Creek Zone, about 6.5 km southwest of Costerfield, on the outer western flank of the Costerfield Dome

• Western Zone, about 1.5 km west of Costerfield, on the western flank of the Costerfield Dome

• Costerfield Zone, near the crest of the dome, centred on the Costerfield township and hosting the major producing mines and deposits

• Robinsons–Browns Zone, 2 km east of Costerfield.

The Costerfield Siltstone-hosted quartz/ sulphide lodes in the Costerfield Zone are controlled by NNW-trending faults and fractures located predominantly on the west flank of the Costerfield Anticline. This is shown in Figure 7-2.

The mineralised structures in the Costerfield Zone, which dip steeply east or west, are likely to be related to the formation of the Costerfield Dome and the subsequent development of the Moormbool Fault. The main reef systems appear to be developed in proximity to the axial region of the Costerfield Dome. However, due to poor surface exposure, the spatial relationship is based solely on limited underground mapping at the northern end of the zone. The mapping also shows that later faulting (conjugate northwest and northeast faulting) has severely disrupted the mineralised system in the north.

Host rocks are Silurian Costerfield Formation siltstones and mudstones (Costerfield Siltstone). These siltstones and mudstones are estimated to be approximately 550 m thick and are the oldest exposed rocks in the local area.

Significant portions of the local area are obscured by alluvium and colluvium deposits, which have washed out over the plains via braided streams flowing east off the uplifted Heathcote Fault Zone. Some of this alluvial material has been worked for gold but workings are small-scale and limited in extent. Most of the previously mined hard rock deposits were found either out-cropping or discovered by trenching within a few metres of surface. The Augusta Deposit was discovered late in the history of the field (1970) by bedrock geochemistry, buried under 2 - 6 m of alluvium which was deposited at the meandering Mountain Creek/ Wappentake Creek confluence.

Locally, the sedimentary succession of the Costerfield area has been deformed into a broad anticlinal dome structure with numerous cross-cutting reverse thrust faults. This domal structure is thought to have resulted from two separate tectonic events, the first producing shortening in an east–west direction (folding and thrust faulting) and the second producing north–south shortening (gentle warping and mild folding). The anticlinal hinge zone of the Costerfield Anticline has been thrust over its eastern limb by the north–south trending King Cobra Fault zone (see Figure 7-2).



Figure 7-2: Costerfield property geology and old workings Source: Mandalay, 2012.

N



7.3 Stratigraphy of the Costerfield Formation Recent stratigraphic investigations, focused around the currently active Augusta workings within the South Costerfield area, have found many previously unrecognised stratigraphic units and structural features. Sub-surface stratigraphic mapping (drill hole data) has indicated that the local host of mineralisation, the Costerfield Formation, is far more stratigraphically complex than previous investigations have documented. This detailed stratigraphic mapping has consequently identified the complex structural geology of the Costerfield area, and these new observations are documented below.

The stratigraphy of the Heathcote-Costerfield region is presented in Figure 7-3. The oldest outcropping strata documented in the region is the Costerfield Formation, and is regarded as lowest-Silurian in age (Sandford and Holloway, 2006). The Costerfield Formation is then overlain by muddy siltstones and sandstones of the lower Silurian-aged Wappentake Formation, then Dargile Formation. Upper Silurian sedimentation is recorded in coarser silici-clastic successions of the McIvor Sandstone that is then finally overlain by the early-Devonian Mt Ida Formation (Figure 7-3). The Mt Ida Formation records the terminal phase of sedimentation in the greater Heathcote region. The overall stratigraphic thickness of this succession is unknown. However, estimations of true stratigraphic thickness have been in the range of 6 - 7 km, without any significant depositional hiatus.

Figure 7-3: Regional stratigraphic chart of the Costerfield region illustrating age relationships of defined sedimentary succession of the Darraweit Guim Province

Modified from Edwards et al., 1997.



The Costerfield Formation (as defined by Talent, 1965) is a series of thickly bedded mudstones and siltstones featuring heavy bioturbation. The ‘Formation’ nomenclature (assigned by Talent, 1965) is used within this report, instead of the later re-assigned name of ‘Costerfield Siltstone’ (redefined by Vandenberg, 1988). It is recommended that the ‘Siltstone’ nomenclature be abandoned as it has become a misleading term inferring that the unit is composed of siltstone-dominant lithologies. Rather, the unit consists of predominantly mudstone lithologies, with siltstones and sandstone representing the lesser constituents as relatively thin interbedded occurrences.

The Costerfield Formation is dominated by weakly bedded mudstones and silty-mudstones with some lesser siltstone and sandstone constituents. The Formation is informally divided into lower and upper portions on the basis of a significant lithological change midway through the succession. Estimations of the true stratigraphic thickness of the Formation are made difficult due to significant faulting in the area; however, it is estimated to be in the range of 450 - 550 m in thickness, with the lower and upper portions of the Formation being around 200 m and 300 m thick respectively. Informal lithostratigraphic units defined from the Lower Costerfield formation are named the Siliciclastic unit and Quartzite beds. Lithostratigraphic units defined from the Upper Costerfield formation are named the Lower siltstone unit, Augusta beds and the Upper siltstone unit (Figure 7-4).

Figure 7-4: Stratigraphic column of the Costerfield Formation illustrating the relative positions of newly defined (informal) stratigraphic units

Source: Thomas Fromhold, Mandalay, 2014.



7.4 Structural Geology of the South Costerfield area Recent resource-definition diamond drilling for the Augusta and Cuffley deposits has resulted in an extensive collection of geological data from the South Costerfield area. From this data, the construction of refined cross section interpretations has been possible. These cross sections have revealed that the Augusta and Cuffley deposits are bounded between two large, low angle west-dipping parallel thrust faults, named the Adder Fault (upper) and the King Cobra Fault (lower). They are typically around 250 m apart in the South Costerfield area where they are recognised.

The area between these two large faults is also heavily faulted, resulting in a defined zone of intense brittle deformation. Three significant second order faults occur within the fault zone. These faults (named the Flat, Red Belly and Tiger faults) are interpreted as having listric geometry, most likely mimicking the larger structure of the Adder and King Cobra faults.

These faults are all observed as brittle structures, with larger faults (namely the Adder and King Cobra faults) occurring as a 1 - 2 m zone of fault pug, with several metres of heavily fractured and sheered rock in both the foot- and hanging-walls. This zone of intense brittle deformation and shortening is bounded by the larger Adder and King Cobra faults, and regarded as representing a regional-scale thrust fault or thrust zone. It has been informally named the Costerfield Thrust. Mandalay interprets the Costerfield Thrust to be the southern extent of the historically recognised ‘Costerfield Fault’. Stratigraphic interpretations suggest that overall shortening (stratigraphic displacement) across the Costerfield Thrust is around 1 km.

An additional series of brittle faults is observed within this thrust system, striking in an NNE direction (i.e. the East Fault). These faults have a sub-vertical dip and are generally observed as 1 - 2 m thick zones of unconsolidated breccia with minor pug on the fault plane itself. The lateral extent of these faults is uncertain; however, they appear to be localised structures, as correlation through the entire suite of drilling data is highly difficult. Offset across these steep dipping faults appears mostly strike-slip and overall vertical offset is tentatively estimated to be less than 50 m. Lateral offset is at present unknown.

Ductile deformation of the Costerfield Formation occurs as a broad anticlinal structure with a wavelength in an estimated range of 1.5 - 2 km. Smaller parasitic folds are observed to have a northerly striking fold axis that dips slightly to the east. These parasitic folds are assumed to mimic the larger scale folding of the area. Ductile to semi-ductile veining/ faulting is evident within the Costerfield Formation and occurs as 20 - 100 mm laminated quartz veins. They are typically bedding parallel, although laminated veins cross-cutting stratigraphy are not uncommon. Displacement across these faults/ veins is uncertain as their bedding-parallel characteristics mean that estimating displacement through stratigraphic observations is not possible. Veins that cross-cut bedding do appear to record displacement in the range of ten to potentially hundreds of metres.



8 Deposit Types The Costerfield field is part of a broad gold-antimony province mainly confined to the Siluro-Devonian Melbourne Zone. Although antimony often occurs in an epithermal setting (in association with silver, bismuth, tellurium, molybdenum, etc.), the quartz-stibnite-gold narrow veins of the Melbourne Zone are mesothermal-orogenic and are part of a 380 - 370 Ma tectonic event. Gold in Central Victoria is believed to have been derived from the underlying Cambrian greenstones. The origin of the antimony is less certain.

The mineralisation occurs as narrow veins or lodes, typically ≤500 mm wide, hosted within low-grade (anchizone) mudstone and siltstone of the Lower Silurian Costerfield formation. Mineralised shoots in the Costerfield Property are structurally controlled by the intersection of the lodes with major crosscutting, puggy, sheared fault structures. Exploration in the property is guided by predictions of where these fault/ lode intersections might be, using data from structural/ geological mapping, diamond drill logging and 3D computer modelling utilising Surpac software.

Large, flat, west- and northwest-dipping reverse faults have displaced the lodes in the Costerfield Mine at the northern end of the mineralisation extent. It has been recognised that such faults occur throughout the field. At the Alison Mine, production stopped in 1922 because the lodes were ‘lost’ on a flat west-dipping fault, since named the Adder Fault. Drilling in 2011 successfully intersected a displaced lode below the fault, now known as the Cuffley Main Lode. Since the discovery of the Cuffley lode, exploration has continued with success at depth and along strike and the persistent low angle west-dipping faults that continue to influence gold-antimony mineralisation.

8.1 Property Mineralisation The targeted mineralisation in the property occurs at the southern end of a system of steeply-dipping quartz-stibnite lodes, with thicknesses ranging from millimetres to one metre and extending over a strike of at least 4 km. Individual lodes can persist for up to 800 m along-strike and 300 m down-dip. The lode system is centred in the core of the doubly-plunging Costerfield Anticline and is hosted by Costerfield siltstones.

Vein fill mineralogical contents and proportions are found to differ from vein to vein throughout the Augusta and Cuffley lodes. However, the texture and chronological order of each vein mineral generation remains remarkably consistent across all lodes. A diagrammatic illustration of this chronological order of vein history (i.e. paragenesis) of the Augusta and Cuffley deposits is shown in Figure 8-1. The overall paragenetic sequence is ordered as follows: laminated quartz, fibrous carbonate (siderite and ankerite), crystalline quartz (rhombic quartz), stibnite, opaline quartz and milky quartz. Acicular stibnite and botryoidal calcite are not generally associated with the main quartz-stibnite vein structures, and are therefore regarded as a post-mineralisation mineralogical occurrence most likely associated with meteoric events.



Figure 8-1: Paragenetic history and vein genesis of the Costerfield region Source: Thomas Fromhold, Mandalay, 2014.

The Costerfield lodes are typically anastomosing, en échelon style, narrow-vein systems, dipping from 25° to 70° west to steeply east (70° to 90°). Mineralised shoots are observed to plunge to the north (when structurally controlled) and south (when bedding controlled).

The mineralisation occurs as single lodes and vein stockworks associated with brittle fault zones. These bedding and cleavage parallel faults that influence the lode structures range from sharp breaks of less than 1 mm to dilated shears up 3 m wide that locally contain fault gouge, quartz, carbonate, and stibnite. Cross faults, such as those seen offsetting other Costerfield lodes, have been identified in both open-pit and underground workings (refer to Section 7-4 of this report for a detailed description of the structural geology).

Mineralised lodes vary from massive stibnite with microscopic gold to quartz-stibnite with minor visible gold, pyrite, and arsenopyrite. Stibnite is clearly seen to replace quartz. Gold can also be hosted by quartz.

Figure 8-2 is a photograph of typical mineralisation as described above, as seen in E Lode at 1,070 mRL. A close-up photograph of the mineralised lode is shown in Figure 8-3.



Figure 8-2: Two views of E Lode at 1,070 Level South, looking south, with annotation on right-hand side

Note: The mining face is approximately 1.8 m wide and 2.8 m high. For scale, the lens cap is 67 mm.

Source: SRK Consulting, 2010.

Figure 8-3: Close-up of mineralisation in west-dipping E Lode at 1070 Level, looking south Note: For scale, the lens cap is 67 mm.

Source: SRK Consulting, 2010.



8.2 Deposit Mineralisation Mandalay has estimated resources within the Augusta, Cuffley and Brunswick deposits of the Costerfield region. The Augusta Deposit currently comprises 19 lodes (Figure 8-7): E Lode, W Lode, C lode, N Main Lode, NV1 Lode, NV2 Lode, NW Lode, NE Lode, NSW Lode, NSP1 to NSP5 Lodes, P1 Lode, P2 Lode, K Lode, B Lode, and Bob Splay Lode. The Cuffley Deposit is currently comprised of eight lode structures, Cuffley Main (CM Lode), Cuffley East (CE Lode), Cuffley South (CS lode), Cuffley South East (CSE lode), Cuffley Deeps (CD Lode), Cuffley Deeps West (CDW lode), Cuffley Deeps Lower (CDL Lode) and Alison South (AS Lode). The Brunswick deposit comprises the main lode structure (Brunswick shear) and takes in a portion of mineralisation between the Penguin and Kiwi faults that has an offset at depth to the west.

E Lode has an approximate strike length of 600 m and a down-dip extent of 200 m. The lode sub-crops beneath shallow alluvium and has been mined by open-pit and underground methods. The dip of the lode varies from 42° in the south to 72° in the north. Its true thickness ranges from <0.1 m up to 2.8 m, with an average width of 0.34 m where it has been mined.

W Lode lies about 50 m west of E Lode and occurs over a strike length of approximately 420 m. It has a known down-dip extent of approximately 350 m and remains open at depth. The dip of W Lode is approximately 55° above the 1,100 mRL. Below this elevation, it gradually steepens to between 70° and 80° at around the 900 mRL. Its true thickness ranges from >0.1 m up to 2.7 m, with an average width of 0.36 m where it has been mined.

C Lode lies an average of 20 m to the west of W Lode. It has a strike extent of 200 m before curving at both the southern and northern extents to intersect W Lode. Its geometry mirrors that of W lode, dipping approximately 60° to the west. It is exposed at surface and extends down dip to 180 m. The true thickness varies from 0.1 m to 1.5 m.

N Main Lode lies approximately 90 m to the west of E Lode. N Main Lode was previously referred to as NE Lode. The lode is sub-vertical and can be interpreted in drilling over a strike length of approximately 700 m with a down-dip extent of approximately 350 m. The true thickness ranges from <0.1 m up to 2.7 m, with an average width of 0.3 m where it has been mined.

The two New Vein Lodes (NV1 and NV2) lie 30 m and 20 m respectively to the east of N Main. They appear as splays off N Main as it changes strike at 4,740 mN and again at 4,870 mN. Both dip at 80° to the West. NV1 sits to the east of NV2 and has a strike extent of 200 m, compared to 180 m for NV1. NV2 averages 130 m vertical extent while NV1 averages approximately 70 m vertical extent. Both exhibit similar Lode widths varying from 0.05 m to 1.6 m, averaging 0.28 m in drilling results and face samples.

NW Lode is located 5 - 10 m to the west of N Main, just above the hanging wall plane of the King Cobra fault. It is a localised lode with a strike extent of 70 m and a vertical extent of 25 m. Mineralisation is sub-vertical along bedding planes. True thickness of the lode ranges from 0.1 m to 2 m and averages 0.4 m.

Five N Splay Lodes (NSP1-5) lie within a structurally complex 20 m wide panel to the east of N Main between 4,800 mN and 4,900 mN. The NSP Lodes sit in a similar structural position to NW Lode, where mineralisation is exploiting bedding parallel fractures above the King Cobra Fault system. However, The NSP Lodes extend up to 50 m vertically. Lode thicknesses are variable, ranging from 0.01 m to 1.8 m and averaging 0.4 m.

P1 Lode lies approximately 80 m to the west of E Lode. The lode dips at approximately 85° to the east and occurs over a strike length of approximately 80 m, with a down-dip extent of approximately 150 m. It is interpreted as a small transfer structure between W Lode and N Main Lode. The true thickness ranges from <0.1 m up to 1.9 m, with an average width of 0.4 m where it has been mined.



P2 Lode is a continuation of the P1 structure beyond its intersection with NM Lode. It is sub-vertical with a strike length of 200 m and a depth extent of 160 m. The majority of this lode structure is sub-economic, with only a small portion of the upper structure being mined.

K Lode lies 30 m west of E lode, immediately north of the W Lode termination. It is a bedded parallel, laminated quartz-stibnite vein. It strikes at approximately 340° extending up to 100 m laterally, and has a consistent 60° dip to the west down, extending 70 m vertically. Mineralisation appears to be less significant on the western side of N Lode. Lode thickness varies from 0.1 m to 0.7 m, averaging 0.27 m.

Bob Lode lies 30 m east of N Lode striking approximately 30°, extending 200 m laterally. It has a vertical extent of 150 m and dips 80° towards the west. Lode widths vary between 0.05 m and 1.8 m, averaging 0.031 m. Bob Splay (BSP) splays off Bob Lode at 4,640 mN striking 10° and dipping 75° west. BSP has a strike extent of 60 m and a vertical extent of 150 m. Lode thicknesses are less than for Bob Lode, ranging from 0.03 m to 0.75 m, averaging 0.016 m.

NE Lode lies approximately 15 m to the east of N Main Lode. It is interpreted to be a sub-vertical lode with a north–south strike of approximately 140 m and a vertical extent of 230 m. It is comprised of one stibnite-quartz vein of 0.1 m to 0.25 m in width.

NSW Lode lies approximately 110 m to the west of E Lode. The lode dips at approximately 55° to the southeast, and occurs over a strike length of approximately 60 m with a down-dip extent of approximately 50 m. It is interpreted as a small splay at the southern end of N Main Lode. The true thickness ranges from <0.1 m up to 2.4 m, with an average width of 0.4 m where it has been mined.

The northwest striking Cuffley Main Lode lies approximately 200 m to the west of E Lode and appears as an offset depth continuation of the Alison east and west lodes (possibly correlates with the ‘west lode’). The upper reaches of Cuffley are separated from the Alison Reefs by a flat (30°- 40°) west dipping fault (Flat Fault) and extend though smaller fault offsets and splay veins to the King Cobra fault over a vertical extent of ~300 m. Along strike stibnite veining is split into two domains by a northwest dipping fault (East Fault). The intersection of the mineralised corridor with this fault identifies the northerly plunging extent of the southern domain of economic mineralisation. The southern domain exists for ~150 m with a gentle strike change leading into the East Fault from west to northwest dipping.

Immediately north of the East Fault, a zone of sub economic mineralisation exists within a small (5 mm - 200 mm) sheared zone. Au and Sb concentrations within the shear increase rapidly, coincident with small-scale faulting ~50 m north of the East Fault, marking the northern domain of the Cuffley deposit. The northern domain exists for ~150 m before the accommodating structure pinches to a series of small veinlets and shears.

The vein dilation within the mined areas of Cuffley exhibits a vein width of up to 4 m. On average, the width is ~0.4 m. The proximity and intersection of sub-parallel quartz veins is observed to be a large control on dilation. Local splaying on veining has been identified as the main Cuffley shear diverts from the underdeveloped cleavage plains of the country rock.

Extending below the Cuffley Lode, the Cuffley Deeps mineralisation was previously thought to be a structurally offset continuation of Cuffley Main below the Tiger Fault. Revised interpretation informed by mining activity suggests that the two lodes are continuous (Figure 8-4). The western shift of Cuffley Deeps relative to Cuffley Main is caused by a shallow rollover of the lode. In the area of this rollover, the lode dips approximately 50° to the west. Cuffley Deeps has a strike extent of 500 m and a depth vertical extent of 200 m below the abovementioned rollover. Cuffley Deeps West sits 30 m to the west of Cuffley Deeps, striking at approximately 340°. It has a strike extent of 250 m and an average vertical extent of 75 m. Cuffley Deeps Lower sits below the interpreted Fox Fault, above the King Cobra



hanging wall. These faults define the upper and lower limits of the lode. It extends 250 m laterally and ~50 m vertically.

Extending below the Cuffley Deeps mineralisation, four zones of mineralisation for Sub King Cobra have been identified (Figure 8-8): the Central Main, Central East, Central LQ and Western targets. These targets sit to the west of, and below, the Cuffley Deeps mineralisation. The known high grade portion of the Central Target extends vertically 100 m and 300 m along strike, and is situated on the eastern limb of a large anticline. The mineralisation is breached by the King Cobra thrust system (Figure 8-4). The Western target sits below and to the west of the Central target, and is likely to be an offset depth extension of the Central target.



Figure 8-4: Schematic cross section 5050N northing, illustrating the ‘apparent offset’ between the Cuffley Deeps lode and Cuffley lodes across the west-dipping series of faults

Source: Mandalay Resources, 2017.

Cuffley System Schematic cross section – 5050N



Figure 8-5: Brunswick cross section 5880N Source: Mandalay Resources, 2017.

The Brunswick Lode lies approximately 600 m northwest of the northernmost point of the Cuffley Lode. The lode is sub-vertical and occurs over a strike length of approximately 450 m, with a down-dip extent of approximately 200 m and an average true thickness of about 1.28 m. Recent exploration drilling has identified mineralisation to the south and at depth below the known Brunswick lode (Figure 8-5). Mineralisation is broken into two zones of mineralisation below the Brunswick lode, called the P-K domain and Brunswick Deeps. The P-K domain is capped by the shallow west-dipping Penguin fault, and extends to the Kiwi Fault (Figure 8-5). The Brunswick Deeps zone is defined by the west-dipping Kiwi Fault and Adder Fault. Similar in nature to Cuffley and N lodes, the mineralisation in the two domains is generally confined to sub vertical quartz–stibnite veins. Drilling is ongoing at the time of this report, and the Brunswick lode resource in this report encompasses mineralisation down to the Kiwi Fault.



Figure 8-7: Deposit wireframes



Figure 8-8: Structural relationship between the Augusta, Cuffley, and Brunswick Lode systems



9 Exploration Exploration work that led to the discovery of the Augusta and Cuffley Deposits has consisted of prospecting, trenching, geological mapping, geophysics, geochemistry, and diamond drill testing of interpreted geological targets. Geochemical methods have proven to be applicable in detecting gold-antimony mineralisation. See Chapter 10 for Mandalay’s diamond drilling details from 2009 to present.

9.1 Costean/ Trenching Previous owners have undertaken trenching, but records of these exploration activities are limited.

9.2 Petrophysical Analysis In 2006, AGD submitted a suite of 22 rock and mineralised samples from all known deposits around Costerfield for testing by Systems Exploration (NSW) Pty Ltd. The aim was to determine their petrophysical properties and to identify the most effective geophysical methods that could be used in the field to detect similar mineralisation.

Of the samples submitted, thirteen were mineralised and were sourced from Augusta, Margaret, Antimony Creek, Costerfield, Bombay, Alison and Brunswick; two were weathered mineralisation and were sourced from Augusta; and seven were waste.

The following petrophysical measurements were taken:

• Mass properties:

− Dry bulk density

− Apparent porosity

− Grain density

− Wet bulk density.

• Inductive properties:

− Magnetic susceptibility

− Diamagnetic susceptibility

− Electromagnetic conductivity.

• Galvanic properties:

− Galvanic resistivity

− Chargeability.

Although there are measurable differences in the physical properties of mineralised and non-mineralised material at Costerfield, they are marginal at best, and it is unlikely that the differences present would result in clear geophysical signatures. The only field techniques recommended for trialling were ground-based magnetic, gravity, and induced polarisation (IP) profiling.

9.3 Geophysics

Ground Geophysics Based on the results of the petrophysical testing program, a limited program of ground magnetic, gravity, and IP profiling, with optimal measurement parameters, was carried out across the Augusta Deposit. None of the techniques were found to be effective for detecting the known mineralisation at Augusta.



Airborne Geophysics A low-level detailed airborne magnetic and radiometric survey was undertaken in 2008 over the AGD tenements, including both Augusta and Cuffley. The airborne survey was conducted on east–west lines spaced 50 m apart, with a terrain clearance of approximately 50 m. Survey details are included in a logistics report prepared by UTS (UTS, 2008).

Magnetic data was recorded at 0.1 second intervals and radiometric data was recorded at 1 second intervals. Additional processing was undertaken by Greenfields Geophysics. Interpretation of the radiometric and magnetic data resulted in regional lineament trends across the tenements, which assisted in interpreting the local buried structures.

9.4 Geochemistry

Mobile Metal Ion Based on historic geochemical surveys over the Augusta Deposit as described by Stock and Zaki in 1972, and informal recommendations by Dr G McArthur of McArthur Ore Deposit Assessments Pty Ltd (MODA), it was decided by AGD geologists to trial mobile metal ion (MMI) analytical techniques on samples collected on traverses across the Augusta lodes in 2005.

Utilising two geophysical traverse lines across the Augusta Deposit, 5 m spaced samples were collected from the soil horizon and submitted to Genalysis Laboratory Services (Genalysis) for MMI analysis of gold, arsenic, mercury, molybdenum, and antimony via inductively coupled plasma (ICP).

While the other elements showed no correlation to the underlying mineralisation, the gold and antimony results appeared to show a broad anomaly across the mineralisation, indicating that the technique could be useful for regional exploration.

Bedrock Geochemistry Augusta South Orientation Bedrock Geochemistry Aircore Drilling (MIN 4644) The effectiveness of bedrock geochemistry was demonstrated by Mid-East Minerals NL (MEM) in 1968 - 1970, when a grid south of South Costerfield/ Tait’s Shafts was sampled. What is now known as the Augusta Gold-Antimony Deposit was highlighted by the resultant anomalies. Although MEM drilled three shallow (22 m - 57 m) diamond drill holes to test the anomalies and intersected stibnite stringers, they did not proceed any further. Both conventional surface soil and bedrock samples were collected to compare techniques, although the surface samples were anomalous, and (cheaper) bedrock samples defined the lodes more precisely.

During March of 2010, Mandalay employed the use of a Mantis 200 drill rig from Blacklaws Drilling, Elphinstone, Victoria, to provide aircore drilling services, with an NQ size bit. Drill holes were at 15 m intervals on traverses 100 m apart. Because of the soft, weathered nature of the bedrock, samples consisted mainly of silt size cuttings (with only rare segments of core). In the later part of the program, the aircore bit was replaced by a reverse circulation (RC) bit. A total of 104 holes were drilled for a cumulative total of 547 m and an average hole depth of 5.2 m.

Cuttings were collected (in a cyclone) at 1 m intervals downhole in each drill hole and laid out in sequence on the ground. Mandalay staff supervising the program determined when bedrock was intersected by observing the cuttings and ensured that the driller penetrated at least 1 m into bedrock. A 5 kg ‘top-of-bedrock’ sample was collected in a plastic sample bag, with the coordinates labelled on it. Each hole was logged and the sample number assigned.

Mandalay staff took a 1 kg subsample via a riffle splitter. The sample was then placed in a calico sample bag labelled with the assigned sample number. The remaining sample was retained in labelled



plastic bags (sample coordinates, sample number) for possible future follow-up assay. This process achieves the best possible sample and preserves it for future reference or quality checks. The aircore drilling method is prone to downhole contamination. However, the likelihood of this was reduced because the sample material was visually inspected by Mandalay geologists to ensure that it was taken from bedrock. The use of certified standards was routine practice throughout the sample assay program.

Plots of the MEM 1970 bedrock geochemistry on the Augusta Lode plans illustrate that antimony mineralisation within 50 m of surface (or top-of-bedrock) can be accurately detected by close-spaced (15 m) sampling on 100 m traverses across strike (Figure 9-1). The antimony halo is subdued where the high-grade lode is greater than 50 m below top-of-bedrock. For example, on cross section 4550N, drill hole M032 intersected E Lode (55% Sb, 59 g/t Au) 35 m below surface and drill hole MH041 intersected E Lode up-dip (35.3% Sb, 36.8 g/t Ay) 15 m below surface. The top-of-bedrock geochemical expression of E Lode on this traverse was a 50 m zone in the 200 - 400 ppm (0.2% - 0.4%) range. Thus, a subdued bedrock geochemistry anomaly could mean either that a low-grade lode exists at shallow depth, or a high-grade lode exists at depth.

Figure 9-1: Augusta South aircore program results with lode locations Source: Mandalay, 2012.

Augusta South Auger Bedrock Geochemistry Drilling (MIN 4644) Mandalay engaged Statewide Drilling to undertake the Augusta South auger geochemistry program with an Edson 200 auger rig. The three traverse programs (at 3600N, 3500N and 3400N), consisted of 224 holes at 10 m spacing, averaged 3.2 m deep, and was completed on 26 May 2012 for a total of 732 m. This program tested the zone between Augusta South and the Margaret Mine (south of the operating Augusta Mine). The three east–west traverses were across cleared grazing paddocks, south of Tobin’s Lane, Costerfield.

Overall, sample recovery was good with sample cuttings from the auger laid out in sequence on the ground. Mandalay staff supervising the program determined when bedrock was intersected by

N



observing the cuttings, and ensured that the driller penetrated at least 1 m into bedrock. A 3 - 5 kg ‘top-of-bedrock’ sample was collected in a plastic sample bag, labelled with the coordinates. Each hole was logged and the sample number assigned.

Mandalay staff re-bagged the 224 samples into calico bags using a sample spear to produce 1 kg subsamples. Each bag was then labelled with its assigned sample numbers and dispatched to Onsite Laboratories in Bendigo. The use of certified standards was routine practice throughout the sample assay program. Using a spear device to subsample introduces potential bias to the resultant assay, as particles do not have an equal opportunity of being sampled. However, it was considered to be a satisfactory method because the results were only used internally within Mandalay as an indication of anomalism, not used as input to resource estimation.

Two parallel NNW-trending anomalous zones were defined by this survey, as shown in Figure 9-2. The western trend appears to be the northern extension of the Margaret Reef Zone.

Figure 9-2: Augusta South (Margaret Zone) – Auger drill 2011 bedrock geochemistry: antimony contours

Source: Mandalay, 2012.

N



Augusta East – Auger Bedrock Drilling Program From December 2011, Mandalay engaged Starwest Pty Ltd to undertake the Augusta East Auger drilling program. This program utilised a Gemco 201B auger drill rig. The Auger drill is a track-mounted, hydraulically powered rig, with 14 cm auger rods, capable of drilling up to 2,000 m per month. A total of 2,615 auger holes were drilled for 7,295.6 m between December 2011 and June 2012.

A soil sampling traverse through the Augusta East area in the 1990s identified a weak N–S anomalous zone. The area was deemed a priority area due to the fact that the crest of the Costerfield Dome is interpreted to pass through the area, and it is therefore prospective for quartz-stibnite lodes similar to those in the Dome crest at the Costerfield-Minerva-Bombay group of mines to the north. The survey revealed three anomalous zones, as shown in Figure 9-3. Further follow-up diamond drilling is planned for this area.

Figure 9-3: Auger geochemistry results displayed as antimony contours Source: Mandalay, 2012.

A total of 1,375 auger holes were then drilled by Mandalay from April to June 2014, for 3,906 m. Holes were drilled over exploration licenses EL 3310 and EL 5432 and mining license MIN 4644 covering six prospect areas: Augusta, Cuffley, Brunswick, North, West Costerfield and Reef. Results from duplicate samples taken during the auger program show excellent correlation, therefore showing that good sampling practices were maintained throughout the program. Table 9-1 summarises all auger drilling totals by prospect location and tenement.



Table 9-1: Auger drill hole totals by prospect

Location Location description Hole IDs No. of

holes Av. Depth

(m) Metres drilled Licence

ALD Cuffley ASAC2100-2341 76 2.26 171.4 MIN 4644

AUG Augusta Mine extension ASAC2127-2355 124 4.15 512.7 MIN 4644,

EL 3310

BRU Brunswick BRA001-0247 247 2.80 690.4 MIN 4644, EL 3310

MAR Margaret’s Reef MAA0001-0542 536 3.00 1,594.6 MIN 4644, EL 3310, EL 5432

NCOS North Costerfield ASAC2237-2292 56 2.94 164.7 MIN 4644

WCOS West Costerfield WCA0642-0977 336 2.30 772.3 MIN 4644, EL 3310

Total 1,375 2.85 3,906.1

A summary of the prospect locations is given below.

Cuffley A total of 76 holes were drilled on two lines over the underground Cuffley deposit, to test the relationship between bedrock geochemistry and known gold-antimony orebodies below surface. The Cuffley orebody does not outcrop at surface due to termination of the vein system at a flat fault approximately 100 m below surface. The depth to the ore zone may explain the low to moderate level of anomalism displayed from the auger drilling. The anomaly covers a broad zone which roughly correlates to the Cuffley orebody at depth.

Augusta Mine Extension To the east, west and south of the existing Augusta mine site, 124 holes were drilled to test for extensions of the known underground orebodies. The auger drilling to the east and west detected no elevated levels of either gold or antimony and no further work is planned in these areas.

The two lines drilled to the south displayed a narrow zone of high anomalism, which correlates directly to extensions of known orebodies. Diamond drilling between this area and the mine intersected no economic mineralisation; therefore, this area represents a low priority drilling target.

Brunswick To the west and south of the Brunswick open cut, 247 auger holes were drilled to test for extensions of the known orebody. No elevated anomalism was detected to the west but a narrow high grade intersection was returned from drilling 500 m south of the Brunswick pit, suggesting extension of the orebody.

In April 2016 diamond drilling recommenced on Brunswick to follow up on the successful 2015 drilling program. Further economic assessment of the current resources at Brunswick has led to the need to test mineralisation below the current resource and to the south along strike. The Brunswick deposit has been mined previously as both an underground and open cut mine, and is conveniently located next to the Costerfield processing plant.

Margaret’s Reef Auger drilling at Margaret’s Reef was carried out on private property 1 km south of current Augusta mining operations, with a total of 536 holes. Previous auger drilling in this area was done on a wider sample spacing of 40 m and was not considered deep enough to provide consistent results, so lines



were re-drilled in this program. Sample spacing of 10 m over the previous anomalous results gave a clearer indication of vein structures at depth.

Margaret’s Reef appears to be made up of several reef vein systems, as suggested from previous RC and diamond drilling. The veins strike approximately northwest, which is a similar vein orientation to veins seen underground at Augusta and Cuffley, and may represent a fault displaced extension of one of these systems. The close proximity to the King Cobra fault to the east appears to have structurally complicated the vein systems, which may explain why the anomalism appears discontinuous in nature. Wide zones of high anomalism correlate to known historic workings over the reef. The highest result received from auger drilling not proximal to current mining operations was received from the northernmost line at Margaret’s Reef, with 5.42 g/t Au and 3.25% Sb, suggesting economic mineralisation at surface.

Several high priority diamond drill targets have been planned, including a target beneath the abovementioned high result, to give further structural information on this vein system. Recent diamond drilling failed to follow up on the high grade intersection from hole MM001 (drilled in 2001) of 1 m at 33 g/t gold and 14% antimony. Further diamond holes are now planned to take into account the bedrock geochemistry to help prove that an economic resource exists in this area. There is a current exploration agreement in place with the landowner to provide immediate access.

West Costerfield A total of 336 auger holes were drilled in 2014 at West Costerfield to test near historic workings to the east and continuity to the south of the previous auger program. The previous geochemistry program delineated the True Blue anomaly to the west, but only the northern section of the West Costerfield reef was explored.

A broad anomaly was defined over the West Costerfield reef, which continues south with high gold values and only moderate antimony results. The anomaly runs along the Mountain Creek drainage zone to the south but widens and slightly changes orientation towards the north, near the small historic pits that define the West Costerfield reef. Though the antimony anomalism is subdued in contrast with the high gold values, there are interpreted gold-antimony veins below surface similar in style to those intersected in a single diamond drill hole over True Blue.

In 2015, 38 follow-up RC holes were drilled to test the anomaly identified in the 2014 auger drilling program. The RC drilling resulted in the identification of mineralisation that warrants follow-up diamond drilling.

9.5 Aerial Photogrammetry Survey In 2005, AGD commissioned Quarry Survey Solutions of Healesville, and United Photo and Graphic Services Pty Ltd of Melbourne, to organise and carry out aerial photogrammetry of the Costerfield Project tenements and the Augusta Mine Site.

High-level (24,000 ft) photo coverage was carried out in November 2005. This was followed by low-level (8,000 ft) coverage over the Augusta Mine Site in January 2006.

A second low-level (4,000 ft) flight was carried out in April 2006, at the time of maximum surface excavation, prior to the commencement of backfilling of the E Lode Pit.

The various photo sets were subsequently used to generate a digital terrain model (DTM) and a referenced orthophotographic scan of the Costerfield central mine area. This area essentially extends from Costerfield south to the Margaret area, thereby encompassing most of Mining Licence MIN 4644.



10 Drilling 10.1 Mandalay Resources (2009 - Present)

On 1 December 2009, Mandalay took over the Costerfield Operations from AGD and continued with exploration across tenements MIN 4644, EL 3310, and EL4848. Drilling conducted on the tenements since then is shown in Table 10-1.

Table 10-1: Drill hole summary

Year Metres (diamond)

Metres (percussion/ auger)

2009 459 547

2010 4,032 0

2011 13,515 0

2012 18,581 7,296

2013 24,329 3,838

2014 20,817 3,906

2015 18,439 2,732

2016 32,995 0

Total 133,167 18,318

10.2 2009/ 2010 From 1 July 2009 to 30 June 2010, drilling mainly consisted of drilling along strike and down-dip from the existing Augusta Resource. In total, 458.9 m of diamond coring was undertaken.

In addition, 547 m of bedrock geochemistry aircore drilling was completed within MIN 4644 at Augusta South.

From 1 July 2009 to 30 June 2010, drilling at Augusta concentrated on the definition of the W Lode Resource. Four drill holes tested the depth extent of W Lode. Another six holes were designed as infill holes to test mineralised shoots and gather geotechnical data. A list of significant intersections for this period was announced to the market in January 2011 (Mandalay, January 2011).

10.3 2010/ 2011 Exploration was undertaken on two projects from 1 July 2010 to 30 June 2011 – the Augusta Deeps project and the Brownfields Exploration project. The Augusta Deeps project was undertaken with the view to extending the current Augusta Resource to depth.

From 1 July 2010 to 30 June 2011, drilling at Augusta concentrated on the infill and extension beneath Augusta to further define the Resource below the 1,000 mRL. In total, 10,622.7 m were drilled beneath the Augusta mine workings, which resulted in the definition of further Indicated and Inferred Resources. A list of significant intersections for this period was announced to the market in August 2011 (Mandalay, August 2011).

10.4 2011/ 2012 From 1 July 2011 to 30 June 2012, exploration was undertaken on four projects . These were: the Augusta Deeps drilling project (W Lode and N Main Lode), the Alison/ Cuffley drilling project, the Brownfields/ Target Testing drilling project and the Target Generation/ Bedrock Geochemistry auger drilling project.



The Augusta Deeps project was undertaken with the view to extend the current Augusta Resource to depth and along strike. N Main Lode was drilled both from underground and surface-based rigs. The Alison/ Cuffley drilling project was designed to infill drill a portion of the lode to Indicated Resource category, and endeavoured to ‘bound’ the limits of the lode to Inferred Resource category.

In total, 18,581.4 m of diamond coring and 7,295.6 m of auger drilling was undertaken on the four projects. All drilling was carried out by Starwest Pty Ltd using one LM75 diamond drill rig, two LM90 diamond rigs, one Kempe underground diamond drill rig and one modified Gemco 210B track-mounted auger rig.

Drilling of the Augusta Deposit from July 2011 to December 2012 was undertaken with the view to extend the W, E and N Main Lodes’ Inferred and Indicated Mineral Resource, and give confidence in the structural continuity of the W and N Main Lodes. A total of 78 holes were drilled from surface and underground, totalling 16,170.4 m of drilling. A list of significant intersections for this period was announced to the market in July 2012 (Mandalay, August 2011).

The Cuffley Lode resource drilling program began in July 2011 (AD series of holes), following the MB007 discovery. As a follow-up program, four holes were drilled (AD001 - AD004). Hole AD004 went through the fault blank, and AD003 appears to have only intersected the Alison Lode above the Adder Fault in the vicinity of some old stopes.

From hole AD005 onwards; the drilling strategy involved drilling at least two holes on each mine grid cross section, with an approximate spacing of 80 - 100 m. Holes have been drilled both from west to east and east to west, depending on site logistics.

A portion of the drilling in 2011/ 2012 was infill drilling (100 m below the Alison Shaft 5 level) at a spacing of 40 m, to define the lode to Indicated Resource category where the planned access decline would first intersect the lode. One deep hole, AD022 (5025N cross section) intersected the Cuffley Lode (1.04 m, 59.7 g/t Au, 0.37% Sb) at 700 mRL, 490 m below the surface. This provided confidence in the depth continuity of the lode to Inferred Resource category initially, to ‘bound’ the extent of the lode.

10.5 2012/ 2013 Cuffley Lode Drilling From 1 July 2012 to 30 June 2013, Mandalay Resources drilled 24,329 m of diamond drilling, targeting the Cuffley Lode from surface. The drill program focused on infill drilling the central, high-grade part of the Cuffley Lode to convert some of the Inferred Mineral Resources to the Indicated category. Longitudinal projections of these intersections relative to the Mineral Resource are provided in Section 14. The Cuffley Lode dips approximately 85° towards 097°. All downhole sample lengths have been converted to true thicknesses using the dip of the lode and the orientation of the drill hole. MH335 and MH336 were drilled as wedges; therefore, they have significantly lower collar elevations and shallower dips than the other drill holes. Due to the narrow high-grade nature of the mineralisation, it is not meaningful to report significantly higher-grade intercepts within lower grade intercepts; therefore they are not reported here.



10.6 2014 Cuffley/ N Lode Drilling In 2014 the focus was on finalising the Cuffley and Augusta Resource Drilling. The goals achieved included:

• Expanding the existing Inferred Resource of the Cuffley Lode, both along strike and at depth

• Increasing the confidence of the central part of the Cuffley Lode to aid mine development

• Stoping the Cuffley Lode


• Infilling and extending the Cuffley resource to the north and south, along with ‘Cuffley Shallows’ in between the flat fault and the Adder Fault.

In total, 20,817 m of diamond drilling and 3,906 m of auger drilling were undertaken on Mandalay’s tenements at Costerfield during 2013/ 2014. A total of 5,735 m were drilled for the purposes of target testing, along with 9,390 m for resource expansion and conversions, and 5,692 m for resource infill drilling. All drilling activity was conducted by Starwest Pty Ltd using two Boart Longyear LM90s, one Boart Longyear LM75, one pneumatic Kempe U2 and one modified Gemco 210B track-mounted auger.

10.7 2015 Cuffley/ N Lode/ Cuffley Deeps/ Sub King Cobra Drilling Exploration in 2015 was focused on extending the Cuffley and Augusta Resource, both along strike and at depth. The expansion of the Cuffley resource included the commencement of drilling in the Cuffley Deeps and Sub King Cobra regions. The goals achieved included:

• Expanding the existing Inferred Resource of the Cuffley Lode along strike

• Defining a resource below the Cuffley Lode at depth

• Commencing drilling at depth below the Cuffley Deposit, into the ‘Cuffley Deeps’ and ‘Sub King Cobra’ areas

• Increasing the confidence of the central part of the Cuffley Lode to aid mine development

• Stoping the Cuffley Lode


• Infilling and extending the Cuffley resource to the north and south, along with ‘Cuffley Shallows’ in between the flat fault and the Adder Fault

• Follow-up RC drilling at West Costerfield to test the geochemical anomaly identified in 2014 by the Auger Bedrock drilling program.

In total, 18,439 m of diamond drilling and 2,732 m of RC drilling were undertaken on Mandalay’s tenements at Costerfield during 2014/ 2015 . All drilling activity was conducted by Starwest Pty Ltd using two Boart Longyear LM90s, one Boart Longyear LM75 and one pneumatic Kempe U2, with the exception of the RC drilling, which was carried out by Blacklaws Drilling utilising a Hanjin surface rig.

10.8 2016 Cuffley Deeps/ Cuffley South/ M and New Lode/ Sub King Cobra Drilling/ Margaret/ Brunswick Exploration from January to December 2016 was focused on extending and upgrading the Cuffley and Augusta Resources to build ‘life of mine’ and replace mine depletion, along with exploring near mine targets in close proximity to current underground infrastructure. The expansion of the Cuffley resource



has included the continuation of drilling in the Cuffley Deeps, Cuffley South and Sub King Cobra regions, along with the addition of new target areas. The goals achieved included:

• Further definition of the Cuffley Deeps and Sub King Cobra resources below the Cuffley Lode at depth

• Continuing infill and exploration drilling in the Cuffley Deeps and Sub King Cobra areas, leading to a resource expansion on Cuffley Deeps and an Inferred resource on the Sub King Cobra domain

• Infill drilling of Cuffley Deeps, delineating further prospective zones and a new ore system; namely mid lode (M Lode) located between the Cuffley line of lode and N Lode

• Further development on Cuffley Lode, informing understandings and increased confidence in Cuffley Deeps at depth and along strike

• Infilling and testing the Cuffley resource to the north and south, along with ‘Cuffley Shallows’ in between the flat fault and the Adder Fault

• Recommencing drilling on Brunswick, and further testing of the deposit to the south and at depth

• Brownfields drilling at Margaret Reef, identifying the Margaret East mineralisation.

In total 32,995 m of diamond drilling was undertaken on Mandalay’s tenements at Costerfield during 2016 (Table 10-2). All drilling activity was conducted by Starwest Pty Ltd using four Boart Longyear LM90s, one Boart Longyear LM75 and one pneumatic Kempe U2.



Table 10-2: Significant Cuffley Main, Cuffley Deeps, Sub King Cobra, Brunswick, Margaret, N Lode and New Vein Lode drill hole intercepts: 2016

Hole ID Intercept Easting (mine

grid, m)

Intercept Northing

(mine grid, m)

Elevation (m)

True width (m)

Gold grade (g/t)

Antimony grade

(%)

AuEq (g/t)

over 1.8 m or TW

Total hole

depth (m)

Target

AD134 15,175 5,027 755 0.16 103.7 34.5 14.6 270.1 Cuffley Deeps

AD140 15,176 5,042 727 0.36 4.1 1.4 1.3 255.9 Cuffley Deeps

AD143 15,176 5,015 775 0.06 2.9 4.7 0.4 224.6 Cuffley Deeps

AD147 15,181 5,053 794 0.10 11.9 3.2 0.9 191.6 Cuffley Deeps

AD148 15,173 5,102 774 0.16 12.1 5.9 1.9 248 Cuffley Deeps

AD149A 15,179 4,992 697 0.99 31.1 6.7 23.4 336.2 Cuffley Deeps

AD151 15,173 5,072 717 0.24 9.8 5.5 2.6 250 Cuffley Deeps

AD152 15,174 5,131 804 0.19 34.6 11.5 5.7 225.1 Cuffley Deeps

AD153A 15,178 5,049 703 0.18 1.6 7.1 1.3 326.4 Cuffley Deeps

AD165 15,175 5,123 773 1.53 11.7 7.2 20.2 320 Cuffley Deeps

AD165W1 15,173 5,124 775 0.27 1.9 2.7 1.0 218.7 Cuffley Deeps

AD144 15,161 5,114 722 0.06 34.9 0.0 1.1 269.8 CD West

AD146 15,157 5,153 698 1.11 11.1 3.8 10.9 300.6 CD West

AD161 15,193 4,560 753 0.68 33.9 0.9 13.4 414.4 M Lode

AD158 15,196 4,552 799 0.10 150.6 50.1 13.1 437.5 M Lode

CSK011 15,171 4,616 545 0.26 9.00 0.00 1.30 750 Central East

CSK011 15,086 4,665 432 0.08 0.18 0.00 0.01 750 Western target

CSK012 15,052 4,809 287 0.19 344.69 0.01 37.18 810.1 Western target

CSK014 15,231 4,807 721 0.11 1.18 25.00 3.03 205 Unnamed

CSK014W2 15,157 4,774 600 0.08 0.29 8.21 0.75 599.5 Unnamed

CSK016A 15,133 4,874 545 0.07 0.80 15.10 1.20 585.7 Central East

CSK016A 15,128 4,875 538 0.11 52.10 0.01 3.11 585.7 Central East

CSK016A 15,095 4,884 489 0.04 7.95 12.70 0.82 585.7 Central main

CSK016W1 15,099 4,881 513 0.10 14.30 0.00 0.77 393.1 Central

CSK016W1 15,095 4,882 508 0.11 39.80 18.50 4.55 393.1 Central

CSK017 15,265 4,844 752 0.00 2.39 7.52 0.00 115.9 Unnamed

CSK017B 15,263 4,845 750 0.00 3.77 2.68 0.00 720.3 Unnamed

CSK019A 15,153 5,025 549 0.20 6.57 6.35 2.16 502.7 Central main

CSK024 15,120 4,899 549 0.19 321.0 0.2 33.4 588 Central East

CSK021 15,091 4,774 482 0.38 7.2 0.0 1.5 597.3 Central Main

CSK022 15,091 4,834 507 0.19 5.3 2.5 1.0 603.1 Central Main

CSK022 15,152 4,832 539 0.05 6.7 0.4 0.2 603.1 Central LQ

CSK022 15,141 4,830 555 0.18 4.8 1.8 0.8 603.1 Central East

CSK023 15,102 4,846 484 2.00 7.1 0.5 8.0 547.7 Central Main

CSK023 15,105 4,845 487 0.24 2.6 1.6 0.7 547.7 Central Main

CSK023 15,138 4,843 523 0.09 4.7 4.6 0.7 547.7 Central East

CSK024 15,106 4,902 536 0.64 9.9 0.5 3.8 588 Central Main

CSK026A 15,127 4,993 547 0.14 25.7 13.0 3.6 445.7 Central Main

CSK026A 15,114 5,001 535 0.18 20.9 0.0 2.1 445.7 Unnamed

CSK026A 15,124 4,994 545 0.84 1.9 2.4 2.7 445.7 Central Main

CSK027 15,129 4,905 516 0.23 4.4 8.2 2.3 500.4 Central Main

CSK027 15,118 4,957 510 0.17 0.2 9.8 1.5 500.4 Unnamed

CSK027 15,081 4,976 469 0.22 7.3 0.1 0.9 500.4 Unnamed



Hole ID Intercept Easting (mine

grid, m)

Intercept Northing

(mine grid, m)

Elevation (m)

True width (m)

Gold grade (g/t)

Antimony grade

(%)

AuEq (g/t)

over 1.8 m or TW

Total hole

depth (m)

Target

BD244W4 14,763 5,692 1,011 0.98 6.5 1.4 4.8 254.7 Brunswick

BD249 14,792 5,764 1,102 2.51 13.3 0.1 13.6 170.2 Brunswick

BD250 14,796 5,779 1,078 2.88 2.5 0.1 2.7 194 Brunswick

BD252 14,774 5,714 1,067 1.57 2.3 0.2 2.3 189.9 Brunswick

BD254 14,775 5,716 1,041 0.40 17.0 29.3 14.8 210 Brunswick

BD255 14,788 5,760 1,039 1.36 8.3 8.7 17.5 220.7 Brunswick

BD256 14,786 5,767 1,012 0.29 4.0 2.4 1.3 241 Brunswick

BD260 14,828 5,920 1,064 0.16 3.4 0.0 0.3 190 Brunswick

BD264W1 14,765 5,736 995 0.48 15.6 4.3 6.1 260 Brunswick

BD265 14,792 5,835 969 0.66 19.5 5.5 10.6 476.2 Brunswick

BD265 14,760 5,833 938 10.35* 19.5 7.2 476.2 Brunswick

MH377 15,363 4,965 1,073 0.1 1.9 7.41 1.40 207.4 N Lode

MH382 15,349 4,870 1,004 0.3 102.5 38.20 33.11 126.3 N Lode

MH383 15,352 4,944 1,001 0.1 19.0 4.78 1.53 140.4 N Lode

MH384 15,348 4,849 1,016 0.5 52.3 6.28 19.86 137.8 N Lode

MH385 15,356 4,833 1,035 0.3 122.7 38.90 30.04 150.3 N Lode

MH386 15,352 4,810 1,049 0.2 39.9 12.80 8.07 166 N Lode

MH388 15,350 4,930 1,056 0.1 19.2 4.38 2.02 146.4 N Lode

MH392 15,362 4,978 1,055 0.2 161.7 3.28 19.00 188.5 N Lode

MH397A 15,376 5,038 1,,070 0.1 1.0 11.60 0.74 214.7 N Lode

MH398A 15,347 4,777 1,037 0.1 96.1 20.10 8.49 186 N Lode

MH399 15,352 4,907 1,054 0.2 1.6 0.71 0.39 143.8 N Lode

MH376 15,366 4,972 1,014 0.1 15.1 6.00 1.47 159.4 New Lode

MH377 15,387 4,982 1,084 0.1 11.0 16.40 2.49 207.4 New Lode

MH381A 15,357 4,902 991 0.3 6.3 3.41 2.39 130.8 New Lode

MH383 15,356 4,947 999 0.6 8.1 5.00 5.73 140.4 New Lode

MH387 15,376 4,968 1,046 0.1 28.3 0.69 2.05 160.8 New Lode

MH388 15,388 4,944 1,070 1.1 113.7 2.57 73.64 146.4 New Lode

MH392 15,375 4,988 1,060 0.1 10.7 6.67 1.46 188.5 New Lode

MH398A 15,358 4,759 1,,039 0.1 28.2 20.50 4.58 186 New Lode

MH399 15,386 4,909 1,066 0.7 13.1 3.45 7.75 143.8 New Lode

MH399 15,367 4,908 1,059 0.1 17.9 11.50 2.84 143.8 New Lode

MM014 15,124 3,010 1,041 0.14 2.99 0.03 0.23 330.2 Unnamed

MM014 15,113 3,011 1,033 0.23 0.05 7.00 1.77 330.2 Unnamed

MM015 15,234 2,801 1,094 0.07 3.95 0.00 0.16 270.5 Margaret East

MM015 15,177 2,801 1,050 0.82 2.62 0.01 1.21 270.5 Unnamed

BD243W1 14,728 5,891 846 0.230 24.20 10.40 5.83 748.8 Brunswick Deeps

BD243W1 14,728 5,891 846 2.250 2.70 3.10 8.99 748.8 Brunswick Deeps

BD242 14,787 5,820 978 0.617 3.31 2.22 2.67 457.6 Brunswick Main

Note: True width is a preliminary estimate only and may not reflect final true width used in resource estimation. AuEq (g/t) over 1.8 m or True Width (TW) if greater. **AuEq*g/t = Au(g) + Sb(%) x (Price per 10SB (kg) * Sb Recovery(%)/ Proce per 1 Au (g) * Au Recovery(%) ***Intercept on a splay vein proximal to the continuous lode structure.



10.9 Drilling Methods Due to the extensive historic drilling conducted throughout the Costerfield area, and the fact that mining has already depleted much of the Augusta Resource above 1,000 mRL, the following sections mainly relate to drilling completed after 1 January 2010 below 1,000 mRL.

The Augusta Deposit has been subject to ongoing development and diamond drilling since commencement of mining operations in 2006. The current Mineral Resource estimates are completed using all historic drilling data, and then depleted for areas already mined.

Between 2006 and 2011, several drilling companies were contracted to provide both surface and underground drilling services at Costerfield. To ensure consistent results and quality of drilling, Starwest Drilling Pty Ltd was made the preferred drilling services supplier in 2011.

Since 2011, underground diamond drilling has been completed using an LM90, which has drilled predominantly HQ-sized drill holes, with some NQ2-sized drilling completed as necessary to pass through areas of bad ground. Some underground drilling has been conducted using LTK46-sized core, but information from these drill holes was used purely for structural information. Surface drilling conducted to identify depth extensions to N Lode has been completed using HQ-sized core, switching to HQ3-sized core 50 - 100 m from the projected lode intercept.

Prior to 2011, various sizes and methods were used for drill holes. These included HQ, HQ3, NQ, NQ2, LTK60, LTK60, LTK48 and 5”1/8’ to 5”5/8’ RC. Details of these holes were not always recorded. However, because the majority of this drilling was in areas that are already depleted by mining, the risk associated with this drilling is perceived to be low. Figure 10-1 illustrates the known drill hole collar locations for both Augusta and Cuffley.



Figure 10-1: Known drill hole collar locations for Augusta and Cuffley including 2016 drilling



10.10 Collar Surveys Since 2006, drill hole collars have been surveyed according to the current Costerfield Mine Grid, either by Mandalay surveyors or by GWB Survey Pty Ltd. Between 2006 and 2011, Adrian Cummins & Associates provided surveying for both underground and surface collar locations.

Currently, initial collar locations are sighted using a hand-held GPS, with drilling azimuths provided by compass. Holes are then surveyed by Mandalay surveyors on completion.

Between the late 1990s and 2001, most drill holes appear to have been located using a GPS. Drill hole collar locations prior to the 1990s were usually sighted via tape and compass. Where possible, historic drill holes were surveyed in 2005 by Adrian Cummins & Associates, but this was not always possible.

Collars surveyed after 2001 are recorded in the acQuire drill hole database as being surveyed. Unsurveyed/ unknown drill holes are recorded as either GPS or unknown, and are given an accuracy within 1 m.

10.11 Downhole Surveys Since 2011, all holes have been downhole-surveyed using an electronic, single-shot survey tool. An initial check survey is completed at 15 m to ensure that the collar set-up is accurate. Thereafter, surveys are conducted at 30 m intervals, unless ground conditions are unsuitable to conduct a survey, in which case the survey is completed when suitable ground conditions are subsequently encountered.

Between 2001 and 2016, all drill holes were surveyed using either electronic single-shot or film single-shot survey tools. Prior to 2001, survey information exists for the majority of holes, but the method and records of these surveys are not readily available.

10.12 Data management In November 2016, Mandalay Resources exploration purchased the Geoscientific Information Management software, acQuire. The acquisition of this software package took place as drilling data collection was rapidly expanding, and this new software will improve overall efficiency in data collection and handling costs. It will improve the overall integrity of data and minimise human error in storing data. Prior to the purchase of acQuire, Excel spreadsheets and Access databases were utilised.

10.13 Logging Procedures The following information only relates to drilling completed after 1 January 2010, below 1,000 mRL, in the Augusta and Cuffley deposits.

Augusta core is geologically logged at the core preparation facility located at the Brunswick Processing Plant site. Core is initially brought to the facility by either the drill crews at the end of shift, or by field technicians who work in the core preparation facility. Core is generally stored on pallets while waiting for processing.

Field technicians initially orientate all core to the alignment provided by the drill crews using an electronic core orientation device. This orientation is transferred along the length of the run, with each drill run having been orientated. If a discrepancy is found between two adjacent runs, the next run is orientated and the two best-matching orientations are used. If no orientation is recorded by the drill crews, the core is simply rotated to a consistent alignment of bedding or cleavage, with no orientation mark made on the core.

Depth marks are made on the core at one-metre intervals using a tape measure, taking into account core loss and overdrill. If core loss is encountered, a block is placed in the zone of core loss.



If problems are encountered with driller core blocks, the drill shift supervisor is advised and depth marking stops until the problem is rectified.

Field technicians then collect rock quality designation (RQD) data directly onto a digital tablet device using acQuire software. RQD data is collected corresponding to drill runs, and includes “from” depth, “to” depth, run length in metres, recovered length in metres, recovery as a percentage, length of recovered core greater than or equal to 10 cm, and number of fractures. From this data, an RQD value is calculated. This data is logged ‘live’ into acQuire via a Toughbook computer, and synced to the company server.

Once depth marks are placed on the core, site geologists then log lithology, sample intervals, structural data, and geotechnical data (if applicable) directly into the acQuire database.

All measurements of structural features such as bedding, cleavage, faults, and shears are made using an orientated core wrap around protractor, and a protractor for alpha and beta measurements using the orientation line on the core. If no orientation line is available, only alpha measurements are made. Measurements are recorded directly into the acQuire database and are also inscribed on the core using wax pencil. As the logging is ‘live’, data is automatically backed up and stored on the company server.

After geological logging has been completed, all trays are photographed before sampling. Once sampling is complete, the trays are put on pallets and moved to a permanent core storage area.

10.14 Drilling Pattern and Quality

Augusta Drilling completed prior to 1 January 2010 informed areas of the resource that have largely been mined; therefore, the following discussion relates to drilling completed after 1 January 2010 and below 1,000 m RL.

Drilling is generally conducted with planned intercepts based on northing at an interval of approximately 40 m and 30 m up- and down-dip. Because most drilling at Augusta is now conducted from underground, and Cuffley Lode is the primary target of the majority of this drilling, the pattern and density achieved on N Main Lode can vary greatly. Where increased geological confidence is required, infill holes specifically targeting NE or E Lodes have been drilled at a nominal 40 m. Surface drilling, targeting depth extensions of the Augusta Deposit, is generally conducted on 100 m sections along strike, with intersections spaced 80 - 100 m up- and down-dip.

Cuffley Initial drilling of the Cuffley Lode was intended to be done in a ‘W’ pattern on an approximate 50 by 50 m offset grid. This pattern started with AD001 and went through to (including) AD004. To aid interpretation, this pattern was changed to a 100 m grid based on mine grid northings, with 50 - 80 m between holes on a given section. This pattern allowed better interpretation to be completed on sections.

For infill drilling between the 820 mRL and 1,020 mRL of the Cuffley Lode, the ‘W’ pattern was used to maximise strike direction information. This is as opposed to depth information, which is gained by drilling on section. This infill drilling was conducted on a nominal 30 m (RL) by 40 m (N) grid.

10.15 Interpretation of Drilling Results Drilling results are interpreted on paper cross sections. Interpretations are then scanned and registered into the mine planning software package Surpac. These sections are then used to interpret



wireframes between drill holes that are snapped to the drill hole in three-dimensional (3D) space. Figure 10-2 illustrates a typical cross-sectional interpretation completed for the Augusta deposit. Mappable stratigraphic units are represented with various colours, faults and lodes marked with heavy black lines.

Figure 10-2: Example of cross section at 4,300 mN post drilling geological interpretation of the Augusta deposit

Source: Mandalay, 2014.

10.16 Factors That Could Materially Impact Accuracy of Results The greatest factor that has the potential to materially impact the accuracy of results is core recovery. Historically, this was an issue for all methods of drilling in the Augusta area. Mandalay has employed methods of drilling and associated procedures to ensure the highest recovery possible. Where recovery is poor, a repeat hole is drilled via a wedge.

Information gained from historical drilling is still used in resource estimation. However, because much of the drilled area has already been depleted by mining, the associated risk is reduced significantly.

Surveying of the collar and downhole follows industry best practice1 given the location of drill collars and the expected deviation encountered during drilling. As such, the potential for significant impacts on results is minimised.

Sampling is also of a consistent and repeatable nature, with appropriate quality assurance/ quality control (QA/QC) methodologies employed. The assay method used is also considered appropriate for this style of mineralisation.

1 Survey techniques are consistent with the Canadian Institute of Mining and Metallurgy and Petroleum (CIM)

Exploration Best Practice Guidelines.



The nature of the geological information collected during the logging process is considered in line with industry best practice2 and consistent with other systems employed at deposits that are similar in nature, i.e. discrete, structurally controlled narrow vein deposits.

Drilling pattern and density are also considered appropriate, as section-based intersections allow for the best possible interpretation. Where required, infill drilling is utilised to gain additional information to assist in the interpretation and estimation processes.

2 Logging and interpretation techniques are consistent with the CIM Exploration Best Practice Guidelines and the

CIM Mineral Resource, Mineral Reserve Estimation Best Practice Guidelines.



11 Sample Preparation, Analyses, and Security 11.1 Sampling Techniques

Samples used to inform the Augusta and Cuffley block model estimates are sourced from both drill core and channel sampling along the ore development drives.

Diamond Core Sampling The mineralisation style at Costerfield is discrete; therefore, not all diamond drill core requires sampling. Sample intervals were determined and marked on the core by Mandalay geologists.

The following general rules apply to the selection of sample intervals:

• All stibnite-bearing veins are sampled.

• A waste sample is taken either side of the mineralised vein (ranging from 30 cm to 100 cm).

• Areas of stockwork veining are sampled.

• Laminated quartz veins are sampled.

• Massive quartz veins are sampled.

• Siltstone is sampled where disseminated arsenopyrite is prevalent.

• Puggy fault zones are sampled at the discretion of the geologist.

A Mandalay Exploration Field Technician samples the core. In order to obtain a consistent sample, the diamond drill core is cut in half with a diamond saw along the top or bottom mark of the orientated core.

Sampling intervals for drill core used for resource estimation purposes are no smaller than 3 cm in length and no greater than 1.0 m in length. The average sample length for drill core samples within the Augusta drill program for 2016 is 40 cm. Drill holes that were designed and drilled for metallurgical analysis have had sample intervals up to 2 m in length.

The northern side of the core is sampled to ensure that the same side of the core is sampled consistently. Where there is a definitive lithological contact that marks the boundary of a sample, the sample is cut along that contact. If, by doing this, the sample is less than 5 cm in length, the boundary of the sample is taken at a perpendicular distance from the centre of the sample, which achieves the 3 cm length requirement.

Samples from RC drilling are not used in the estimation of the Augusta Mineral Resource, although some RC and hammer drilling is used to establish pre-collars for deeper diamond drilling.

Underground Face Sampling Approximately 80% of all drive faces are sampled. Each development cut is approximately 1.8 m along strike. Samples are taken at a frequency between 1.8 and 5 m along strike. Underground face samples are collected using the following method:

• The face is marked out by the sampler to show the limits of the lode and the bedding angle; any geological structures that may offset the lode are marked on the face.

• Sample locations are marked out so that the sample is taken in a direction that is perpendicular to the dip of the lode from footwall to hanging wall.

• The face and length of the sample are measured.

• The face is labelled with the heading, dated and photographed.



• Each sample is collected as a channel sample using pick or pneumatic chisel and placed into a sample bag.

• Care is taken to obtain a representative sample.

• Where there are two or more lode structures in the face, samples of the intervening siltstone are also taken.

• Samples are between 0.5 kg and 2 kg in weight.

• The sample length can vary from 5 cm to 1.5 m across the structures.

• Each sample bag is assigned a unique sample ID.

• A sample book is used to record the sample IDs; a ticket stub from the sample book is placed into the sample bag for use by the assay laboratory.

• The face is sketched on a face sample sheet and sample details are recorded.

• The location of the face is derived from survey pickups of the floor and backs of the ore drive.

• Face samples are taken at the appropriate orientation to the mineralisation and are representative of the mine heading being sampled.

11.2 Data Spacing and Distribution Within the Augusta and Cuffley deposits, the distance between drill hole intercepts is approximately 40 by 30 m. This is reduced to 20 by 20 m in areas of structural complexity. Face sampling along drives is done at a frequency of 1.8 - 5.0 m along strike and 5 - 10 m down dip.

11.3 Testing Laboratories Assaying of the drill core and face samples is predominantly completed by Onsite Laboratory Services (Onsite) in Bendigo. This laboratory is independent of Mandalay and holds a current ISO 9001 accreditation. Genalysis (Brisbane and Perth), ALS Global (ALS) in Brisbane and Bureau Veritas (BV) in Perth have also been used to verify the accuracy of Onsite’s testing.

After Mandalay dispatches the core or face samples, the assaying laboratory’s personnel undertake sample preparation and chemical analysis. Results are returned to Mandalay staff, who validate and input the data into the relevant databases.

11.4 Sample Preparation The following sample preparation activities are undertaken by Mandalay staff:

• Sample material is placed into a calico bag pre-marked with a sample number.

• Sample characteristics are marked on a sample ticket stub and placed in the bag.

• Calico bags are loaded into plastic bags which are not to weigh more than 10 kg.

• An assay request sheet is completed and placed in the sample bag.

Plastic bags containing samples are sealed and transported to Onsite in Bendigo via private courier.

The following sample preparation activities are undertaken by Onsite staff:

• Samples are received and checked against the submission sheet.

• A job number is assigned and worksheets and sample bags are prepared.

• Samples are placed in an oven and dried overnight at 80°C.



• The entire sample (up to 3 kg) is jaw-crushed to approximately 2 mm – if the sample weighs less than 3 kg, the sample is split and 50% of the sample is used.

• The entire sample is milled and pulverised to 90% passing 75 μm.

• Samples are split, with 200 g for analysis and the remaining sample returned to its sample bag for storage and eventual return to Mandalay.

11.5 Sample Analysis Augusta drill core and face samples are assayed for gold, antimony, arsenic and iron. Gold grades are determined by fire assay/ AAS. The following procedure is undertaken by Onsite for gold and antimony:

• 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.

• Once the solids have settled, the gold content is determined by flame AAS.

Antimony grades are determined using acid digest/ AAS. Where the sample contains antimony in excess of 0.6% concentration, the following procedure is undertaken:

• 0.2 g of sample is added to a flask of distilled water (20 mL).

• 30 mL of 50% nitric acid is added.

• 20 mL of tartaric acid is added.

• 80 mL of 50% HCl is added and allowed to stand for 40 minutes.

• 5 mL of hydrobromic acid (HBr) is added.

• The solution is mixed for one hour and left to stand overnight until fuming ceases.

• Sample is heated until colour changes to light yellow and white precipitate dissolves.

• When cool, the sample is diluted to 200 mL with distilled water.

• Antimony content is determined by AAS.

11.6 Laboratory Reviews Mandalay personnel conduct periodic visits to the Onsite laboratory in Bendigo and meet regularly with the laboratory managers. In 2016, monthly visits were scheduled at the Onsite lab and quarterly visits at the Costerfield office.

Tours of the laboratory are normally completed in the presence of Onsite’s Laboratory Manager, Mr Rob Robinson.

Notes and minutes from laboratory visits and meetings with laboratory staff are stored on the Mandalay server.

Mandalay conducted check assay programs in May and November 2016. This process involves obtaining a pulped sample from Onsite, splitting it, and submitting pulps to three laboratories for comparison. In 2016, the same samples were sent to Onsite, ALS and Bureau Veritas. The results of the most recent program (November 2016) are outlined in Section 11.7.4.



11.7 Assay Quality Assurance and Quality Control

Standard Reference Material In total, five standards were used for quality control in 2016. Two were made from material collected from Augusta (MR04/06 and MR30/20), and routinely submitted to Onsite. Both MR04/06 and MR30/20 have antimony and gold only standards. Figure 11-1 through Figure 11-5 present the results.

Mandalay also routinely uses a commercially available, gold-only standard (G310-6) sourced from Geostats Pty Ltd and provided to Onsite.

A standard is sent with each batch of exploration samples (on average 1 standard per 25 samples) and with each batch of underground face samples (on average 1 standard per 10 samples).

Results from January 2016 to December 2016 for gold and antimony are displayed in Figure 11-1 through Figure 11-11. A standard assay result is considered to be compliant when it falls inside the three standard deviation (SD) limits defined by the standard certification. When a batch fails to comply with the three SD limits defined by the standard certification, all significant assay results from that batch are re-assayed. Significant assay results are defined as samples that may be used in a future resource estimate. Any actions or outcomes are recorded as comments on the QA/QC database.

A review of the results shows the following:

• Gold (and antimony) standard MR30/20 (CRM Assay 23.64 g/t gold) displays excellent compliance for the period, with no assays outside the ±3 SD limits. There is no evidence of bias and drift in the accuracy of the assay results for the period (Figure 11-1).

• Gold standard 310-6 (CRM 0.61 g/t Au) displayed in Figure 11-2 shows good compliance for 2016 with one outlier outside the ±3 SD limits, which is likely due to a typing error.

• Gold (and antimony) standard MR04/06 (CRM Assay 7.51 g/t gold) shows excellent compliance for the period, with no assays outside the ±3 SD limits. There is no evidence of any significant drift in the accuracy of the assay results for the period (Figure 11-3).

• Antimony standard MR30/20 shown in Figure 11-10 (CRM Assay 31.08% antimony) shows good compliance for the period, with one assay outside the ±3 SD limits. There is no evidence of bias or any significant drift in the accuracy of the assay results for the period.

• Antimony standard MR04/06 displayed in Figure 11-11 (CRM Assay 3.50% antimony) shows good compliance for the period, with the exception of one assay outside the ±3 SD limits. There is no evidence of bias or any significant drift in the accuracy of the assay results for the period; and

• Mandalay considers that the level of compliance and bias displayed by the standards is good and demonstrates the reliability of the gold and antimony grades used to inform the block model estimate.



Figure 11-1: MR30-20 Gold Standard Reference Material – Assay Results for 2016

Figure 11-2: G310-6 Gold Standard Reference Material – Assay Results for 2016



Figure 11-3: MR04/6 Gold Standard Reference Material – Assay Results for 2016

Figure 11-4: MR30/20 Sb Standard Reference Material – Assay Results for 2016



Figure 11-5: MR04/6 Antimony Standard Reference Material – Assay Results for 2016

Blank Material Mandalay sends uncrushed samples of basalt as blank material (1 in every 30 samples) to Onsite to test for contamination. Greater than two times the detection limit is regarded as an unacceptable assay on blank material. The detection limit for gold at Onsite is >0.02 g/t.

Figure 11-12 and Figure 11-13 show the performance of the blanks for 2016. This data is a combination of all mine data and exploration drill hole data and provides a good representation of the blank values received from Onsite. The 2016 results display some exceedance of a handful of samples over 0.05, but the majority of samples lie within the ±3 SD limits. The variability of results can likely be attributed to a small amount of contamination. In comparison to previous years, it appears that the regular use of quartz wash seems to keep the blank readings relatively low. Given the high gold grades at the mine, it is not considered by Mandalay and SRK to be an issue that will have any material impact on the Mineral Resource estimate.

It also became apparent that the high blank readings were, in some cases, typing errors at the laboratory.

Figure 11-6: Au Blank assay results for 2016



Figure 11-7: Sb Blank assay results for 2016

Duplicate Assay Statistics A summary of laboratory duplicate statistics assayed by Onsite, for original gold assays versus duplicate gold assays for 2016, is presented in Table 11-1. The duplicates are assayed on separate aliquots of the same sample pulp. A scatter plot of this data is presented in Figure 11-8, which shows no significant bias between the original and duplicate assays.

A relative paired difference (RPD) plot using the same duplicate dataset is presented in Figure 11-9. It is desirable to achieve 90% of pairs at less than 10% RPD in the same batch, or less than 20% in different batches or different laboratories (Stoker, 2006). The duplicate dataset achieves 75% of pairs at less than 10% RPD and 97% of pairs at less than 20% RPD, which demonstrates acceptable precision in the gold assays by Onsite.

Table 11-1: Summary of Onsite gold duplicate statistics

Statistic Original Duplicate

Number of samples 302 302

Mean 45.24 45.61

Maximum 610.20 668.00

Minimum 0.02 0.03

Population Std Dev. 72.66 75.55

CV 1.61 1.66

Bias -0.80%

Correlation Coefficient 0.99

Percent of samples < 10% RPD 74.83



Figure 11-8: Scatter plot for Onsite gold duplicates (g/t) for 2016

Figure 11-9: Relative paired difference plot for Onsite gold duplicates (g/t) in 2016

Check Assay Program – Sample Pulps Duplicate statistics resulting from a gold and antimony check assay programme comparing Onsite, ALS and BV are presented in Table 11-2 and Table 11-3. The duplicates are assayed on separate aliquots of the same sample pulp. Two pulp check assay programs were conducted – in May and November 2016. The results displayed in this report are from the most recent program undertaken by Mandalay in November 2016.

The gold scatter plot of this data (Figure 11-13) shows a large scatter for samples greater than 80 g/t, with the high grades from ALS and BV returning higher grades than Onsite. The gold RPD plot shows that, on average, 88% of duplicate pairs show less than 20% RPD (Table 11-2). On further investigation it was found that ALS and BV did not use the same assay techniques. In addition, because of the large range of gold values, a number of different techniques are utilised depending on



the grade range. ALS’s standard method for assays greater than 100 g/t is GRA21; however, laboratory reports show a method, titled DIL25, is used for samples greater than 100 g/t. This issue is currently being investigated further to determine whether the ALS samples greater than 100 g/t may have had an incorrect or inappropriate method applied, resulting in an under-call of the true assay value. The grade comparisons below 80 g/t show no significant bias.

The antimony statistics show a bias of an average 3.14%, with Onsite reporting lower antimony grades than both ALS and BV. The antimony RPD plot shows that approximately 98.6% of duplicate pairs are less than 20% RPD (Figure 11-11), which demonstrates good reproducibility across laboratories. However, Onsite has been reporting slightly lower antimony results, which may be due to an older atomic absorption method being used. The reproducibility between laboratories (for antimony for the 2016 check assay program) has markedly improved, from 91.3% RPD in 2015 to 98.6% RPD in 2016.

Both ALS and BV reported higher gold grades from Onsite original pulp. Higher gold grades reported by ALS may not be representative. An investigation of this issue in early 2016 revealed that when ALS has been used in certification programs for CRMs, there is a small, but clear, high bias. Communication with Geostats Pty Ltd also confirms that ALS has shown a high bias when assessed in the Geostats round robin of laboratories; this is more pronounced at higher grade ranges.

Mandalay considers this issue important, and it is the focus of ongoing investigation in order to quantify its potential impact on the resource.

Table 11-2: Summary of Onsite original vs Onsite duplicate, ALS, Bureau Veritas gold duplicate statistics

Statistic Onsite Check ALS Bureau Veritas

Number of samples 50 50 50 50

Mean 47.74 46.27 49.16 47.31 Maximum 242.10 242.80 175.00 234.00

Minimum 1.45 1.71 1.70 1.88

Population Std Dev. 45.25 44.42 40.47 44.01

CV 0.95 0.96 0.82 0.93 Bias 3.09% -2.96% 0.90% Correlation Coefficient 1.00 0.90 0.99 Percent of samples < 20% RPD 98.00 86.00 80

Table 11-3: Summary of Onsite original vs Onsite duplicate, ALS, Bureau Veritas antimony duplicate statistics

Statistic Onsite Check ALS Bureau Veritas

Number of samples 50 50 50 50 Mean 28.44 28.34 27.27 26.96 Maximum 60.20 61.90 66.90 64.7 Minimum 4.42 4.26 4.53 4.49 Population Std Dev. 15.92 16.07 15.74 15.79 CV 0.56 0.57 0.58 0.59 Bias 0.33% 4.11% 5%

Correlation Coefficient 0.99 0.99 0.96

Percent of samples < 20% RPD 100.00 100.00 96



Figure 11-10: Relative pair difference plot for Onsite Original vs Onsite duplicate, ALS, BV gold duplicates (g/t)

Figure 11-11: Relative pair difference plot for Onsite original vs Onsite duplicate, ALS, BV antimony duplicates (%) with BV outliers



Figure 11-12: Scatter plot for Onsite original vs Onsite duplicate, ALS, BV antimony duplicates (%)

Figure 11-13: Scatter plot for Onsite original vs Onsite duplicate, ALS, BV gold duplicates (g/t)

11.8 Sample Transport and Security Sample bags containing sample material and a ticket stub with a unique identifier are placed in heavy duty plastic bags in which the sample submission sheet is also included. The plastic bags are sealed with a metal twisting wire. This occurs for both underground face samples and drill core samples. The bags are taken to a storage area that is under constant surveillance. A private courier collects samples twice daily and transports them directly to Onsite in Bendigo, where they are accepted by laboratory personnel. Sample pulps from Onsite are returned to Mandalay for storage. The pulps are stored undercover and wrapped in plastic.



12 Data Verification On 18 November 2014, SRK full-time employee Danny Kentwell (QP for Sections 6 - 12 and Section 14) visited the Augusta and Brunswick mine sites, escorted by Chris Davis, Resource and Exploration Manager for Costerfield Operations. All drill core for the Costerfield property is processed at the Brunswick exploration core shed. Discussions with site geologists were held regarding the following aspects, for data verification:

• Sample collection

• Sample preparation

• Core mark-up

• Core recovery

• Core cutting procedures

• Sample storage

• QA/QC

• Data validation procedures

• Collar survey procedures

• Downhole survey procedures

• Geological interpretation

• Exploration strategy

• Grade control sampling and systems

• Inspection of Brunswick core shed facilities and drill core intersections (Augusta and Cuffley).

An underground tour was conducted. The Cuffley Lode was observed in strike drives on the lode at two adjacent levels. The Cuffley East Lode was also observed via a strike drive on the lode. Obvious stibnite mineralisation and visible gold was sighted in both the faces (Figure 12-1) and in rock fragments along the drives.



Figure 12-1: Cuffley Main lower drive south end showing mineralised structures Note: gold bearing quartz = pink, stibnite = yellow and bedding

Danny Kentwell also visited the site in August 2015 and November 2016 to review current operations, but did not go underground.

SRK assisted Mandalay with mid-year modelling updates and, as part of this process, examined the underground face sampling database. A number of inconsistencies between original paper face mapping and sampling logs and the final details in the electronic database were noted. These were communicated to Mandalay along with recommendations to make the transfer of logging and sampling data information a more robust process. During December 2016, SRK re-examined the face sampling database and found that inconsistencies still existed in some areas. As part of Mandalay’s QA/QC procedures, they examine all face samples before use for resource modelling; if inconsistencies are noted they are either resolved or the sample is rejected from use in resource estimation. The level of face sampling initially found to be inconsistent and requiring verification is not recorded. The level of rejections (unresolved inconsistencies) is approximately 3% of the face sampling data. SRK considers this acceptable. SRK has recommended further revision of the data transfer process to ensure that maximum use is made of the effort involved in the face sampling process.

In SRK’s opinion, the geological data used to inform the Augusta and Cuffley block model estimates were collected in line with industry best practice as defined in the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Exploration Best Practice Guidelines and the CIM Mineral Resource, Mineral Reserve Best Practice Guidelines. As such, the data are suitable for use in the estimation of Mineral Resources.



13 Mineral Processing and Metallurgical Testing 13.1 Metallurgical Testing

Extensive ongoing metallurgical testwork has been undertaken on samples from the Augusta deposit since 2004 and Cuffley deposit since 2012. The Brunswick deposit, currently classified as a resource but potentially a feed source in a future Life of Mine (LOM) extension, has undergone testing from 2016 onwards. Routine mill feed blend characterisation tests and metallurgical tests are ongoing.

The following reputable, appropriately experienced laboratories were involved with various aspects of the original metallurgical evaluation:

• ALS Ammtec – New South Wales (previously Metcon Laboratories) (Metcon)

• AMDEL Ltd Mineral Services Laboratory – South Australia (Amdel)

• Australian Minmet Metallurgical Laboratories – New South Wales (AMML).

The metallurgical testwork on Augusta and Cuffley ore has since been superseded by operational data. Using comprehensive historical operating data is a more accurate way to forecast future metallurgical behaviour when processing similar ores, rather than using old and limited testwork data. The Brunswick Processing Plant has been operated by Mandalay since late 2009, and three years of operating data exists for the current Cuffley/ Augusta ore blend. These are the same two deposits being treated in the forward LOM plan, so this provides appropriate reference data.

This allows relationships developed from historical operating data to be used to forecast future throughput and recoveries, as the ratio of ores is scheduled to continue unchanged for the LOM. These relationships account for other influencing variables such as feed and concentrate grades. This provides a much better understanding of the processing behaviour expected on these and similar ores.

Brunswick Testwork Metallurgical testwork has been undertaken on the Brunswick Deposit, which may provide future plant feed. It is not yet included in the LOM production schedule. This testwork has shown that the Brunswick ores will demonstrate similar metallurgical behaviour to the current ores feeding the plant. Metal recoveries were high and reflected historical plant performance. It is expected that, with further optimisation of the testwork conditions, recoveries can be further increased.

Only sighter level physical (comminution) testing has been undertaken. The single Bond Ball Mill Work Index (BBMWi) test resulted in a lower value (12.9 kWh/t) than the ore that is currently processed (Cuffley at 16.0 kWh/t and Augusta at 15.5 kWh/t). Flotation testing has shown the recoveries to be relatively insensitive to a grind size between 38 and 75 microns. Based on this testing, it is likely that the plant throughput will be maintained, or potentially marginally increased, when feeding a blend including a Brunswick ore component.

Ten flotation tests have been undertaken on a Brunswick composite sample. The sample had a feed grade of 8.65 g/t gold and 3.31% antimony. Tests consistently achieved an antimony recovery above 98% and gold recovery at 87% once test conditions were refined and assuming a recycle of the cleaner tails. The later results were stable across a range of grind sizes and reagent addition regimes. The best results from Test 7 achieved a recovery of 87.1% and 98.3% for Au and Sb respectively, even after disregarding the cleaner tailings, which in practice will be recycled on the plant.

SRK notes that the arsenic grades in the Brunswick feed and concentrate products are higher than in the current feed blend. The recovery of arsenic was approximately 90% in the same testing, with a feed grade of 0.5% and a resultant concentrate grade of approximately 3.2%. This is higher than the current arsenic feed grades of 0.06%. If Brunswick were to be brought into production, there would



be a need to blend either feed and/ or concentrate products, to manage the arsenic to target levels. Future consideration will need to be given to whether the optimum processing conditions are sufficiently similar that the different ores could be processed as a blend or whether they would need to be campaign processed through the processing plant. If the latter is the case, more emphasis would be required on blending the final products to ensure a homogenous product arsenic grade while maintaining consistent product export to sustain cash flow to the operation. Currently any shipments over the target of 3.0% will incur modest penalty payments, but this is not considered a risk to the project.

SRK understands that additional testing will be undertaken in the future, as additional Brunswick samples become available, to confirm their metallurgical behaviours and develop grade and recovery relationships and arsenic product grades, to further de-risk the introduction of the Brunswick ores into the LOM plan and help develop future blending strategies.

13.2 Ore Blend Effect on Throughput and Recovery Forecasts Since January 2014, Cuffley ores have been processed in a blend with Augusta ores (previously only Augusta ore was processed). The following ratios of Augusta and Cuffley ores have been processed historically, and a blend will be continued to be processed:

• 2014: 44% Augusta and 56% Cuffley



• LOM 2017 onwards: 32% Augusta, 68% Cuffley.

Therefore, throughput and recovery data from 2014 - 2016 is reflective of that expected from 2017 onwards, based on the equivalent ore blend to be processed in the forward LOM. This assumes ores of similar lithology and oxidation state operating under similar processing conditions and feed grades.

13.3 Throughput The forecast throughput used for LOM planning is based on historical throughput. This is the best method of forecasting throughput on a similar ore blend. The historical throughput (2014 - 2016) used for forecasting is comparable to the feed blend of Augusta and Cuffley ores used in the forward LOM plan.

Several processing plant upgrades have seen production throughput increase from approximately 2,000 t/mth in 2007 to approximately 10,000 t/mth in 2012. Since then, through ongoing optimisation and minor, low capital cost debottlenecking projects, the capacity has been increased to the current 2017 capacity, which can exceed 13,000 t/mth. Historic throughput can be seen in Figure 13-1 and Figure 13-2.



Figure 13-1: Historic Brunswick Processing Plant throughput: Apr 2007 to Dec 2016

Figure 13-2: Historic Brunswick Processing Plant throughput: Jan 2012 to Dec 2016

Although throughput began to plateau through 2014 - 2016, as shown in Figure 13-3, there was still a modest increase from an average 12,445 t/mth in 2014 to 12,822 t/mth in 2015, and then to 12,867 t/mth in 2016. The 2016 monthly throughput was regularly over 13,000 t, and as high as 13,762 t. It is expected that ongoing process optimisation and minor, low capital cost debottlenecking projects will continue the ongoing production increase. This is considered to be the project upside. It is not incorporated into the budgeted or LOM model throughput or recovery.



Figure 13-3: Historic Brunswick Processing Plant throughput: Jan 2014 to Dec 2016

Forecast throughput for the 2017 LOM is 13,000 t/mth, based on current mining rates. Plant budgeted throughput is 12,833 t/mth. The average forecast throughput for the remaining LOM to 2020 is 13,000 t/mth, before it tapers off after mid-2020. Based on the 2014 and 2015 throughput, the LOM and budgeted milling forecast is supported and there is potentially upside in plant throughput.

The Brunswick mill currently operates with approximately 20,000 tonnes of ROM stockpile inventory allowing for fluctuations in mining production and/or a further increase in plant throughput. There is further capacity on the existing ROM pad if required. Remaining ROM stockpiles will be processed at the end of the LOM.

13.4 Recovery Forecast antimony and gold recoveries used for LOM planning and economic modelling are based on feed grades, historical recoveries and concentrate grades. This is the best method for forecasting throughput on the same ore blend. The historical recoveries (2014 - 2015) used for forecasting are the equivalent of the feed blend of Augusta and Cuffley ores used in the forward LOM.

Grade vs Recovery Figure 13-4 shows the grade versus recovery relationship from January 2014 to December 2016. There is a relationship between head grade and recovery for both gold and antimony. This is a common phenomenon based on a (relatively) constant tail, and one that is expected. The 2015 gold grade versus recovery data is stronger than the 2014 data, due to operational changes. Development of specific relationships used for forecasting recovery is discussed further below.



Figure 13-4: Cuffley/ Augusta blend feed grade vs recoveries: 2014 - 2016

Antimony Recovery Using the 2015 daily operating data, an antimony recovery relationship was developed based on the interaction between the feed head grade, concentrate grade and recovery. This relationship is shown in Figure 13-5.

Figure 13-5: Antimony yield to concentrate vs upgrade for Cuffley/ Augusta ore blend: 2015

Where:

Yield to concentrate = Sb head grade (%) x Sb recovery (%)/ Concentrate Sb grade (%)

Upgrade = Concentrate Sb grade (%)/ Sb head grade (%)

The resultant relationship is shown below:

Yield to concentrate = 110.52 * (Upgrade)-1.056

The relationship that has been developed has an extremely high correlation coefficient of 0.999, demonstrating a high degree of confidence in the relationship even though there are other processing



variables such as grind size, plant stability, reagent addition, fluctuations in feed blend and minor changes to lithology, oxidation state and mineralogy that would also be expected to impact on the relationship.

Because of the confidence in this relationship across a range of feed grades, it provides the best means of estimating the antimony recovery at variable head grades. This relationship assumes a constant final antimony concentrate grade of 54%, which is consistently what the processing plant produces.

The data for 2016 has been assessed in the same model and supports the ongoing use of the formula, with a high correlation coefficient maintained. This formula is used in the LOM model to predict the Sb recovery for a given head grade for the remaining LOM. SRK notes that the average Sb head grade is lower for the remaining LOM; however, it is still strongly supported by the model. The recent historical and forecast Sb mass weighted recoveries for the LOM are:

• LOM model (2017 - 2021) = 92.9% Sb recovery at a 2.8% Sb feed grade

• 2016 actual Sb recovery = 95.4% at a 3.7% Sb feed grade

• 2015 actual Sb recovery = 95.1% at a 4.0% Sb feed grade.

Gold Recovery Using the daily 2015 operating data, a similar gold recovery relationship was developed for the gold reporting to the antimony and gold concentrate, based on the interaction between the flotation feed grade, concentrate grade and recovery. It too has a very high correlation coefficient. However, the flotation recovery only makes up part of the overall gold recovery, as 30% - 40% typically reports to the gravity gold concentrate. The gravity gold recovery has a relatively high level of variability, complicating the application of the flotation gold recovery relationship to forecast the overall gold recovery.

Overall gold recovery appears to be quite independent of the gravity recovery, i.e. the flotation circuit is efficient at recovering what the gravity gold circuit is not recovering. This was demonstrated in 2015, when a drop in gravity recovery that corresponded with an increase in feed grade (likely due to changes in mineralogy and/or finer gold, as well as low grade ore being fed into the blend) was offset by higher flotation gold recovery.

A number of alternative gold recovery relationships have been assessed. A strong relationship is evident for gold feed grade versus tailings grade, particularly in the 2016 calendar year. The tailings grade is relatively consistent, ranging between 0.81 g/t and 1.28 g/t on a monthly basis, depending on the gold feed grade. Based on the relationship between reconciled monthly feed and tails gold grades for 2015, a relationship was formulated to calculate gold recovery for the LOM model. This relationship was used for the 2015 and 2016 LOM forecasting.

Gold recovery (%) = ((head grade-(0.0197*head grade+0.9647))/head grade)*100.

This exercise was repeated for the 2016 calendar year, resulting in a similar relationship with a stronger correlation coefficient of 0.8. The revised 2016 relationship and a combined 2015 - 2016 relationship were both compared against the 2015 gold feed grade versus tailings grade relationship. The 2015 relationship is more conservative, discounting the recovery by approximately 1.5% in absolute terms, and has been used for the 2017 financial modelling. This allows for any variability in the recoveries during future months of scheduled lower gold feed grade that are below monthly historical feed grades.



A fixed gravity gold component of 36.5% was then subtracted from the overall recovery to allow the gravity and flotation gold recoveries to be separated. This value was also based on historic data: 37.4% between 2014 and 2016, and 35.7% in the 2016 calendar year only.

The 2015 and 2016 historical gold recovery data used to forecast the LOM recovery was as follows:

• 2015 actual – Total gold recovery of 89.8% and gravity recovery of 34.0% on 10.7 g/t head grade (resultant tailings grade of 1.17 g/t and flotation recovery of 55.8%)

• 2016 actual – Total gold recovery of 90.1% and gravity recovery of 35.7% on 10.3 g/t head grade (resultant tailings grade of 1.08g/t and flotation recovery of 54.3%).

LOM model forecast gold recoveries weighted averages are:

• LOM model (2017 - 2021) = 81.2% on 6.38 g/t head grade

• LOM model (2017 - 2021) fixed gravity recovery assumption of 36.5%.

The overall gold recovery and breakdown into gravity and flotation recoveries is suitably supported by historical operating data. The recovery relationships will be updated during the next review and applied to the LOM planning and LOM financial modelling.

Throughput Effect on Recovery

Figure 13-6 shows that the mill throughput was relatively consistent from 2014 to 2016, with a slight increase in the year on year average recoveries due to incremental (continuous) improvements and efficiency creep. Based on this relatively stable flow data, and no clear recovery relationship existing within the throughput range of the operation, throughput has been assumed to have a negligible effect on the mill recovery calculations at the forecast LOM throughput (a minor exception possibly being October 2016; nevertheless, the correlation is not strong over the entire period).

Figure 13-6: Mill recoveries and throughput for Cuffley/ Augusta ore blend: 2014 – 2016



Table 13-1: Brunswick Processing Plant key processing results – January 2014 to December 2016 Jan-14 Feb-14 Mar-14 Apr-14 May-14 Jun-14 Jul-14 Aug-14 Sep-14 Oct-14 Nov-14 Dec-14 2014

Milled (DMT)* 12,783 11,377 13,042 11,756 12,739 11,593 13,023 13,618 12,376 12,788 12,345 11,897 12,445 Antimony recovery (%) 94.0 94.1 94.5 94.3 95.8 94.1 93.7 94.3 95.3 94.9 94.2 95.1 94.6 Gold recovery (%) 91.0 90.5 90.1 88.6 88.6 88.0 87.7 87.1 90.6 91.4 90.0 91.6 89.9 Antimony feed grade (%) 4.0 3.7 3.3 3.3 4.3 3.8 3.8 4.1 5.0 4.4 3.9 4.1 4.0 Gold feed grade (g/t) 9.1 7.9 7.9 8.3 8.2 8.6 9.1 9.2 11.8 11.7 10.7 12.1 9.5 Concentrate (DMT) 877 750 770 705 985 784 878 1002 1089 975 831 854 875 Ave. Conc Sb grade (%) 54.3 53.0 52.4 52.4 53.9 52.8 52.5 52.6 54.6 54.4 54.4 54.6 53.5

Jan-15 Feb-15 Mar-15 Apr-15 May-15 Jun-15 Jul-15 Aug-15 Sep-15 Oct-15 Nov-15 Dec-15 2015 Milled (DMT) 13,629 11,702 13,271 12,529 13,106 12,349 12,643 13,359 13,139 13,564 12,453 12,127 12,822 Antimony recovery (%) 94.8 94.8 95.3 94.9 94.2 95.8 95.5 95.5 95.0 95.0 94.6 96.0 95.1 Gold recovery (%) 89.8 90.1 89.8 89.7 88.0 89.3 90.2 89.5 89.7 89.5 90.2 90.9 89.9 Antimony feed grade (%) 3.9 3.9 4.3 3.8 3.3 4.3 4.1 4.2 3.9 3.9 3.8 4.3 4.0 Gold feed grade (g/t) 10.1 11.2 10.6 10.0 9.2 10.4 11.0 10.5 11.1 10.8 11.3 11.7 10.7 Concentrate (DMT) 922 799 983 836 752 928 916 978 900 950 835 931 894 Ave. Conc Sb grade (%) 54.1 54.3 54.8 54.4 54.1 54.3 53.6 54.2 54.2 53.1 53.3 53.8 54.0

Jan-16 Feb-16 Mar-16 Apr-16 May-16 Jun-16 Jul-16 Aug-16 Sep-16 Oct-16 Nov-16 Dec-16 2016 Milled (DMT) 12,913 12,960 13,762 13,749 13,370 12,430 12,194 12,113 11,674 12,886 13,284 13,074 12,867 Antimony recovery (%) 95.6 95.4 95.1 95.1 95.6 94.5 95.1 95.2 95.7 95.4 96.0 96.4 95.4 Gold recovery (%) 90.5 91.3 90.1 90.8 90.6 90.8 89.4 90.7 90.1 87.8 89.0 89.1 90.1 Antimony feed grade (%) 4.1 4.1 4.0 3.9 3.8 3.8 3.8 3.8 3.7 3.8 3.4 2.8 3.7 Gold feed grade (g/t) 11.3 12.4 11.7 12.3 10.8 11.6 9.8 10.4 9.6 7.9 7.8 7.4 10.3 Con DMT 927 943 988 941 899 829 817 775 743 839 802 685 849 Ave. Conc Sb grade (%)) 54.6 54.1 53.2 53.9 54.7 53.4 53.6 56.3 55.5 55.1 53.4 52.0 54.1

*DMT = dry metric tonnes



Figure 13-7: Mill recoveries and throughput 2016

Figure 13-8: Feed grade and recoveries 2016



14 Mineral Resource Estimates 14.1 Introduction

Gold and antimony grades and lode thickness were estimated using the two dimensional (2D) accumulation method. The 2D accumulation method requires that gold and antimony are multiplied by true thickness to give a gold accumulation and antimony accumulation. This method assigns weights to composites of different lengths during estimation. The interpolation method used was ordinary kriging with the exception of Cuffley Main Domain 6, Brunswick PK, CSE 213, CSE 212, Sub King Cobra and NSP lodes which used inverse distance squared. The estimated grade was then back-calculated by dividing estimated gold accumulation and estimated antimony accumulation by estimated true thickness.

Not all models have been re-estimated due to the absence of new data being captured during 2016 within these areas. Table 14-1 is a summary of the changes made to the models within this estimate.

Table 14-1: Changes made to models at year-end 2016

Lode Zone Code New data captured

during 2016 New

Estimation New Resource Classification

Depleted during 2016

Reported above new

cut-off

E Lode 10 No No No Yes Yes

B Lode 15 Yes Yes Yes Yes Yes

BSP Lode 16 Yes Yes Yes Yes Yes

W Lode 20 No No No Yes Yes

NM Lode 40 Yes Yes No Yes Yes

NE Lode 41 Yes Yes No Yes Yes

NV Lode 43,44 Yes Yes Yes Yes Yes

NSW Lode 45 No No No Yes Yes

NSP 1 Lode 49 Yes Yes Yes Yes Yes

NSP 2 Lode 38 Yes Yes Yes No Yes


NSP 4 Lode 39 Yes Yes Yes Yes Yes


NW Lode 47 No No No Yes Yes

P1 Lode 55 No No No Yes Yes

P2 Lode 56 No No No Yes Yes

K Lode 60 Yes Yes Yes Yes Yes

C Lode 30 Yes Yes Yes Yes Yes

CM Lode 210 Yes Yes No Yes Yes

CE Lode 211 Yes Yes No Yes Yes

CI Lode 215/216/217 No No No No Yes

CD Lode 220 Yes Yes Yes No Yes

CDL Lode 225 Yes Yes Yes No Yes

CS Lode 212 Yes Yes Yes No Yes

CSE Lode 213 Yes Yes Yes No Yes

AS Lode 230 No No No Yes Yes



Lode Zone Code New data captured

during 2016 New

Estimation New Resource Classification

Depleted during 2016

Reported above new

cut-off

Brunswick 300 Yes Yes No Yes Yes

Brunswick PK 310 Yes Yes Yes No Yes

SKC C 410 Yes Yes Yes No Yes

SKC CE 400 Yes Yes Yes No Yes

SKC LQ 405 Yes Yes Yes No Yes

SKC W 420 Yes Yes Yes No Yes

For sample statistics, top cutting and estimation parameters of models not updated during this estimation, the 2015 NI 43-101 report (SRK, 2015) can be referenced. From here on, the models not re-estimated at 2015 year-end are called “2015 models” (E, W, NSW, NW, P1, P2, CI and AS Lodes).

14.2 Diamond Drill Hole and Underground Face Sample Statistics Statistics for gold and antimony grades and true thickness for newly re-estimated lodes are presented in Table 14-2.

The tabulated data shows the unweighted average gold and antimony grade being higher within the face sample data than the drill holes. This is attributed to two factors:

1). Face sample data is collected representatively within ore drives however these ore drives exist only in areas of the deposit that are deemed economically viable. Therefore the average grade of these samples is expected to be higher than that of the drilling data which includes intercepts within areas that will never be mined as they are sub-economic.

2). When drill core is sampled the core is cut at an angle perpendicular to the long axis of the core rather than along the boundary of the targeted vein. The sample is taken so that the entire vein is within the sample therefore there is invariably a wedge of waste rock that is included with the lode sample. During face sampling the material is only collected within the vein boundary. This difference in sampling technique manifests as lower average grades and higher average widths within drill data when compared to face sample data. The contained metal calculated though each sampling method is the same.

Table 14-2: Face and diamond drilling sample statistics

Lode Type Variable No. of Samples Min. Max. Mean CV

NM Lode

Drill hole

Au (g/t)

211

0.0 319.30 29.13 1.63

Sb (%) 0.0 59.50 11.73 1.16

Vein Width (m) 0.0 2.29 0.61 0.90

Face sample

Au (g/t)

2068

0.0 989.0 52.30 1.20

Sb (%) 0.0 67.10 25.09 0.76

Vein Width (m) 0.0 2.21 0.65 0.78

CM Lode

Drill hole

Au (g/t)

205

0.0 353.07 29.97 1.72

Sb (%) 0.0 62.70 11.50 1.22

Vein Width (m) 0.0 2.19 0.7 0.71

Face sample

Au (g/t)

2048

0.0 1794.0 50.52 1.72

Sb (%) 0.0 66.8 18.3 0.94

Vein Width (m) 0.03 3.01 0.81 0.72

CE Lode Drill hole Au (g/t) 29 0.03 262.04 51.54 1.24




Sb (%) 0.0 55.40 16.36 0.86

Vein Width (m) 0.03 0.90 0.35 0.69

Face sample

Au (g/t)

180

0.06 821.90 73.57 1.45

Sb (%) 0.02 62.80 20.30 0.93

Vein Width (m) 0.0 1.68 0.64 0.75

CD Lode 220 Drill hole

Au (g/t)

64

0.02 116.45 12.20 1.43

Sb (%) 0.0 34.5 7.19 1.48

Vein Width (m) 0.04 4.54 1.65 0.87


Au (g/t)

14

0.11 62.0 9.55 1.20

Sb (%) 0.02 13.10 2.42 1.10

Vein Width (m) 0.06 1.11 0.57 0.74


Au (g/t)

8

0.16 367.3 46.44 2.06

Sb (%) 0.01 8.96 0.76 2.74

Vein Width (m) 0.03 0.83 0.39 0.91

B Lode

Drill hole

Au (g/t)

46

0.0 3021.60 17.78 10.31

Sb (%) 0.0 16.90 2.80 1.42

Vein Width (m) 0.04 1.90 0.86 0.67

Face sample

Au (g/t)

217

0.03 822.0 26.14 1.59

Sb (%) 0.12 62.80 12.35 1.27

Vein Width (m) 0.01 1.96 0.77 0.74

BSP Lode

Drill hole

Au (g/t)

13

0.16 44.0 6.02 1.35

Sb (%) 0.04 26.60 8.26 0.86

Vein Width (m) 0.04 0.76 0.39 0.50

Face sample

Au (g/t)

29

0.28 231.50 27.19 1.48

Sb (%) 0.14 53.10 17.32 0.76

Vein Width (m) 0.01 0.62 0.24 0.85

NV Lode 43

Drill hole

Au (g/t)

20

0.01 415.0 37.74 2.29

Sb (%) 0.01 14.70 4.29 0.79

Vein Width (m) 0.07 0.93 0.53 0.54

Face sample

Au (g/t)

31

0.04 146.20 27.32 1.49

Sb (%) 0.02 45.0 11.64 1.36

Vein Width (m) 0.02 1.61 0.70 0.78

NV Lode 44

Drill hole

Au (g/t)

23

0.12 29.50 6.11 1.38

Sb (%) 0.0 18.20 4.48 1.05

Vein Width (m) 0.03 1.26 0.65 0.66

Face sample

Au (g/t)

8

2.91 45.40 10.94 0.83

Sb (%) 0.26 19.10 8.19 0.79

Vein Width (m) 0.04 1.03 0.70 0.49

K Lode

Drill hole

Au (g/t)

24

0.01 65.40 5.26 1.73

Sb (%) 0.04 35.60 3.33 1.25

Vein Width (m) 0.06 0.86 0.48 0.48

Face sample Au (g/t)

12 2.41 54.0 19.46 0.78

Sb (%) 2.38 48.60 17.63 0.57




Vein Width (m) 0.08 0.28 0.16 0.38

C Lode Drill hole

Au (g/t)

72

0.0 176.0 21.28 1.73

Sb (%) 0.0 32.05 6.54 1.07

Vein Width (m) 0.03 1.54 0.87 0.43

CS 212 Lode Drill hole

Au (g/t)

10

1.18 48.0 6.76 1.45

Sb (%) 0.02 30.10 5.77 1.15

Vein Width (m) 0.05 0.79 0.53 0.45

CSE 213 Lode Drill hole

Au (g/t)

10

0.0 24.56 5.83 1.4

Sb (%) 0.0 16.04 2.53 2.16

Vein Width (m) 0.07 0.99 0.7 0.34

SCK CE 400 Drill hole

Au (g/t)

14

0.05 37.70 3.05 1.61

Sb (%) 0.01 33.70 1.55 2.29

Vein Width (m) 0.04 1.25 0.68 0.49

SCK LQ 405 Drill hole

Au (g/t)

12

0.0 321.0 35.20 2.50

Sb (%) 0.0 8.64 1.24 1.98

Vein Width (m) 0.07 0.44 0.27 0.54

SCK C 410 Drill hole

Au (g/t)

14

0.11 55.96 13.57 1.55

Sb (%) 0.07 11.60 1.50 1.24

Vein Width (m) 0.12 3.04 1.90 0.54

SCK W 420 Drill hole

Au (g/t)

8

0.02 1360.8 50.93 4.35

Sb (%) 0.0 0.11 0.04 0.99

Vein Width (m) 0.05 0.98 0.67 0.42

Brunswick Drill hole

Au (g/t)

113

0.0 115.78 5.58 1.10

Sb (%) 0.0 34.21 2.14 1.83

Vein Width (m) 0.09 11.50 2.74 1.04

Brunswick PK Drill hole

Au (g/t)

9

0.0 15.6 6.914 0.73

Sb (%) 0.0 5.75 2.76 0.82

Vein Width (m) 0.0 3.76 2.04 0.63

NE Lode

Drill hole

Au (g/t)

48

0.05 135.20 18.41 1.31

Sb (%) 0.01 52.80 11.55 1.34

Vein Width (m) 0.04 0.54 0.25 0.62

Face sample

Au (g/t)

60

1.32 131.90 39.53 0.78

Sb (%) 0.09 56.90 22.51 0.71

Vein Width (m) 0.01 1.33 0.48 0.76

NSP1 Lode

Drill hole

Au (g/t)

16

0.19 686.80 21.18 4.94

Sb (%) 0.09 20.84 9.72 0.82

Vein Width (m) 0.05 1.31 0.77 0.58

Face sample

Au (g/t)

28

0.04 135.40 17.10 1.74

Sb (%) 0.06 38.40 5.31 1.45

Vein Width (m) 0.02 1.80 1.01 0.49

NSP2 Lode Drill hole

Au (g/t)

5

0.11 59.0 17.84 1.07

Sb (%) 1.75 13.10 6.45 0.69

Vein Width (m) 0.05 0.58 0.36 0.66




Face sample

Au (g/t)

1

12.90 12.90 12.90 0

Sb (%) 9.95 9.95 9.95 0

Vein Width (m) 0.03 0.03 0.03 0


Au (g/t)

10

0.08 62.09 12.06 1.56

Sb (%) 0.07 24.0 4.44 1.24

Vein Width (m) 0.08 0.46 0.34 0.41

NSP4 Lode

Drill hole

Au (g/t)

18

0.14 62.80 11.68 1.55

Sb (%) 0.0 26.10 7.96 1.01

Vein Width (m) 0.07 0.71 0.48 0.44

Face sample

Au (g/t)

16

0.38 126.0 10.38 2.33

Sb (%) 0.06 20.33 9.87 0.63

Vein Width (m) 0.09 1.30 0.80 0.41


Au (g/t)

6

0.16 13.10 3.69 1.20

Sb (%) 0.02 24.90 4.91 1.58

Vein Width (m) 0.03 0.39 0.27 0.39

In order to show an example of the two factors described above, a study was undertaken on N Lode due to its large dataset. To negate the influence of low grade intercepts on the drilling information, all data was cut to the boundaries of the faces sampling data (Figure 14-1).

Figure 14-1: Long section of data collected for sample comparison analysis showing face sample (red) and drill hole (blue) data

This subset of N Lode data shows a higher average grade and a smaller vein width within the face sample data (Table 14-13). When this data is accumulated, by multiplying the grade by the vein width (SbAcc and AuAcc), the resulting numbers are representative of each other. Accumulation by vein true thickness produces similar, but more accurate, statistics than simple length weighting.



Table 14-3: Sample statistics of the data selected for analysis

Variable Type No. of Samples Min Max Mean

Sb (%) Drill 39 0.0 59.5 23.1

Face 871 0.0 66.5 32.9

Au (g/t) Drill 39 0.0 166.5 46.2

Face 871 0.0 989.0 65.9

Vein Width (m) Drill 39 0.065 0.994 0.332

Face 871 0.005 2.700 0.280

SbAcc Drill 39 0.0 20.1 7.3

Face 871 0.0 52.5 7.6

AuAcc Drill 39 0.0 97.1 16.8

Face 871 0.0 197.8 14.5

14.3 Data Interpretation and Domaining Each lode structure has been modelled separately and assigned a numeric zone code (Table 14-1). The identified intervals within both drill data and face sample data are incorporated into a wireframe of the lode structure. This wireframe is then used to flag the selected data with the corresponding zone code. The assays are then composited over the full width of the domained intersections. Data and observations from drill logs, core photography, underground face mapping, face photography and backs mapping were considered during the process of wireframe modelling.

Sub-domaining of NM, CM, CD, NE, CE and Brunswick Lodes was required to separate high grade and low grade populations to an acceptable degree. Structural controls on mineralisation were analysed to help determine where a break in subdomain boundary was located (Figure 14-2).

Figure 14-2: Long section of NM Lode showing fault intercepts identified as controls on mineralisation



14.4 Grade Capping The process of grade capping was altered slightly from the previous estimate. Statistical analysis of each subdomain was completed, including histograms, log probability plots and square root of the variogram/ madogram plots to determine appropriate grade caps for gold accumulation and antimony accumulation (Table 14-5). Monthly model reconciliations indicated that some of the lode model grades may have been overestimated locally. Therefore, the previously applied spatial limit to high grades (which used only the uncapped high grades for estimation) was modified to incorporate the surrounding lower grades. The same search distance parameters as the grade capped data estimate are then used.

The process is outlined below using AuACC in Domain 1 of the Cuffley Main model as an example.

1). Additional variables in the composite data file are created; AuACC_BC (Au accumulated bottom cut), which copies only intervals greater than the cut value from AuACC; and AuACC_NC which copies all values from AuACC (Table 14-4). The equivalent true thickness value is also copied across to a new variables called ‘TT_BC’.

2). To then differentiate between AuACC and AuACC_NC, the AuACC is then capped at the top cut grade, in this case 200 g/t (Table 14-4).

Table 14-4: CM lode Domain 3 composite file with new AuACC_BC and AuACC_NC variables created, and grade capping (200 g/t) applied to AuACC

3). An estimate is run using the AuACC_BC variable and a restricted search distance limited to the block size (either ~4 m or ~20 m) as well as the equivalent modelled variogram parameters. This results in a highly restricted set of blocks (Figure 14-3), which is used solely to identify the blocks to be estimated with the uncapped data.

Figure 14-3: Estimate produced using the AuACC_BC variable compared to composites



4). Two estimates are then run simultaneously, both using the same search parameters based on variogram ranges and QKNA; one estimate uses the AuACC variable (grade capped) and one estimate uses the AuACC_NC (non-grade capped) (Figure 14-4).

Figure 14-4: AuACC (left) and AuACC_NC (right) estimates compared to composites

5). Any blocks flagged by AuACC_BC are then allocated the AuACC_NC estimate in place of the AuACC estimate (Figure 14-5).

Figure 14-5: Final AuACC estimate; cells delineated by the AuACC_BC estimate have their AuACC value replaced with the AuACC_NC value (circled)

This process allows blocks close to high grade samples to be estimated with the full uncapped dataset, but blocks further than a defined distance away (beyond the AuACC_BC search parameters) to be estimated with the capped dataset. This process limits the spread of very high grades but retains the high local values in the close blocks (Figure 14-5). This process was applied to all grade capped data outlined in Table 14-5. Where datasets did not include face samples, and were therefore estimated using the larger block size of 1 x 20 x 20 m, the process required more manipulation of the search parameters and grade capping to ensure overestimation did not occur. High grades located where the larger block size was required were rare, due to the prevalence of higher grades in the face samples, rather than drill samples (i.e. where the small block size of 1 x 2.5 x 5 m is used).



Table 14-5: Accumulated gold statistics before and after top cuts

Variable Lode Domain Top Cut

No. Samples

No. Cut Samples Variable Min. Max. Mean CV

Au Acc CE 1 125 184 3 Uncut 0.01 231.69 21.75 1.29

Cut 0.01 125 20.7 1.07

Au Acc CE 2 12 25 1 Uncut 0.02 16.28 2.62 1.31

Cut 0.01 12 2.45 1.15

Au Acc NM 1 60 176 1 Uncut 0.0 197.80 5.65 2.89

Cut 0.0 51.17 4.56 1.76

Au Acc NM 2 100 137 1 Uncut 0.02 171.84 16.37 1.14

Cut 0.02 100 15.84 0.94

Au Acc NM 3 250 1147 1 Uncut 0.005 902.4 16.97 1.88

Cut 0.005 250 16.40 1.19

Au Acc NM 4 150 631 1 Uncut 0.003 448.48 13.03 1.84

Cut 0.003 150 12.56 1.34

Sb Acc NM 4 60 631 1 Uncut 0.003 111.17 6.15 1.33

Cut 0.003 60 6.07 1.21

Au Acc NM 5 120 188 1 Uncut 0.009 133.60 13.71 1.61

Cut 0.0 120 13.64 1.59

Sb Acc NM 5 35 188 2 Uncut 0.0 46.43 4.80 1.49

Cut 0.0 35 4.72 1.43

Au Acc CD 220 1 30 12 2 Uncut 0.0 163.41 20.63 2.18

Cut 0.0 30 8.16 1.49

Sb Acc CD 220 1 15 12 2 Uncut 0.0 138.57 18.94 2.07

Cut 0.0 15 4.53 1.43

Au Acc CD 220 2 15 35 3 Uncut 0.0 48.0 4.24 2.22

Cut 0.0 15 2.79 1.49

Sb Acc CD 220 2 3 35 2 Uncut 0.0 6.69 0.84 1.69

Cut 0.0 3 0.66 1.22

Au Acc CD 220 3 5 17 3 Uncut 0.04 10.52 1.97 1.38

Cut 0.04 5 1.56 1.12

Sb Acc CD 220 3 1 17 1 Uncut 0.0 1.9 0.4 1.16

Cut 0.0 1 0.35 0.94

Au Acc CD 221 - 5 14 2 Uncut 0.03 12.34 2.16 1.62

Cut 0.003 5 1.43 1.28

Sb Acc CD 221 - 2 14 1 Uncut 0.0 4.27 0.54 1.98

Cut 0.0 2 0.38 1.44

Au Acc CM 1 100 221 6 Uncut 0.0 841.22 22.58 3.1

Cut 0.0 100 15.88 1.38

Sb Acc CM 1 30 221 3 Uncut 0.0 58.18 4.68 1.54

Cut 0.0 30 4.52 1.40

Au Acc CM 2 20 294 7 Uncut 0.0 31.43 3.35 1.58

Cut 0.0 20 3.19 1.45

Sb Acc

CM

2

25

294

1

Uncut 0.0 29.1 1.63 2.4

Cut

0.0

25

1.61

2.37




No. Samples


Au Acc CM 3 200 717 3 Uncut 0.0 567.54 22.07 1.74

Cut 0.0 200 21.16 1.42

Sb Acc CM 3 50 717 3 Uncut 0.0 73.28 6.8 1.34

Cut 0.0 50 6.75 1.32

Au Acc CM 4 10 45 3 Uncut 0.0 20.24 1.96 1.90

Cut 0.0 10 1.68 1.57

Sb Acc CM 4 5 45 2 Uncut 0.0 13.74 0.90 2.54

Cut 0.0 5 0.66 1.83

Au Acc CM 5 200 969 4 Uncut 0.0 408.53 20.52 1.61

Cut 0.0 200 19.95 1.41

Sb Acc CM 5 100 969 1 Uncut 0.0 123.43 9.08 1.31

Cut 0.0 100 9.05 1.29

Sb Acc Brunswick 1 15 56 2 Uncut 0.0 36.59 4.15 1.61

Cut 0.0 15 3.53 1.27

Sb Acc Brunswick 2 5 45 3 Uncut 0.0 20.33 1.11 2.91

Cut 0.0 5 0.68 1.83

Au Acc C Lode - 40 72 3 Uncut 0.0 222.53 12.99 2.59

Cut 0.0 20 3.79 1.28

Sb Acc C Lode - 20 72 1 Uncut 0.0 24.59 3.85 1.32

Cut 0.0 40 8.29 1.28

Au Acc CS 213 - 10 10 1 Uncut 0.0 17.31 2.97 1.67

Cut 0.0 10 2.24 1.29

Sb Acc CS 213 - 8 10 1 Uncut 0.0 11.31 1.27 2.65

Cut 0.0 8 0.94 2.5

Au Acc NV 43 - 50 51 3 Uncut 0.0 148.81 9.17 2.71

Cut 0.0 50 7.85 1.82

Sb Acc NV 43 - 10 51 1 Uncut 0.0 28.27 2.25 1.85

Cut 0.0 10 1.90 1.17

Au Acc NV 44 - 10 31 2 Uncut 0.02 17.45 2.29 1.59

Cut 0.02 10 2.0 1.29

Sb Acc NV 44 - 5 31 4 Uncut 0.0 10.18 1.68 1.52

Cut 0.0 5 1.32 1.21

Au Acc NE Lode - 35 108 3 Uncut 0.02 49.65 5.75 1.53

Cut 0.02 35 5.46 1.39

Au Acc Brunswick PK - 20 9 1

Uncut 0.0 43.72 8.02 1.69

Cut 0 20 5.39 1.32

Sb Acc Brunswick PK - 10 9 1

Uncut 0.0 21.65 3.33 1.99

Cut 0.0 10 2.03 1.53

Au Acc B Lode - 100 263 3 Uncut 0.0 228.0 7.52 2.64

Cut 0.0 100 6.76 1.99

Sb Acc B Lode - 20 263 3 Uncut 0.0 33.42 3.12 1.38

Cut 0.0 20 3.02 1.25

Au Acc BSP Lode - 10 42 1 Uncut 0.0 20.30 2.45 1.50

Cut 0.0 10 2.21 1.2




No. Samples


Sb Acc BSP Lode - 10 42 1 Uncut 0.0 12.6 1.85 1.36

Cut 0.0 10 1.78 1.27

Au Acc NSP4 Lode - 20 34 3 Uncut 0.0 32.16 4.83 1.54

Cut 0.0 20 4.27 1.33

Au Acc NSP1 Lode - 40 44 3 Uncut 0.0 81.24 7.64 2.23

Cut 0.0 40 5.76 1.78

Sb Acc NSP1 Lode - 15 44 3 Uncut 0.0 23.20 2.77 1.84

Cut 0.0 15 2.51 1.62

Au Acc Sub KC 405 - 15 12 1 Uncut 0.0 58.81 6.45 2.48

Cut 0.0 15 2.80 1.58


Cut 0.0 30 5.28 1.59


Cut 0.0 20 5.04 1.72

14.5 Estimation Domain Boundaries Structural controls on mineralisation have been identified through underground mapping and structural interpretation of drill core (Figure 14-2). For Cuffley Main, N Main, Cuffley Deeps (220), N East, Cuffley East and Brunswick lodes these relationships have been used to guide estimation domain boundaries, all of which are hard boundaries (Figure 14-6 to Figure 14-11).

Figure 14-6: CM Lode Estimation domain boundaries and composite samples



Figure 14-7: NM Lode Estimation domain boundaries and composite samples

Figure 14-8: CD (220) Lode Estimation domain boundaries and composite samples

Figure 14-9: Brunswick Lode Estimation domain boundaries and composite samples



Figure 14-10: NE Lode Estimation domain boundaries and composite samples

Figure 14-11: CE Lode Estimation domain boundaries and composite samples

14.6 Vein Orientation Domains To estimate true thickness from the drill hole intersections, and convert the 2D tonnes and grade estimates to 3D tonnes and grade estimates, dip and dip-direction domains were interpreted in long section. Dip and strike domains were identified visually from the wireframe of the lode structure. The dip and strike of each domain was found by adjusting a plane to best fit the dip and strike of the domain. The details of this plane were then recorded and added to the drill data within the particular domain.



These dip and strike domains were used to create volume correction factors within the Z and Y directions using the following formula:

Z Correction Factor = 1/ sin (dip)

Y Correction Factor = Absolute (1/ sin (strike))

Volume Correction Factor = Z Correction Factor x Y Correction Factor.

The vein orientation domains are numbered and illustrated for CM and NM Lodes in Figure 14-12 and Figure 14-13.

Figure 14-12: CM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor

Figure 14-13: NM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor



14.7 Bulk Density Determinations Estimation of bulk density was assessed using the following two methods:

1). Bulk density for both Augusta and Cuffley was estimated using the analysed antimony grade using the following formula, which is based on the stoichiometry of stibnite and gangue.

BD = ((1.3951*Sb%)+(100-(1.3951*Sb%)))/(((1.3951*Sb%)/4.56)+((100-(1.3951*Sb%))/2.74))

2). Bulk density determinations of drill core samples were measured using a water immersion method. The samples used were whole pieces of diamond drill core, which were not coated in wax.

Figure 14-14 shows the measured bulk density values compared with the values calculated using the above bulk density formula. The bulk density determined from the immersion method shows a good concordance with the calculation.

For the resource estimate, bulk density was assigned using the formula method in line with previous estimates.

Figure 14-14: Bulk density determinations

The waste rock density of 2.74 has been averaged from 430 samples of drill core taken over a 2 year period exhibiting assays of below 0.01% antimony.

14.8 Variography Variographic analysis was carried out on true thickness, gold accumulation and antimony accumulation on the combined composited face and drill hole samples. The aim was to identify the directions of continuity of grade and thickness, and to assist in the selection of estimation ranges. The variography was undertaken in 2D after projecting the data to a constant easting.



Anisotropic normal score variograms were modelled by domain for each lode in Supervisor software after top cutting had taken place. The orientation of best continuity in grade accumulation and thickness was selected based on variographic analysis, verified by observations made during underground mapping and used to create ellipse wireframes to be compared with composite long sections. Due to the 2D variograms analysis, the downhole variograms were determined using the omnidirectional variograms at short lags. Examples of experimental, final back-transformed variograms and variogram ellipses for Domain 3 in the CM Lode are illustrated in (Figure 14-15 to Figure 14-26).

Where insufficient data was present within a domain to create a variograms for grade accumulations and true thickness, such as in Domain 6 of CM Lode, variogram ranges from comparable domains were used and the search orientations modelled visually.

Figure 14-15: Major direction experimental variogram for gold accumulation in Domain 3 of CM Lode



Figure 14-16: Semi major experimental variogram for gold accumulation in Domain 3 of CM Lode

Figure 14-17: Back transformed final variogram model for gold accumulation in Domain 3 of CM Lode



Figure 14-18: Long section view of CM Lode Domain 3 composites coloured by gold accumulated grade and search ellipse used

Figure 14-19: Major direction experimental variogram for antimony accumulation in Domain 3 of CM Lode



Figure 14-20: Semi major experimental variogram for antimony accumulation in Domain 3 of CM Lode

Figure 14-21: Back-transformed final variogram model for antimony accumulation in Domain 3 of CM Lode



Figure 14-22: Long section view of CM Lode Domain 3 composites coloured by antimony accumulated grade and search ellipse used

Figure 14-23: Major direction experimental variogram for True Thickness in Domain 3 of CM Lode



Figure 14-24: Semi major experimental variogram for True Thickness in Domain 3 of CM Lode

Figure 14-25: Back transformed final variogram model for antimony accumulation in Domain 3 of CM Lode



Figure 14-26: Long section view of CM Lode Domain 3 composites coloured by True Thickness grade and search ellipse used

14.9 Estimation Parameters True thickness, gold accumulation, and antimony accumulation were estimated in the 2D vertical plane using Ordinary Kriging for all domains, apart from Cuffley Main Domain 6, Brunswick PK, CSE 213, CSE 212, Sub King Cobra and NSP lodes, where an inverse distance squared method was applied. Search ellipsoids were orthogonal to the 2D block model orientation. A summary of the resource estimation process is as follows:

• Drill hole and face samples were projected into the plane of the estimate at 15,300 mE.

• The orientation of the major and semi major directions of the search ellipsoid for each lode was guided by the maximum continuity observed in the variography.

• The anisotropy of the search ellipsoid for each lode was guided by the anisotropy observed in the grade and thickness distribution.

The variogram parameters for the models estimated at year end 2016 are listed in Table 14-6. Each estimate involved three search passes with increasing search ellipse diameters on the second and third pass, as listed in Table 14-7. The influence limitation search parameters for domains that were grade capped are listed in Table 14-8..

Estimation was undertaken using a combined dataset of face sample and drill hole data. In the case of Cuffley Main, N Main, Cuffley Deeps (220), N East, Cuffley East and Brunswick grade domains were identified so estimation occurred individually for each domain. The domains were then cut to their boundaries and combined to complete the model. This process resulted in hard boundaries encapsulating these domains.



The resource was based on a minimum true lode width of 1.2 m, which is the practical minimum mining width applied at the Augusta and Cuffley Mines. For blocks with widths less than 1.2 m, diluted grades were estimated by adding a waste envelope with zero grade and 2.74 t/m3 bulk density to the lode.

The same true lode width approach has been applied to the Brunswick model, due to the lack of prior underground mining experienced at Brunswick. However, given that the lode widths are higher on average at Brunswick (Figure 14-27), SRK suggests that subsequent updates be completed using a more conventional 3D estimate (i.e. estimating in the X direction as well) to ensure the grade variability noted in the X direction is reflected during estimation. The current use of full length composites across the lode has smoothed the estimate in the thicker areas and negated the possibility of selectivity in the cross-strike direction.

Figure 14-27: Cross section 5818 mN of Brunswick lode, composites coloured by true thickness



Table 14-6: Variogram Model Parameters

Variable Lode Domain Rotations

Nugget 1st Spherical Structure 2nd Spherical Structure

Z Y X Major Semi major Minor Sill Major Semi Major Minor Sill

Au Acc

CM 1

-95 44 -75 0.38 37 28 20 0.26 107 38 21 0.28

Sb Acc -88 30 -85 0.21 37 20 20 0.36 112 31 21 0.38

True Thick -88 20 -80 0.4 21 28 16 0.21 50 33 17 0.33

Au Acc

CM 2

-85 0 -90 0.44 47 16 12 0.29 61 29 13 0.17

Sb Acc -84 -10 -80 0.26 48 11 8 0.4 67 16 9 0.21

True Thick -75 -5 -85 0.35 15 14 20 0.31 54 24 21 0.28

Au Acc

CM 3

-75 -30 -80 0.34 32 32 20 0.39 62 35 21 0.19

Sb Acc -75 -30 -80 0.22 24 23 20 0.43 65 42 21 0.3

True Thick -80 -20 -85 0.34 39 22 20 0.33 56 39 21 0.28

Au Acc

CM 4

135 75 45 0.47 72 34 20 0.1 85 59 21 0.32

Sb Acc 135 75 45 0.65 28 20 20 0.08 61 35 21 0.17

True Thick 135 75 45 0.3 28 12 20 0.23 82 14 21 0.4

Au Acc

CM 5

-80 10 -80 0.23 22 16 20 0.4 59 37 21 0.31

Sb Acc -85 0 -85 0.25 23 13 14 0.45 61 35 21 0.25

True Thick -85 0 -85 0.28 17 18 20 0.28 46 33 21 0.4

AuAcc

NM 1

15 -85 0 0.23 31 32 20 0.3 69 54 21 0.38

SbAcc 15 -85 0 0.11 27 21 20 0.43 54 48 21 0.42

True Thick 15 -85 0 0.27 19 6 6 0.29 23 12 7 0.34

AuAcc

NM 2

-90 55 -90 0.64 12 14 14 0.03 18 17 15 0.32

SbAcc -90 60 -90 0.67 13 17 14 0.09 38 19 15 0.34

True Thick -90 60 -90 0.70 21 20 16 0.08 46 23 17 0.28

AuAcc

NM 3

-75 40 -100 0.70 52 19 20 0.21 100 77 21 0.18

SbAcc -78 34 -100 0.48 44 16 20 0.35 94 64 21 0.33

True Thick -79 30 -100 0.32 64 18 20 0.35 125 65 21 0.35






AuAcc

NM 4

-73 35 -100 0.45 35 33 20 0.18 129 63 21 0.28

SbAcc -76 20 -100 0.22 26 29 20 0.32 100 52 21 0.41

True Thick -76 30 -100 0.44 30 39 20 0.18 71 54 21 0.3

AuAcc

NM 5

-84 5 -100 0.27 23 20 20 0.24 78 27 21 0.41

SbAcc -84 5 -100 0.20 19 20 20 0.36 94 31 21 0.37

True Thick -78 10 -100 0.47 10 17 12 0.23 27 19 13 0.25

AuAcc

BRU 1

-76 -35 -84 0.45 172 132 20 0.24 206 175 21 0.36

SbAcc -77 -30 -85 0.29 54 67 20 0.20 99 91 21 0.44

True Thick -62 -30 -85 0.34 149 40 20 0.26 174 96 21 0.38

AuAcc

BRU 2

110 -45 97 0.25 50 43 20 0.33 74 50 21 0.36

SbAcc -75 -45 -82 0.28 98 48 20 0.32 109 68 21 0.27

True Thick -42 72 -125 0.30 58 62 20 0.33 105 87 21 0.32

AuAcc

NE 1

-90 70 -90 0.44 56 52 20 0.29 75 53 21 0.27

SbAcc -90 30 -90 0.34 24 23 20 0.33 34 27 21 0.32

True Thick -90 75 -90 0.43 59 21 20 0.22 100 40 21 0.35

AuAcc

NV 43 1

-68 25 -95 0.22 33 30 20 0.35 50 46 21 0.31

SbAcc -63 24 -105 0.43 20 27 20 0.14 33 30 21 0.36

True Thick -61 29 -105 0.44 31 18 16 0.16 33 21 17 0.44

AuAcc

NV 44 1

91 -5 105 0.30 52 38 20 0.21 71 71 21 0.42

SbAcc 105 -35 100 0.55 49 20 20 0.08 66 40 21 0.28

True Thick 110 0 105 0.10 16 20 20 0.24 62 35 21 0.63

Au Acc

K 1

122 -15 100 0.35 35 22 20 0.22 58 38 21 0.2

Sb Acc 122 -30 95 0.20 20 16 20 0.29 61 39 21 0.47

True Thick 124 -40 96 0.43 15 20 20 0.25 51 31 21 0.24

Au Acc 2 -90 15 -90 0.30 109 39 20 0.24 110 43 21 0.34






Sb Acc CD 220

-80 0 -90 0.20 89 41 20 0.32 112 51 21 0.43

True Thick -90 -20 -90 0.40 75 49 20 0.25 81 50 21 0.3

Au Acc CD 220 3

-66 34 -96 0.20 91 37 20 0.33 116 48 21 0.4

Sb Acc -60 65 -100 0.25 94 41 20 0.21 142 44 21 0.5

True Thick -75 60 -90 0.34 116 69 20 0.23 126 80 21 0.33

Au Acc

CE 1

-90 -70 -90 0.41 30 12 9 0.14 37 24 10 0.45

Sb Acc -90 -80 -90 0.30 34 27 8 0.30 45 29 10 0.4

True Thick -90 -80 -90 0.44 37 9 11 0.26 43 22 12 0.3

Au Acc

B 1

122 -15 100 0.35 35 22 20 0.22 58 38 21 0.2

Sb Acc 122 -30 95 0.20 20 16 20 0.29 61 39 21 0.47

True Thick 124 -40 95 0.43 15 20 20 0.25 51 31 21 0.24

Au Acc

BSP 1

-70 0 -110 0.50 18 14 12 0.16 22 25 13 0.35

Sb Acc -82 -5 -110 0.45 12 16 14 0.22 19 18 15 0.33

True Thick -85 0 -110 0.39 15 14 13 0.23 29 25 14 0.38

AuAcc

C 1

-110 0 -120 0.27 37 20 9 0.31 52 28 10 0.35

SbAcc -110 0 -125 0.33 9 11 9 0.33 17 12 10 0.28

True Thick -105 0 -120 0.20 91 9 6 0.39 112 14 7 0.39



Table 14-7: Search Parameters for top cut estimate

Variable Lode Domain

First Pass Second Pass Third Pass Octant search

Min octant

Min sample

per octant

Max sample

per octant

Search Distance Samples Search Distance Samples Search

Distance Samples

Long axis

Short axis Min Max Long

axis Short axis Min Max Long

axis Short axis Min Max

Au Acc

CM 1

100 50 2 5 200 100 3 10 300 150 3 16 -

Sb Acc 100 50 2 5 200 100 3 10 300 150 3 16 -

True Thick 75 50 2 5 150 100 3 10 225 150 3 16 -

Au Acc

CM 2

75 50 2 6 150 100 3 10 225 150 4 20 -

Sb Acc 75 50 2 6 150 100 3 10 225 150 4 20 -

True Thick 75 50 2 6 150 100 3 10 225 150 4 20 -

Au Acc

CM 3

75 50 2 5 187.5 125 3 10 450 300 4 20 -

Sb Acc 75 50 4 9 187.5 125 3 10 450 300 4 20 -

True Thick 75 50 2 6 150 100 3 10 225 150 4 20 -

Au Acc

CM 4

75 50 8 12 187.5 125 3 16 450 300 3 20 0

Sb Acc 75 50 2 12 187.5 125 3 16 450 300 3 20 1 1 1 8

True Thick 75 50 6 12 187.5 125 3 16 450 300 3 20 1 1 2 8

Au Acc

CM 5

100 50 4 8 200 100 2 10 600 300 3 20 -

Sb Acc 100 50 4 8 200 100 2 10 600 300 3 20 -

True Thick 75 50 2 5 150 100 2 16 300 200 5 20 -

Au Acc

NM 1

75 50 4 8 150 100 5 16 450 300 6 16 -

Sb Acc 75 50 4 12 150 100 5 16 450 300 6 20 -

True Thick 75 50 2 12 150 100 3 16 450 300 6 20 -

Au Acc

NM 2

45 15 4 12 112.5 37.5 4 16 270 90 5 20 -

Sb Acc 45 20 4 12 112.5 50 4 16 270 120 5 20 -

True Thick 45 25 4 12 112.5 62.5 4 16 270 150 5 20 -

AuAcc NM 3

100 75 3 12 250 187.5 4 16 600 450 5 20 -

SbAcc 75 50 2 12 187.5 125 3 16 450 300 4 20 -





Min octant

Min sample

per octant

Max sample

per octant


Distance Samples

Long axis




True Thick 75 50 3 12 187.5 125 4 16 450 300 5 20 -

Au Acc

NM 4

100 75 3 6 250 187.55187 4 12 600 450 6 20 -

Sb Acc 100 75 4 10 250 187.5 4 12 600 450 6 20 -

True Thick 100 75 4 10 250 187.5 4 12 600 450 6 20 -

AuAcc

NM 5

100 50 4 8 250 125 2 10 600 300 3 20 -

SbAcc 100 50 6 10 250 125 3 10 600 300 3 20 -

True Thick 100 50 3 5 250 125 2 10 600 300 3 20 -

AuAcc

BRU 1

70 60 2 4 140 120 2 6 210 180 1 10 -

SbAcc 60 50 4 6 120 100 2 8 180 150 1 15 -

True Thick 80 60 2 4 160 120 2 8 240 180 1 15 -

AuAcc

BRU 2

50 60 2 4 100 120 2 6 150 180 1 15 -

SbAcc 70 60 2 6 140 120 2 8 210 180 1 15 -

True Thick 70 40 2 4 140 80 2 8 210 120 1 15 -

AuAcc BRU PK 1

55 55 4 10 137.5 137.5 5 12 220 220 5 12 -

SbAcc 55 55 4 10 137.5 137.5 5 12 220 220 5 12 -

True Thick 55 55 4 10 137.5 137.5 5 12 220 220 5 12 -

AuAcc

NE 1

75 50 2 6 187.5 125 3 10 300 200 2 16 -

SbAcc 75 50 2 6 187.5 125 3 10 300 200 2 16 -

True Thick 75 50 2 6 187.5 125 3 10 300 200 2 16 -

AuAcc

NV 43 1

75 50 2 6 150 100 3 8 225 150 1 15 -

SbAcc 75 50 2 6 150 100 3 8 225 150 1 15 -

True Thick 75 50 2 6 150 100 3 8 225 150 1 15 -

AuAcc NV 44 1

75 50 1 6 150 100 2 8 225 150 3 15 -

SbAcc 75 60 1 6 150 120 2 8 225 120 3 15 -





Min octant

Min sample

per octant

Max sample

per octant


Distance Samples

Long axis




True Thick 75 50 1 6 150 100 2 8 225 150 3 15 -

AuAcc

K 1

80 50 4 10 200 125 5 12 320 200 5 12 -

SbAcc 80 50 4 10 200 125 5 12 320 200 5 12 -

True Thick 80 50 4 10 200 125 5 12 320 200 5 12 -

AuAcc

C 1

50 50 4 6 100 100 3 10 150 150 2 15 -

SbAcc 50 50 4 6 100 100 3 10 150 150 2 15 -

True Thick 50 50 4 6 100 100 3 10 150 150 2 15 -

AuAcc CD 220 1

50 25 1 3 100 50 2 6 150 75 3 10 -

SbAcc 50 25 2 3 100 50 2 8 150 75 3 10 -

True Thick 50 25 1 3 100 50 4 6 150 75 3 10 -

AuAcc CD 220 2

75 50 1 3 150 100 3 6 225 150 3 12 -

SbAcc 75 50 1 2 150 100 3 6 225 150 3 12 -

True Thick 75 50 6 8 150 100 3 10 225 150 3 12 -

AuAcc CD 220 3

75 50 4 6 150 100 3 10 225 150 2 20 -

SbAcc 75 50 1 2 150 100 2 10 225 150 3 20 -

True Thick 75 50 2 2 150 100 3 10 450 300 3 20 -

AuAcc CD 221 1

50 50 2 4 100 100 2 4 150 150 3 8 -

SbAcc 50 50 3 4 100 100 2 4 150 150 3 8 -

True Thick 50 50 3 4 100 100 2 4 150 150 3 8 -

AuAcc CD 225 1

40 40 5 10 100 100 5 12 200 200 5 12 -

SbAcc 40 40 5 10 100 100 5 12 200 200 5 12 -

True Thick 40 40 5 10 100 100 5 12 200 200 5 12 -

AuAcc CE 1

75 50 2 15 187.5 125 1 20 225 150 1 20 1 1 1 8

SbAcc 75 50 2 15 187.5 125 1 20 225 150 1 20 1 1 1 8





Min octant

Min sample

per octant

Max sample

per octant


Distance Samples

Long axis




True Thick 75 50 2 15 187.5 125 1 20 225 150 1 20 1 1 1 8

AuAcc

B 1

50 50 2 4 100 100 3 6 150 150 1 15 -

SbAcc 50 50 2 4 100 100 3 6 150 150 1 15 -

True Thick 50 50 2 4 100 100 3 10 150 150 1 15 -

AuAcc

BSP 1

50 50 4 6 100 100 2 8 150 150 1 15 1 2 1 8

SbAcc 50 50 4 6 100 100 2 8 150 150 1 15 1 2 1 8

True Thick 50 50 4 6 100 100 2 8 150 150 1 15 1 2 1 8

AuAcc

CS 212 1

40 40 4 10 100 100 5 12 160 160 5 12 -

SbAcc 40 40 4 10 100 100 5 12 160 160 5 12 -

True Thick 40 40 4 10 100 100 5 12 160 160 5 12 -

AuAcc

CS 213 1

40 40 4 10 100 100 5 12 160 160 5 12 -

SbAcc 40 40 4 10 100 100 5 12 160 160 5 12 -

True Thick 40 40 4 10 100 100 5 12 160 160 5 12 -

AuAcc SubKC

400 1

40 40 4 10 100 100 5 12 160 160 5 12 -

SbAcc 40 40 4 10 100 100 5 12 160 160 5 12 -

True Thick 40 40 4 10 100 100 5 12 160 160 5 12 -

AuAcc SubKC

405 1

30 30 4 10 75 75 5 12 120 120 5 12 -

SbAcc 30 30 4 10 75 75 5 12 120 120 5 12 -

True Thick 30 30 4 10 75 75 5 12 120 120 5 12 -

AuAcc SubKC

410 1

40 40 4 10 100 100 5 12 160 160 5 12 -

SbAcc 40 40 4 10 100 100 5 12 160 160 5 12 -

True Thick 40 40 4 10 100 100 5 12 160 160 5 12 -

AuAcc SubKC 420 1

50 50 4 10 125 125 5 12 200 200 5 12 -

SbAcc 50 50 4 10 125 125 5 12 200 200 5 12 -





Min octant

Min sample

per octant

Max sample

per octant


Distance Samples

Long axis




True Thick 50 50 4 10 125 125 5 12 200 200 5 12 -

AuAcc

NSP1 1

40 20 4 10 100 50 5 12 160 80 5 12 -

SbAcc 40 20 4 10 100 50 5 12 160 80 5 12 -

True Thick 40 20 4 10 100 50 5 12 160 80 5 12 -

AuAcc

NSP2 1

40 20 4 10 100 50 5 12 160 80 5 12 -

SbAcc 40 20 4 10 100 50 5 12 160 80 5 12 -

True Thick 40 20 4 10 100 50 5 12 160 80 5 12 -

AuAcc

NSP3 1

40 20 4 10 100 50 5 12 160 80 5 12 -

SbAcc 40 20 4 10 100 50 5 12 160 80 5 12 -

True Thick 40 20 4 10 100 50 5 12 160 80 5 12 -

AuAcc

NSP4 1

40 20 4 10 100 50 5 12 160 80 5 12 -

SbAcc 40 20 4 10 100 50 5 12 160 80 5 12 -

True Thick 40 20 4 10 100 50 5 12 160 80 5 12 -

AuAcc

NSP5 1

40 20 4 10 100 50 5 12 160 80 5 12 -

SbAcc 40 20 4 10 100 50 5 12 160 80 5 12 -

True Thick 40 20 4 10 100 50 5 12 160 80 5 12 -



Table 14-8: Search Parameters for influence limitation estimate


First Pass

Search Distance Samples

Long axis Short axis Minimum Maximum

AuAcc_bc CM 1

5 4 1 3

SbAcc_bc 5 4 1 3

AuAcc_bc CM 2

5 3 1 3

SbAcc_bc 5 4 1 3

AuAcc_bc CM 3

5 4 1 3

SbAcc_bc 5 4 1 3

AuAcc_bc CM 4

5 4 1 3

SbAcc_bc 5 4 1 3

AuAcc_bc CM 5

5 4 1 3

SbAcc_bc 5 4 1 3

AuAcc_bc NM NM

1 5 4 1 3

AuAcc_bc NM NM

2 2

5 4 1 3

AuAcc_bc NM NM

3 5 3 1 3

AuAcc_bc NM 4

5 3 1 3

SbAcc_bc 5 4 1 3

AuAcc_bc NM 5

5 3 1 3

SbAcc_bc 5 3 1 3

AuAcc_bc BRU PK -

4 4 1 2

SbAcc_bc 4 4 1 2

AuAcc_bc CD 220 1

15 15 1 3

SbAcc_bc 10 10 1 3

AuAcc_bc CD 220 2

15 15 1 3

SbAcc_bc 15 15 1 3

AuAcc_bc CD 220 2

10 10 1 3

SbAcc_bc 10 10 1 3

AuAcc_bc CD 221 -

15 10 1 3

SbAcc_bc 15 10 1 3

AuAcc_bc CD 225 -

4 4 1 2

SbAcc_bc 4 4 1 2

AuAcc_bc B -

5 4 1 2

SbAcc_bc 5 4 1 2

AuAcc_bc CS 212 -

4 4 1 2

SbAcc_bc 4 4 1 2

AuAcc_bc CS 213 -

4 4 1 2

SbAcc_bc 4 4 1 2




First Pass

Search Distance Samples

Long axis Short axis Minimum Maximum

AuAcc_bc Sub KC 400 -

20 20 1 2

SbAcc_bc 20 20 1 2


20 20 1 2

SbAcc_bc 15 15 1 2


20 20 1 2

SbAcc_bc 20 20 1 2


20 20 1 2

SbAcc_bc 20 20 1 2

14.10 Block Model Estimation Grade accumulation and true thickness were estimated into 2D block models, whose cell centroids had an arbitrary easting of 15,300 mE. The 2D estimates were run with all data, including face samples and diamond drill hole samples, for two different cell sizes resulting in two models. Where influence limitation top cutting was applied, the models were run twice for each cell size. Once domain models were combined (if required), the high density face samples were used as a guide to delineate the Measured Resource, which retained the smaller block size. The remaining large block size Indicated and Inferred areas were regularised to the small block size, and the Measured smaller size block model was then added to create the final model. The cell sizes were selected based on the sample spacing of each area. Areas of high sample density are almost always face sampling (development) and areas of low sample density are usually from drill intercepts only, ranging from 20 m to 80 m spacing. Sub-cells were used in the Y and Z directions to better define the mining depletion and domain boundaries. The block model origins and number of cells are specific to the modelled lode. The common specifications for the models are given in Table 14-9.

Table 14-9: Block Model Dimensions

High Sample Data Density Low Sample Data Density

Block Dimensions Discretization

Block Dimensions Discretization

X 1 1 1 1

Y 2.5 2 20 3

Z 5 3 20 3

After the models were depleted and resource categories applied, the models were adjusted to replace the X cell dimension (XINC) with the estimated true thickness. The models from different lode structures were then combined into a global model, which was used for mine planning and creation of the reserve. The specifications for the global model are given in Table 14-10.

Table 14-10: Mine Planning Regularised Block Model Dimensions

Block model origin Block dimension (m) Number of cells

X 14500 1200 1

Y 4000 0.5 4200

Z 0 0.5 2400



14.11 Block Model Validation All statistics presented in this section refer to grades prior to dilution by the minimum mining width of 1.2 m. The resource estimate was validated by:

• Visual comparison of the sample thickness and grades to the estimated model grades in long section

• Visual comparison of the estimated grades to previous estimates in long section

• Comparison of the de-clustered accumulated sample grades with the model grades

• Plotting the sample and block estimated true thickness, gold and antimony accumulation grades on Swath plots.

There was a correlation when visually comparing the sample thickness and grades with the estimated grades. As expected, a greater degree of smoothing of the grades was evident where there were fewer samples available. The Slope of Regression (SoR) was analysed for each domain that used Ordinary Kriging as an estimation technique, to determine whether conditional bias was occurring. Where data was abundant in each domain, the SoR was close to 1, indicating conditional unbiasedness and suggesting the estimates were appropriate.

Results from the comparison of the declustered, thickness weighted, mean composite grades with the mean model grades are shown in Table 14-11.

Declustering was completed as required for each domain of each lode based on the sample spacing. Where sample population is small but sample spacing is large, differences greater than 10% often exist between composite and model values and are considered acceptable. An example of this was CD (221) Lode. There were also localised variations in thickness and grade in some domains, due to the presence of small-scale structural controls on mineralisation. Subsequent declustering over these areas led to some mixing of grade populations; in these cases declustering was not used.



Table 14-11: Declustered thickness weighted, composited sample grades and widths compared to estimated model grades and true thickness values Au (g/t) Sb (%) True Thickness (m) Model/Composite

Lode Domain Declustering Declustered Composites Model Declustered

Composites Model Declustered Composites Model Au (g/t) Sb (%) Vein

Width (m)

CM 1 25 x 25 52.96 40.98 13.46 11.99 0.37 0.34 77% 89% 92%

CM 2 25 x 25 18.13 13.93 7.11 5.14 0.30 0.31 77% 72% 105%

CM 3 None 71.84 61.22 19.82 18.61 0.36 0.37 85% 94% 101%

CM 4 75 x 75 4.88 3.86 1.74 1.50 0.39 0.33 79% 86% 85%

CM 5 100 x 100 28.88 23.23 11.72 9.66 0.35 0.36 80% 82% 103%

CM 6 None 3.32 3.05 0.008 0.003 0.24 0.22 92% 38% 92%

NM 1 75 x 75 19.76 14.78 10.16 8.41 0.21 0.20 75% 83% 93%

NM 2 None 68.22 62.32 40.16 39.43 0.26 0.25 91% 98% 96%

NM 3 25 x 25 65.97 59.48 30.49 28.76 0.32 0.32 90% 94% 100%

NM 4 75 x 75 31.11 24.76 15.03 10.73 0.25 0.24 80% 71% 86%

NM 5 50 x 75 42.03 34.48 10.79 10.09 0.29 0.30 82% 94% 105%

BRU - 100 x 100 4.98 4.7 1.86 1.72 1.11 1.13 94% 92% 102%

BRU PK - None 4.72 4.62 1.54 1.74 1.05 1.06 98% 113% 101%

NE - 50 x 50 25.72 24.39 14.98 14.50 0.15 0.14 95% 97% 95%

NV 43 - 25 x 25 37.87 23.51 7.67 5.68 0.34 0.33 62% 75% 97%

NV 44 - 50 x 50 8.46 7.73 6.01 5.77 0.27 0.24 91% 96% 89%

K - 10 x 10 6.94 7.48 5.22 5.67 0.29 0.28 108% 109% 97%

C - 25 x 25 14.26 12.09 6.59 5.67 0.49 0.52 85% 86% 106%

CD 220 1 None 6.74 5.97 5.99 3.29 1.24 1.16 89% 55% 94%

CD 220 2 25 x25 13.41 9.22 3.71 2.99 0.34 0.33 69% 81% 97%

CD 220 3 25 x 25 12.55 6.03 3.09 1.92 0.42 0.32 48% 62% 76%

CD 221 - 50 x 50 10.05 6.55 2.39 1.43 0.23 0.22 65% 60% 95%

CD 225 - None 73.2 72.0 1.70 1.50 0.19 0.17 98% 91% 87%

CE - 25 x 25 67.81 60.33 19.42 17.07 0.25 0.25 89% 88% 100%



Au (g/t) Sb (%) True Thickness (m) Model/Composite

Lode Domain Declustering Declustered Composites Model Declustered

Composites Model Declustered Composites Model Au (g/t) Sb (%) Vein

Width (m)

B - 25 x 25 40.52 22.69 8.80 8.98 0.34 0.34 56% 102% 100%

BSP - 25 x 25 13.63 10.73 10.4 9.41 0.20 0.19 79% 90% 93%

CS 212 - 25 x 25 6.97 8.48 4.5 5.48 0.31 0.30 122% 122% 97%

CS 213 - 25 x 25 3.92 3.84 1.15 1.16 0.57 0.53 98% 101% 93%

SubKC 400 - 25 x 25 3.17 3.25 1.43 1.86 0.42 0.42 102% 131% 101%

SubKC 405 - 25 x 25 25.17 23.49 1.43 1.63 0.42 0.19 93% 114% 45%

SubKC 410 - 25 x 25 10.40 5.84 1.49 1.74 1.10 1.05 56% 117% 95%

SubKC 420 - 50 x 50 10.02 13.58 4.64 6.31 0.22 0.21 136% 136% 96%

NSP1 - 50 x 50 15.10 13.40 6.78 6.047 0.31 0.47 89% 89% 150%

NSP2 - 25 x 25 16.90 17.25 6.33 6.23 0.14 0.21 102% 98% 152%

NSP3 - 50 x 50 10.02 13.58 4.64 6.31 0.22 0.21 136% 136% 96%

NSP4 - 25 x 25 10.24 10.46 8.83 8.02 0.44 0.35 102% 91% 80%

NSP5 - 25 x 25 3.92 4.86 4.62 7.01 0.22 0.21 125% 152% 97%



Swath plots were created for each domain of each modelled lode (Figure 14-28 to Figure 14-33). They compare declustered (where used, see Table 14-11), composited, accumulated grades with modelled accumulated grades.

Figure 14-28: NM Lode Domain 3 Gold Accumulation Swath Plot by Northing

Figure 14-29: NM Lode Domain 3 Gold Accumulation Swath Plot by Elevation



Figure 14-30: NM Lode Domain 2 Antimony Accumulation Swath Plot by Northing

Figure 14-31: NM Lode Domain 2 Antimony Accumulation Swath Plot by Elevation



Figure 14-32: CM Lode Domain 1 Gold Accumulation Swath Plot by Northing

Figure 14-33: CM Lode Domain 1 Gold Accumulation Swath Plot by Elevation



Figure 14-34: CM Lode Domain 1 Vein Width Swath Plot by Northing

Figure 14-35: CM Lode Domain 1 Vein Width Swath Plot by Elevation

14.12 Mineral Resource Classification Classification of the Mineral Resources takes into account Mandalay’s experience in mining the deposit, the comparable reconciliation observed between previous block model resource estimates, and the processing plant head grade during 2015 and 2016. Mandalay’s ongoing mining experience continues to improve the geological confidence and understanding of the controls on mineralisation,



which guides decisions made during the construction of the geological model. The classification criteria include the following:

• The Measured Resources are located within, and are defined by, the developed areas of the mine. This criterion ensures the estimate is supported by close spaced underground channel sampling and mapping.

• The Indicated Resources are located where drilling spacing on a nominal 40 mN x 40 mRL grid and there is high geological confidence in the geological model.

• The Inferred Resources have irregular or widely-spaced drill intercepts, is difficult to interpret due to multiple splays, or the structure does not have a demonstrated history of predictable mining.

The classification criteria are consistent with the previous Resource estimate conducted by SRK reported in March 2016 (SRK, 2016).

14.13 Mineral Resources The Mineral Resources are stated here for the Augusta, Cuffley and Brunswick deposits with an effective date of 31 December 2016. The Mineral Resource is depleted for mining up to 31 December 2016. The estimate excludes remnant mineralisation that remains in pillars.

The Augusta, Cuffley and Brunswick deposits consist of a combined Measured and Indicated Mineral Resource of 1,098,000 t at 6.9 g/t Au and 2.9% Sb, and an Inferred Mineral Resource of 611,000 t at 5.5 g/t Au and 1.5% Sb.

This Mineral Resource includes an ROM stockpile comprised of 20,000 t at 6.4 g/t Au and 2.2% Sb.

The Mineral Resources are reported at a cut-off grade of 3.5 g/t gold equivalent (AuEq), with a minimum mining width of 1.2 m. The gold equivalence formula used is calculated using typical recoveries at the Costerfield processing plant, a gold price of USD1,400 per ounce and an antimony price of USD10,000 per tonne as follows:

• AuEq =Au (g/t) + 1.76 x Sb (%)

• All relevant diamond drill hole and underground face samples available as of 30 November 2016 for the Augusta, Cuffley and Brunswick deposits, were used to inform the estimate. The estimation methodology was amended slightly (changes were made to the influence limitation grade capping estimation) from the approach used by SRK in the December 2015 Resource Estimate, reported in an NI 43-101 Technical Report on the Property (SRK, 2015).

Table 14-12: Mineral Resources at Costerfield, Inclusive of Mineral Reserves, as of 31 December 2016

Category Inventory (kt)

Gold Grade (g/t)

Antimony Grade (%)


Contained Antimony (kt)

Measured 286 9.5 4.0 88 11.4

Indicated 812 5.9 2.5 155 20.6

Measured + Indicated 1098 6.9 2.9 242 32.0

Inferred 611 5.5 1.5 108 9.0

Notes: 1) Mineral Resources estimated as of December 31, 2016, and depleted for production through December 31, 2016. 2) Mineral Resources are stated according to CIM guidelines and include Mineral Reserves. 3) Tonnes and contained gold (oz) are rounded to the nearest thousand; contained antimony tonnes are rounded to the

nearest hundred. 4) Totals may appear different from the sum of their components due to rounding. 5) A cut-off grade of 3.5 g/t AuEq over a minimum mining width of 1.2 m is applied where AuEq is calculated at a gold

price of USD1,400/oz, antimony price of USD10,000/t and exchange rate USD:AUD of 0.75.



6) The AuEq value is calculated using the formula: AuEq = Au g/t + 1.76 * Sb %. 7) Geological modelling and sample compositing was performed by Cael Gniel, who is a full-time employee of Mandalay

Resources, Chris Davis, MAusIMM, who is a full-time employee of Mandalay Resources and was independently verified by Danny Kentwell, FAusIMM, who is a full-time employee of SRK Consulting.

8) The Mineral Resource estimation was performed by Kirsty Sheerin, MAusIMM, who is a full-time employee of SRK Consulting, Cael Gniel, who is a full-time employee of Mandalay Resources and Chris Davis, MAusIMM, who is a full-time employee of Mandalay Resources. The resource models were verified by Danny Kentwell, FAusIMM, who is a full-time employee of SRK Consulting. Danny Kentwell, FAusIMM, a full-time employee of SRK Consulting is the Qualified Person under NI 43-101 and the Competent Person for the Resource.

Details of the Augusta, Cuffley and Brunswick Mineral Resources are stated in Table 14-13. Longitudinal projections of the models re-estimated at the end of 2016 are displayed in Figure 14-36 to Figure 14-47.

Table 14-13: Summary of the Augusta, Cuffley and Brunswick Mineral Resource, inclusive of Mineral Reserve

Lode Name Resource Category Tonnes Au (g/t) Sb (%) Au (oz) Sb (t)

Augusta D

eposit

E Lode

Measured 55,000 5.4 2.9 9,600 1,600

Indicated 54,000 4.1 2.4 7,000 1,300

Inferred 72,000 3.6 2.7 8,400 2,000

B Lode Measured 8,000 7.0 2.9 1,900 200

Indicated 28,000 6.6 1.9 5,900 500

B Splay Measured 2,000 3.0 2.2 200 100

Inferred 10,000 1.9 1.8 600 200

W Lode

Measured 22,000 7.1 3.4 5,100 800

Indicated 25,000 5.3 2.8 4,300 700

Inferred 30,000 5.1 2.3 4,900 700

C Lode Inferred 72,000 4.4 2.4 10,300 1,700

NM Lode

Measured 69,000 10.7 4.7 23,600 3,200

Indicated 141,000 6.3 2.8 28,700 3,900

Inferred 89,000 4.1 1.1 11,800 1,000

NE Lode

Measured 4,000 5.6 3.0 800 100

Indicated 33,000 4.3 2.5 4,600 800

Inferred 18,000 3.8 2.2 2,200 400

New Vein Measured 1,000 3.9 2.2 200 0

Indicated 67,000 6.6 1.4 14,000 1,000

NSW Lode Indicated 13,000 3.8 2.5 1,500 300

NW Lode Measured 1,000 6.2 3.7 100 0

Indicated 3,000 4.9 3.3 400 100

N Splay 1 Measured 1,000 8.8 2.5 200 0

Indicated 7,000 5.4 2.4 1,200 200

N Splay 2 Indicated 2,000 3.4 1.2 200 0


N Splay 4 Indicated 11,000 3.5 3.0 1,300 300


P1 Lode Measured 12,000 9.4 2.3 3,600 300

Indicated 9,000 10.2 2.6 3,000 200

P2 Lode Measured 4,000 6.2 3.4 700 100



Lode Name Resource Category Tonnes Au (g/t) Sb (%) Au (oz) Sb (t)

Indicated 18,000 4.3 2.1 2,500 400

Inferred 29,000 3.2 1.2 3,000 300

K Lode

Measured 2,000 2.8 2.5 200 0

Indicated 24,000 2.6 1.9 2,100 500

Inferred 21,000 2.2 1.5 1,500 300

Cuffley D

eposit

CM Lode

Measured 75,000 13.6 5.2 32,900 3,900

Indicated 64,000 6.7 2.6 13,900 1,700

Inferred 8,000 4.9 1.3 1,300 100

CE Lode Measured 11,000 13.2 4.7 4,600 500

Indicated 13,000 6.5 2.1 2,600 300

CS Lode Indicated 10,000 2.2 1.9 700 200

Inferred 2,000 2.1 1.0 100 0

CSE Lode Indicated 1,000 4.8 3.6 100 0

CD Lode Indicated 97,000 7.2 3.5 22,500 3,300

Inferred 5,000 1.3 1.6 200 100

CDW Lode Inferred 1,000 2.8 0.9 100 0

CDL Lode Inferred 26,000 7.4 0.1 6,200 0

AS Lode Indicated 28,000 4.8 3.7 4,300 1,000

Inferred 9,000 0.8 2.3 200 200

Brunswick Deposit Indicated 157,000 6.5 2.4 32,800 3,800

Inferred 59,000 4.5 1.6 8,600 1,000

Sub K

ing Cobra

SKC CE Inferred 13,000 2.5 1.0 1,000 100

SKC LQ Inferred 11,000 9.5 0.2 3,400 0

SKC C Inferred 71,000 7.5 1.2 17,100 800

SKC W Inferred 64,000 13.3 0.0 27,400 0

Stockpile Measured 20,000 6.4 2.2 4,000 400

Measured and Indicated 1,098,000 6.9 2.9 242,400 32,000

Inferred 611,000 5.5 1.5 108,300 9,000



Figure 14-36: NM Lode Resource Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.

Figure 14-37: CM Main Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.



Figure 14-38: CE Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.

Figure 14-39: CD (220) Lode Block Model with Resource Category Boundaries

N







Figure 14-42: CS (212) Lode Block Model with Resource Category Boundaries

Figure 14-43: CS (213) Lode Block Model with Resource Category Boundaries



Figure 14-44: NV (43) Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.

Figure 14-45: NV (44) Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.



Figure 14-46: K Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.

Figure 14-47: C Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.



Figure 14-48: BSP Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.

Figure 14-49: B Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate.



Figure 14-50: NE Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.

Figure 14-51: Brunswick Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.



Figure 14-52: Brunswick PK Lode Block Model with Resource Category Boundaries

Figure 14-53: NSP1 Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.



Figure 14-54: NSP2 Lode Block Model with Resource Category Boundaries


Figure 14-56: NSP4 Lode Block Model with Resource Category Boundaries Note: Black dots represent face and drill hole samples used in the resource estimate. White areas denote mined-out areas.




14.14 Cut-off Grade Calculations A 3.5 g/t AuEq cut-off grade over a minimum mining width of 1.2 m is applied where AuEq is calculated at a gold price of USD1,400/oz, an antimony price of USD10,000/t and a USD:AUD exchange rate of 0.75. The 3.5 g/t is derived by Mandalay based on recent cost, revenue, mining and recovery data.

14.15 Reconciliation The following graphs show the reconciliation between mined grades and 2016 Resource model grades reported within the same area mined, and split by month. Antimony grade shows good correlation over the year. Gold Grade shows a model overcall of 9% over 2016. This overcall is being addressed by changes to the top-cutting methodology as described in previous sections. Increased data density in high gold grade areas of the model, attained during 2016 via face sampling, also addresses the overcall seen in 2016.

Figure 14-43: Model Reconciliation – Antimony Grade



Figure 14-44: Model Reconciliation – Gold Grade

14.16 Other Material Factors SRK is not aware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, or political factors that could materially influence the Mineral Resources other than the modifying factors already described in other sections of this report.



15 Mineral Reserve Estimate From the Mineral Resource, a mine plan was designed based only on Measured and Indicated Resource blocks, predominantly using the cemented rock fill blast hole stoping method. A cut-off grade of 4.0 g/t AuEq and minimum mining widths of 1.8 m for development and 1.2 m for stoping were used. Planned and unplanned dilution was applied at zero grade.

The 2016 cut-off grade (4.0 g/t) did not change from the 2015 cut-off grade.

Financial viability of the mine plan was demonstrated at metal prices of USD1,200 /oz Au, USD8,000/t Sb and a conversion rate of 0.75 USD:AUD. The Measured and Indicated Resources within the mine plan were converted to the Proven and Probable Mineral Reserves estimate presented in Table 15-1.

Table 15-1: Mineral Reserves at Costerfield, as of 31 December 2016

Category Reserve (kt)

Gold grade (g/t)

Antimony grade

(%)


Contained Antimony

(kt)

Proven 184 8.1 3.5 48 6.4

Probable 434 5.7 2.6 80 11.1

Proven + Probable 619 6.5 2.8 128 17.5

Notes: 1) Mineral Reserve estimated as of 31 December 2016, and depleted for production through to 31 December 2016. 2) Tonnes are rounded to the nearest thousand; contained gold (oz) is rounded to the nearest thousand; contained

antimony (kt) is rounded to the nearest hundred. 3) Totals are subject to rounding error. 4) Lodes have been diluted to a minimum mining width of 1.2 m for stoping and 1.8 m for ore development. 5) A 4.0 g/t Au Equivalent (AuEq) cut-off grade is applied. 6) Commodity prices applied are: gold price of USD1,200 /oz, antimony price of USD8,000 /t and USD:AUD exchange

rate of 0.75. 7) The Au Equivalent value (AuEq) is calculated using the formula: AuEq = Au g/t + 1.64 * Sb %. 8) The Mineral Reserve is a subset of the Measured and Indicated only schedule of a Life of Mine Plan that includes

mining of Measured, Indicated and Inferred Resources. 9) The Mineral Reserve estimate was prepared by Chloe Cavill, BSc, who is a full-time employee of Mandalay, and was

independently verified by Peter Fairfield, FAusIMM, CP (Mining), who is a full-time employee of SRK Consulting and a Qualified Person under NI 43-101.

The net decrease of 17 koz Au in Proven and Probable Reserves for 2016 relative to 2015 consists of a total of 46 koz Au depleted from the 2015 Reserves, which has been positively offset by the addition of 29 koz Au added by resource conversion and mining re-evaluation. The 1.4 kt Sb net decrease in Proven and Probable Reserves consists of 5.5 kt Sb depleted from the 2015 Reserves, offset by the 4.1 kt added by resource conversion and mining re-evaluation.

15.1 Modifying Factors

Mining Dilution and Recovery Mining Dilution Airleg development, jumbo development, longhole bench stoping with cemented rock backfill and longhole uphole stoping with no backfill are the current mining methods utilised at Costerfield for the extraction of underground Mineral Reserves.

Planned and unplanned dilution has been considered for establishing the production schedule.

Planned dilution includes low grade material or waste rock that will be mined and is not segregated from the design.



Sources of planned dilution include:

• Waste rock (or low grade material) that is drilled and blasted within the drive profile and the overall grade of the blasted material justifies delivery to the mill

• Waste rock (or low grade material) within the confines of the stope limits. This includes footwall and/or hanging wall rock that has been drilled and blasted to maximise mining recovery and/or maintain favourable wall geometry for stability.

Due to the narrow width of the Augusta and Cuffley Lode mineralisation, Mineral Resources include an element of planned mining dilution, as Mineral Resources are reported to conform to a minimum 1.2 m mining width.

Unplanned dilution includes waste rock (or low grade material) and/or backfill from outside the planned production drive profile or stope limits that overbreaks or sloughs into the mining void and is bogged and delivered to the mill.

Unplanned dilution is the sum of overbreak (from deficient blasting practices) and fall off (as in wedge failure). Surveys of mined development drives and stopes to date are consistent with the figures presented in Table 15-2. The longhole overbreak and dilution factors are considered to be consistent with operational experience, as these measurements are based on stope inspection sheet measurements as well as stope scans that produce a 3D model of the open void, which is then interrogated using mine planning software to generate the final void volume.

Operating practices attempt to mine the stope as close to the lode width as possible, to limit the amount of planned and unplanned dilution reporting to the stope drawpoint. All planned and unplanned mining dilution is assumed to have zero grade.

The percentage of overall (i.e. planned and unplanned) dilution for proposed mining methods is shown in Table 15-2.

The overall dilution assumptions include unplanned dilution such as overbreak, fall off and bogging dilution. This dilution has been based on the experience obtained from current operations, and there is a good reconciliation (Section 14.15) between forecast tonnes and actual tonnes. Development dilution is based on end of month survey reports, which compare actual drive area against the designed area. The remaining stoping dilution figures are based on actual historical measurements as outlined above.

Mining Recovery The tonnage recovery factors shown in Table 15-2 represent the recovered proportion of planned mining areas for the different mining methods, and include in-situ ore plus dilution material. Recovery of half-upper stopes has increased since 2015, from 75% to 90%, due to the ongoing successful use of remote loaders for extraction of these stopes.

For stoping areas, visual inspections are carried out to estimate the stope void volume and to determine if any ore is left in the stopes. This information is recorded on stope inspection sheets. Stope scans are also conducted to confirm the data captured. All of this data is used to estimate the recovery factor.



Table 15-2: Costerfield Mine recovery and dilution assumptions

Mining Method Unit Airleg Development

Longhole Uphole N Lode

Longhole Uphole

Cuffley East

Longhole Uphole

All Others Longhole CRF

N Lode Longhole CRF Cuffley East

Longhole CRF All Others

Planned width m 1.8 1.2 1.2 1.2 1.2 1.2 1.2

Overall dilution % 15 40 0 25 40 0 25

Tonnage recovery factor % 100 90 90 90 95 95 95



Mine Design and Planning Process All mine design work has been completed using GEOVIA Surpac, Deswik ASD and Studio 5D Planner (S5DP), with all mine scheduling being undertaken in Enhanced Production Scheduler (EPS).

The mining shapes designed were assessed against the block model in S5DP in order to calculate tonnes and grade. A combined 3D block model was used for the interrogation and mining depletion process of the designed mining shapes.

The mining schedule included mining areas within the Augusta and Cuffley Mineral Resources above the gold equivalent cut-off grade.

After completing the interrogation and depletion processes, dependency rules including schedule constraints were applied to the designed shapes to link all the mining activities in a logical manner within the S5DP project. This information was then exported to an EPS file for scheduling purposes.

EPS software was used for life of mine scheduling. The key assumptions such as cut-off grade, mining dilution and recovery factors, assigning resources and rate of mining were included in scheduling calculations.

These design and scheduling programs are also in use for quarterly and monthly planning purposes.

Cut-off Grade Estimation of the design cut-off grade was based on past mining experience gained within both the Augusta and Cuffley deposits, as well as historical performances, both physical and economic, of the mining and processing methodologies.

The distribution of grade over the measured and indicated Mineral Resources was considered to enable grade-tonnage curves to be generated, from which cut-off equations were determined.

Based on the historical cost estimates from the Augusta Mine, as well as the planned future capital and operating cost structure for the entire operation, a cut-off grade of 4.0 g/t AuEq was calculated and applied to the Mineral Resource to determine the mine design shapes (i.e. stope shapes) that possessed acceptable economic grade to warrant mining.

The mine design shapes enabled a mine plan to be developed and independently verified as both technically achievable and economically viable once all Modifying Factors were considered.

For reporting purposes, 4.0 g/t AuEq was applied to determine the Mineral Reserve estimate.



16 Mining Methods 16.1 Introduction

The Augusta Mine is serviced by a decline haulage system developed from a portal within a box-cut with the majority dimensions of 4.8 m high by 4.5 m wide at a gradient of 1:7 down. The majority of the decline development has been completed with a twin boom jumbo; however, development of the decline from the portal to 2 Level was completed with a road-header. This section of the decline is 4 m high and 4 m wide. The decline provides primary access for personnel, equipment and materials to the underground workings.

The Augusta Mine is scheduled to produce mill feed from three different mining methods: full face development, longhole CRF stoping and uphole stoping. All mined material is transported to the Augusta box-cut before being hauled to either the Brunswick ROM Pad or Augusta Waste Rock Storage Facility.

Mandalay is extracting the mineral resources contained in the Cuffley Lode from the 1035 mRL to the 804 mRL. Access to the Cuffley Lode is via a single decline that connects to the existing Augusta decline at the 1030 mRL. The Cuffley Decline currently extends down to approximately 895 mRL. At the 935 mRL, the Cuffley Incline extends off the Cuffley Decline and accesses the resources from the 945 mRL to the 1020 mRL. A second decline within Cuffley, known as the 4800 Decline, accesses the southern part of the Cuffley Lode which is positioned south of the East Fault. This Decline commences at the 960 mRL and extends to 824 mRL.

The proposed stope outlines are presented in Figure 16-1.



Figure 16-1: Long section of proposed Underground Mine Design along Cuffley (base) and N Lode (top)



16.2 Geotechnical

16.1.1 Rock Properties Lithology and Structures Active underground mine workings are hosted within weakly metamorphosed siltstones of the lower Silurian-aged Costerfield formation. Underground operations target the NNW striking, sub-vertical dipping mineralised structures that are typically less than 500 mm in true width.

Targeted mineralised structures are bounded up-dip and down-dip by the Adder and King Cobra thrust faults respectively. The King Cobra Fault is observed as separate hanging wall and foot wall structures filled with strongly deformed siltstone and quartz horsetails. The zone of deformation within the King Cobra Fault can be up to 10 m wide. Offset across the King Cobra Fault is unknown. The Adder Fault is also filled with quartz and rubble and varies in width from less than 0.3 m to more than 2 m.

Significant second order structures include the northeast striking, northwest dipping faults that offset lode mineralisation (East and Brown faults). Variable offsets (less than ~2 m) are observed on shallow dipping structures that have caused strong shearing at lode intersections (Flat, Krait, Red Belly, Bushmaster, Tiger and Fox faults), as shown in Figure 16-2. A 3D structural model of all intersected mine-scale faults is maintained and is a key driver of pre-emptive ground control strategies.

Figure 16-2: Schematic cross section of the Augusta and Cuffley systems

Rock Strength A total of 58 Unconfined Compressive Strength (UCS) tests have been completed on the Costerfield formation siltstones since 2009. Intact rock strength increases with depth due to sustained weathering in the upper strata. At levels lower than 100 m below surface, intact rock strength exceeds 80 MPa.



Rock Stress In situ stress measurements have not been undertaken at Costerfield. In situ stress is approximated using the paper by Lee et al. (2010), for which over 1,000 stress measurements taken on the Australian continental tectonic plate were collated. The Costerfield Region (i.e. Augusta Mine) lies within the Lachlan Province, where the principal stress orientation is WNW-ESE with magnitudes of δ1= 55: δ2 = 35: δ3 = 30 for measurements greater than 500 m below surface. Stress measurements obtained from other mines in the region suggest a major principal stress magnitude of ~40 MPa at 565 - 575 m below surface, with a strong correlation between increased depth and increased in situ stress.

In situ stress in levels below 895 mRL has caused minor convergence (squeezing ground) in isolated areas around major fault zones. The magnitude of this squeezing is small enough to be contained by dynamic support.

Rock Mass Alteration Rock mass in the vicinity of mineralised structures is heavily fractured with multiple joint orientations, often with a portion of clay fill and smooth planar joint surfaces. In waste rock, away from mineralised lodes and discrete structures, the rock mass improves with lower fracture frequency and rough, tightly healed joint surfaces.

Hydrogeology The regional hydrogeology comprises the following two main aquifers – the Shallow Alluvial Aquifer (SAA) and the Regional Basement Aquifer (RBA):

• The SAA is comprised of silts, sands and gravels and is a perched groundwater system occurring across the site within the confines of the creek and valley floors. There is clear evidence that this aquifer is perched, laterally discontinuous and less common in the area.

• The RBA is comprised of Silurian metasediments and forms the basement aquifer, where groundwater mainly occurs within and is transmitted through fracture systems beneath the upper weathered profile, at depths of in excess of 20 m below natural surface.

Dewatering of underground workings is achieved via the two underground pump stations that pump to the Cuffley Pump Station, which is connected to the Brunswick Pit via a dedicated dewatering rising main. Inflow rates have increased as the mine advances deeper and to the northern and southern extents of mining targets. Water inflow peaked at 1.95 ML per day in March 2014 and has ranged from 1.0 - 1.4 ML per day in recent months.

16.1.2 Mine Design Parameters Mining Methods The dominant mining method is longitudinal longhole open stoping with cemented rock fill (CRF). Panels generally consisting of four operating levels are mined from the bottom up over the CRF, with longitudinal retreat to a quasi-central access. The following other mining methods are applied in order to access and optimise extraction of ore at Costerfield:

• Capital development with twin or single-boom jumbo

• Level development with airleg (hand-held rock drill)

• Overhand cut and fill (flat backing ore level development)

• Floor benching of level ore development

• Blind up-hole longitudinal longhole open stoping

• Longitudinal longhole open stoping with CRF



• Down hole vertical crater retreat (VCR)

• Avoca stoping with CRF (‘reverse fill’)

• Avoca stoping with rock fill (‘reverse fill’)

• Airleg rise mining.

Mining methods are selected to suit ore drive/ lode geometry and maximise ore recovery while minimising unplanned dilution.

Development Geometry Standard development profiles adopted at Costerfield include:

• 1.8 m wide x 2.8 m high ore drives

• 2.0 m wide x 2.8 m high access crosscuts

• 4.5 m wide x 4.8 m high decline

• 5.0 m wide x 4.8 m high level access

• 5.0 m wide x 6.5 m high truck tips

• 4.5 m wide x 4.8 m high ore stockpiles.

Non-standard development profiles may be mined to allow for major infrastructure (pump stations, explosives magazines, fan chambers etc.) or to suit applied mining methods (flat backing, floor benching etc.). Development spans and associated ground support are designed using empirical data to ensure the stability of mined spans.

Stope Geometry In response to observed ground conditions and production drill capability, inter-level spacing at Costerfield is variable. Stope strike length varies based on the applied mining method, observed ground conditions and machinery capability.

Stope geometry parameters include:

• Stope height: up to 17 m

• Stope strike length: 2.8 - 10.5 m

• Stope design width: 1.5 m

• Stope dip: 70° - 90°.

Non-standard stope geometry may be mined to maximise ore extraction under unique circumstances (i.e. remnant mining or geological complexity). The empirical stope performance chart is consulted to ensure that designed stope spans will allow safe efficient extraction of target mineralisation.

Pillars and Offsets In mine design and planning, the following pillars and offsets are observed to ensure stability of mined excavations:

• Decline development is designed and mined with a 30 m offset to target mineralised structures; to date stope production blasting has not influenced decline stability having applied the 30 m offset.

• The minimum inter-level pillar width to height ratio is 1:2 (i.e. for 1.8m wide ore drives, the minimum inter-level spacing is 3.6 m).

• The minimum horizontal clearance between sub-parallel ore drives is 2 m.

• The minimum pillar strike between unfilled blind up-hole longitudinal open stopes is 3 m.



Backfill CRF is the most commonly used backfill material at Costerfield. CRF is exposed vertically in the longitudinal retreat of CRF-filled longhole open stopes, and horizontally in the mining of sill pillars at the toe of blind up-hole longitudinal open stopes. Loose rock fill is used where vertical and horizontal exposures to filled voids are not required (i.e. level close-out stopes and adjacent to waste pillars).

16.1.3 Ground Support Development Ground Support Ground support elements installed in standard development profiles include:

• 2.4 m 20 mm diameter galvanised resin bolts

• 2.4 m 47 mm diameter galvanised friction bolts

• 1.5 m 33 mm diameter galvanised friction bolts

• 2.4 m x 3.6 m 5.6 mm diameter gauge galvanised mesh



• 2.4 m x 1.5 m 4.0 mm diameter gauge galvanised mesh.

When spans exceed 5.5 m in development intersections, or in response to deteriorating ground conditions and discrete structures, cable bolts (single strand, non-galvanised, bulbed, 4.5 - 6.0 m) are installed to ensure the stability of development profiles.

Additional ground support may be installed to support non-standard development profiles or in response to particularly poor ground conditions. Fibrecrete, resin injection, spiling, sets and straps have been installed in the past to support poor ground, development/ stoping interactions and faults/ shear zones. Bolts – 2.4 m and 1.8 m yield “lok” bolts – are installed in areas where squeezing ground is expected.

Stoping Ground Support Additional support for designed stopes is installed on an ‘as required’ basis in response to compromised stope geometry, poor rock mass, interactions with faults/shears or interactions with other stopes and development. Single strand, non-galvanised, bulbed, 4.5 - 6 m cable bolts are generally installed as secondary support for stopes. Other forms of ground support including resin bolts, friction bolts, mesh, fibrecrete, resin injection and straps may also be installed to provide secondary support for designed stopes.

16.3 Mine Design

16.1.4 Method Selection Longhole CRF stoping has been selected as the preferred mining method for the remaining Augusta and Cuffley Mineral Resources. This is based on the orebody geometry as well as the favourable experience gained through the use of this method. Other mining methods have been trialled.

This process necessitates that crown/ sill pillars are left in place on a regular basis to ensure local mine stability. Recovery of these pillars is planned to be undertaken by means of uphole production stoping.



16.1.5 Method Description Mining within the Augusta Mine targets several individual lodes (including W, NM, NE, P, E and Cuffley lodes) which vary in width from 0.1 m to 1.5 m and dip between 45° and 85°. This lode geometry is favourable for longhole CRF stoping using mechanised mining techniques.

The sub-level spacing of 10 m floor to floor (7.2 m backs to floor) has been established to ensure stable spans, acceptable drilling accuracies and blast hole lengths. A sub-level spacing of 15 m has been trialled for one level in the Cuffley South area. This involved drilling up from the lower level to 8 m and drilling and firing the remainder from the upper level using downholes. While successful, the trial is limited to this one area at this point.

The production cycle for longhole CRF stoping, as illustrated in Figure 16-3, comprises the following steps:

• Develop access to the orebody.

• Establish bottom sill drive and upper fill drive.

• Drill production blast holes in a staggered ‘Dice 5’ pattern depending on ore width; the nominal stope design width is 1.2 m.

• Blast 2.5 m strike length of holes and extract ore.

• Place rock bund at brow of stope and place rock tube in stope; the rock tube is tightly rolled steel mesh placed in leading edge of stope prior to filling and eliminates the need to bore reamer holes in next stoping panel.

• Place CRF into stope.

• Remove rock bund at brow of stope.

• Commence extraction of adjacent stope once CRF has cured.

Figure 16-3: Longhole CRF stoping method Source: after Potvin, Thomas, Fourie, 2005.



The uphole stope geometry is engineered to minimise the need to use tele-remote loaders for ore mucking and the mining cycle comprises the following steps:

• Drill up to 10 m blind production longholes for a strike length of 3 m.

• Blast holes and extract ore (use tele-remote loader once brow exposed).

• Leave a 3 m rib pillar where required by ground conditions.

• Commence next stope.

16.1.6 Materials Handling All underground ore is trucked to surface via the Augusta decline and dumped in the box-cut. A private contractor rehandles the ore and transfers it to the Brunswick ROM pad where it is stockpiled, screened, blended and crushed prior to be being fed into the Brunswick Processing Plant.

All other waste is trucked to surface and stockpiled at the Augusta Waste Rock Storage Facility or used in the CRF backfill for stopes. A portion of suitable material is screened and utilised underground as road base.

16.4 Mine Design Guidelines The mining schedule follows a bottom-up sequence, mining from the northern and southern extents retreating toward the central access. This sequence enables a consistent production profile to be maintained as it allows for dual development headings on each level.

The presence of the Cuffley Incline has allowed the bottom-up mining sequence to commence from the 935 mRL with no crown pillars planned in between this level and the 1035 mRL. Elsewhere, crown pillars are formed at every fourth level, allowing the stoping panel to consist of three CRF stoping blocks.

16.1.7 Level Development Production drive development is carried out so as to ensure the ore is positioned in the face and the hanging wall exposure is minimised. All production development is completed under geology control. Production drives are excavated and supported by handheld mining methods that have been proven to be generally stable and productive during the past 24 months of operation.

16.1.8 Vertical Development Ventilation rises of 3.5 x 3.5 m have been excavated between levels to extend the existing primary exhaust system both above and below the Cuffley exhaust shaft bottom. Ladder rising has been utilised for installation of escapeways providing a second means of egress.

16.1.9 Stoping A default strike length of 10 m is assumed for all stopes. Material is assumed to have a swell factor of 30% and non-mineralised material is allocated a default relative density of 2.72 t/m3. The relative density of mineralised material is estimated in the Mineral Resource model.

16.1.10 Mine Design Inventory The planned stope and development inventory for the mine is summarised in Table 16-1 and a breakdown of major lodes by level is shown in Table 16-2.



Table 16-1: Stope and Development Inventory by lode

Level Development Stoping Total

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

AS 7,857 3.3 2.8 25,071 4.2 3.4 32,928 4.0 3.3

B 3,592 9.0 1.3 24,096 7.2 1.9 27,688 7.5 1.9

BSP 150 2.5 1.6 450 1.7 1.6 600 1.9 1.6

CD 8,563 7.3 4.3 66,235 6.4 3.8 74,798 6.5 3.8

CE 1,026 5.3 1.7 5,663 9.0 2.4 6,689 8.4 2.3

CM 4,327 5.4 1.7 73,933 11.3 4.2 78,260 10.9 4.0

CS 150 2.3 1.8 1,246 1.8 2.0 1,396 1.9 2.0

CSE 151 2.9 2.2 1,028 3.4 2.5 1,179 3.4 2.5

E 2,123 2.7 2.0 27,663 3.5 2.4 29,786 3.4 2.4

K 1,048 2.4 1.4 23,159 2.4 1.8 24,207 2.4 1.8

NE 3,470 3.6 2.7 15,065 3.7 3.0 18,535 3.7 2.9

NM 32,474 6.1 2.3 170,681 6.9 2.9 203,155 6.7 2.8

NSP2 56 2.9 0.8 683 2.6 1.0 739 2.7 1.0

NSP3 0 0.0 0.0 2,252 3.5 1.7 2,252 3.5 1.7

NSP4 2,860 2.6 2.2 8,497 3.1 2.8 11,357 3.0 2.6

NV43 12,618 5.1 1.0 40,072 6.4 1.2 52,690 6.1 1.2

NV44 0 0.0 0.0 1,217 2.6 1.0 1,217 2.6 1.0

NW 0 0.0 0.0 145 4.3 3.4 145 4.3 3.4

P2 3,515 3.4 1.7 8,873 4.4 2.2 12,388 4.1 2.0

W 3,091 5.4 2.7 13,658 6.1 3.1 16,749 6.0 3.0

Total 87,071 5.4 2.2 509,687 6.7 3.0 596,758 6.5 2.9

Table 16-2: Development and Stoping Inventory by level for major lodes (CE, CS, CSE, NSP2-4, NV44, NW, P2 and W lodes not included)


Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

AS Lode 7,857 3.3 2.8 25,071 4.2 3.4 32,928 4.0 3.3

954 605 4.2 3.0 1,820 6.8 4.7 2,425 6.1 4.3

964 907 6.1 2.8 2,612 6.9 3.9 3,519 6.7 3.6

974 910 3.9 3.2 2,596 5.2 4.1 3,506 4.8 3.9

984 908 3.6 2.9 2,784 4.5 4.4 3,692 4.3 4.0

994 1,211 2.8 3.0 4,893 3.4 3.1 6,104 3.3 3.1

1004 1,358 2.6 2.5 4,656 3.3 3.2 6,014 3.1 3.0

1014 1,284 2.2 2.7 3,805 3.0 2.9 5,089 2.8 2.8

1024 674 1.9 2.6 1,905 2.8 1.6 2,579 2.5 1.9

B Lode 3,592 9.0 1.3 24,096 7.2 1.9 27,688 7.5 1.9

939 450 7.7 1.5 3,683 7.1 2.4 4,133 7.2 2.3

949 754 7.1 2.7 4,134 5.4 2.3 4,888 5.7 2.4

959 0 0.0 0.0 169 2.2 2.4 169 2.2 2.4

969 0 0.0 0.0 1,580 7.9 2.5 1,580 7.9 2.5

979 150 1.7 2.0 3,631 4.1 2.1 3,781 4.0 2.1




Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

989 0 0.0 0.0 4,602 4.2 2.1 4,602 4.2 2.1

999 895 11.2 0.5 1,980 17.1 0.7 2,875 15.2 0.6

1005 743 11.1 0.4 2,044 13.3 0.8 2,787 12.8 0.7

1019 300 8.0 1.2 1,235 8.3 1.7 1,535 8.3 1.6

1029 300 8.3 1.3 1,038 7.4 1.5 1,338 7.6 1.5

CM Lode 4,327 5.4 1.7 73,933 11.3 4.2 78,260 10.9 4.0

804 896 5.5 0.9 3,930 8.3 1.9 4,826 7.8 1.7

814 1,656 5.9 2.1 4,248 6.8 3.0 5,904 6.6 2.7

824 600 4.4 1.4 6,200 7.1 2.5 6,800 6.9 2.4

834 96 11.5 1.7 6,957 8.8 3.4 7,053 8.8 3.4

844 26 2.2 1.2 7,316 11.2 5.5 7,342 11.1 5.5

854 299 4.1 1.2 8,429 10.4 4.4 8,728 10.1 4.3

864 301 4.1 1.9 10,886 15.7 6.8 11,187 15.3 6.7

874 0 0.0 0.0 373 1.7 1.6 373 1.7 1.6

884 0 0.0 0.0 6,959 17.9 5.7 6,959 17.9 5.7

894 0 0.0 0.0 17,433 10.9 3.1 17,433 10.9 3.1

904 0 0.0 0.0 705 9.9 0.6 705 9.9 0.6

954 453 5.0 2.7 497 8.9 4.6 950 7.1 3.7

964 0 0.0 0.0 0 0.0 0.0 0 0.0 0.0

E Lode 2,123 2.7 2.0 27,663 3.5 2.4 29,786 3.4 2.4

1 0 0.0 0.0 1,929 5.4 2.0 1,929 5.4 2.0

2 0 0.0 0.0 119 8.6 3.4 119 8.6 3.4

5 0 0.0 0.0 354 14.8 4.5 354 14.8 4.5

6 0 0.0 0.0 1,671 4.8 4.4 1,671 4.8 4.4

980 0 0.0 0.0 625 1.4 1.8 625 1.4 1.8

1000 192 2.0 1.5 3,352 1.9 1.6 3,544 1.9 1.5

1020 621 2.6 1.8 6,304 3.0 2.3 6,925 3.0 2.3

1034 1,310 2.9 2.2 5,600 2.8 2.2 6,910 2.8 2.2

1040 0 0.0 0.0 2,022 4.0 3.2 2,022 4.0 3.2

1060 0 0.0 0.0 1,974 3.8 2.9 1,974 3.8 2.9

1070 0 0.0 0.0 3,713 3.6 2.4 3,713 3.6 2.4

K Lode 1,048 2.4 1.4 23,159 2.4 1.8 24,207 2.4 1.8

1 0 0.0 0.0 2,588 2.2 1.9 2,588 2.2 1.9

2 150 1.6 1.5 3,727 2.1 1.9 3,877 2.0 1.9

3 0 0.0 0.0 1,466 2.1 2.0 1,466 2.1 2.0

1040 0 0.0 0.0 982 3.3 2.0 982 3.3 2.0

1050 0 0.0 0.0 983 3.9 2.2 983 3.9 2.2

1069 0 0.0 0.0 0 0.0 0.0 0 0.0 0.0

1079 0 0.0 0.0 1,377 2.0 1.3 1,377 2.0 1.3

1089 448 2.4 1.1 2,069 2.8 1.5 2,517 2.8 1.4

1099 150 3.5 1.5 3,103 2.5 1.5 3,253 2.5 1.5

1109 150 2.5 1.4 3,333 2.2 1.7 3,483 2.2 1.7




Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

1119 150 2.2 1.9 2,393 2.2 1.9 2,543 2.2 1.9

1129 0 0.0 0.0 1,138 2.1 2.0 1,138 2.1 2.0

NE Lode 3,470 3.6 2.7 15,065 3.7 3.0 18,535 3.7 2.9

989 0 0.0 0.0 0 0.0 0.0 0 0.0 0.0

990 0 0.0 0.0 0 0.0 0.0 0 0.0 0.0

1050 901 3.5 1.9 2,166 2.1 1.6 3,067 2.5 1.7

1069 0 0.0 0.0 2,007 2.0 1.7 2,007 2.0 1.7

1079 0 0.0 0.0 1,272 2.1 1.9 1,272 2.1 1.9

1089 150 1.9 1.7 1,136 2.5 2.2 1,286 2.4 2.1

1099 451 2.7 2.2 1,180 4.4 3.6 1,631 3.9 3.3

1109 453 3.4 2.9 1,176 4.8 4.1 1,629 4.4 3.7

1119 606 3.8 3.1 2,829 5.1 4.0 3,435 4.8 3.8

1129 909 4.4 3.4 3,299 5.3 3.9 4,208 5.1 3.8

NM Lode 32,474 6.1 2.3 170,681 6.9 2.9 203,155 6.7 2.8

1 0 0.0 0.0 1,771 6.4 2.8 1,771 6.4 2.8

2 753 4.4 2.3 3,468 3.8 3.0 4,221 3.9 2.9

3 0 0.0 0.0 4,288 3.9 2.8 4,288 3.9 2.8

829 513 3.1 0.7 1,301 4.8 0.9 1,814 4.3 0.9

839 298 4.7 0.9 1,299 3.4 1.9 1,597 3.6 1.7

849 2,119 6.2 3.1 13,218 7.1 4.0 15,337 6.9 3.8

859 1,676 8.0 2.8 11,480 6.9 2.7 13,156 7.0 2.7

860 971 5.9 2.4 2,208 4.9 2.1 3,179 5.2 2.2

869 0 0.0 0.0 4,677 6.2 2.7 4,677 6.2 2.7

870 231 3.4 1.2 588 3.4 1.4 819 3.4 1.3

879 899 5.8 1.8 4,013 5.2 1.1 4,912 5.3 1.2

889 0 0.0 0.0 967 6.6 3.3 967 6.6 3.3

899 0 0.0 0.0 1,576 5.7 3.4 1,576 5.7 3.4

909 0 0.0 0.0 2,487 5.9 2.9 2,487 5.9 2.9

969 0 0.0 0.0 5,367 7.4 2.9 5,367 7.4 2.9

979 0 0.0 0.0 4,058 6.9 3.9 4,058 6.9 3.9

980 149 4.0 0.7 7,722 7.9 2.9 7,871 7.8 2.9

990 603 5.7 2.5 7,060 10.1 3.5 7,663 9.7 3.4

1000 2,618 7.9 2.6 7,600 8.7 3.3 10,218 8.5 3.1

1010 3,341 5.1 2.5 8,642 6.5 3.3 11,983 6.1 3.1

1020 3,330 5.0 2.6 10,979 5.4 2.7 14,309 5.3 2.7

1029 0 0.0 0.0 937 4.1 2.8 937 4.1 2.8

1030 2,029 3.4 1.7 7,076 4.5 2.1 9,105 4.2 2.0

1040 2,697 4.4 1.5 13,432 6.2 2.6 16,129 5.9 2.4

1050 1,796 3.6 1.1 13,897 5.6 2.4 15,693 5.4 2.3

1069 302 6.1 2.7 5,363 6.5 3.7 5,665 6.4 3.7

1079 1,510 4.4 2.5 5,445 6.6 3.1 6,955 6.1 3.0

1089 1,055 7.8 2.3 3,828 10.7 3.1 4,883 10.0 2.9




Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

Inventory (t)

Grade Au (g/t)

Grade Sb (%)

1099 1,507 7.9 2.3 4,064 10.6 3.0 5,571 9.9 2.8

1109 1,359 10.3 2.5 4,055 14.3 3.3 5,414 13.3 3.1

1119 1,360 12.3 2.7 3,521 12.5 3.4 4,881 12.5 3.2

1129 1,358 7.2 2.6 4,294 4.2 3.2 5,652 4.9 3.0

NV43 Lode 12,618 5.1 1.0 40,072 6.4 1.2 52,690 6.1 1.2

990 745 9.2 0.6 2,604 9.1 0.7 3,349 9.1 0.7

1000 1,643 4.2 0.9 5,562 4.7 1.2 7,205 4.6 1.1

1010 1,792 4.1 0.9 6,573 4.7 1.2 8,365 4.5 1.2

1020 2,092 4.8 1.1 5,812 6.0 1.3 7,904 5.7 1.3

1030 2,092 4.8 1.1 7,594 6.2 1.4 9,686 5.9 1.3

1040 2,243 5.4 1.1 5,381 6.9 1.3 7,624 6.5 1.2

1050 2,011 5.7 1.0 6,546 8.6 1.2 8,557 7.9 1.2

Total 67,509 5.4 2.0 399,740 6.9 2.9 467,249 6.6 2.7

16.5 Ventilation

16.1.11 Ventilation Circuit The current mine ventilation circuit comprises fresh air being sourced from three surface intakes – the Augusta portal, the Augusta ladderway and the Augusta Fresh Air Rise (FAR). Exhaust ventilation flow exits the active mine workings via two return airways being the Cuffley Return Air Raise (RAR) and Augusta Magazine exhaust rise.

Production from the Augusta decline has ceased and primary ventilation in this area has been reduced to a level that provides for the inspection of infrastructure. Fresh air travels to the bottom of the Augusta workings via internal rises and enters at the 900 mRL, at which point it flows up the Augusta Decline and enters the Cuffley Decline where it joins the primary flow that is distributed throughout the Cuffley workings. The circuits for Augusta and Cuffley are presented in Figure 16-4 and Figure 16-5.

The following specifications apply to the existing Augusta and Cuffley ventilation rises:

• Augusta Ladder Rise (surface to 900 mRL), 2.4 m diameter

• Augusta FAR (1020 mRL to surface), 3.0 m diameter

• Augusta Magazine Exhaust (1020 mRL to surface), 1.2 m diameter

• Cuffley RAR (950 mRL to surface), 3.0 m diameter

• Cuffley RAR (above the 955 mRL), 3.5 m x 3.5 m diameter

• Cuffley RAR (below the 955 mRL), 3.5 m x 3.5 m diameter.

Three single stage 110 kW axial fans have been built into a bulkhead at the 950 mRL and act as the Cuffley primary ventilation fans. The Cuffley primary ventilation fans have been designed with a final duty of 140 m³/s.

The existing Cuffley primary fans are three Clemcorp CC1400 MK4 single stage 110 kW axial fans installed in a bulkhead on the 950 mRL. The operating parameters of each fan are:

• Lower operating fan static pressure of 650 Pa at 50 m3/s

• Higher operating fan static pressure of 2,600 Pa at 30 m3/s.



At present, all three primary fans are operating at a measured pressure of 1,077 Pa for a combined volume capacity of 133 m3/s. These fans have sufficient capacity to support any further expansion of capital access development.

The exhaust air from the underground explosives magazine exits the mine through a 1,200 mm Protan ventilation ducting that has been installed in the Augusta FAR and exhausts to the surface away from the Augusta FAR.

Based on a current primary flow of 133 m3/s, the main Cuffley decline airflow has a maximum velocity of 5.6 m/s, which is below the upper design limit of 6 m/s, the point at which dust would become airborne. The Cuffley RAR has a velocity of approximately 17.9 m/s, which is at the middle range of velocities at which main exhaust shafts should be operated.

Figure 16-4: Augusta Ventilation Circuit



Figure 16-5: Cuffley Ventilation Circuit

16.6 Backfill The practice of placing CRF in stope voids has been undertaken at Cuffley to improve local ground stability, reduce unplanned dilution and improve mining recoveries. The CRF uses waste rock sourced from development with the addition of a cement slurry mix that results in a final product composed of 4% cement.

CRF is mixed in batches of varying sizes with a Caterpillar 1700G loader. The cement slurry is delivered underground to mixing bays via a cement agitator truck. The hydrated cement mix is batched on surface using a cement silo on hire from the cement supplier. The quality of the CRF is ensured by use of PLC control of the cement batching plant and standardised bucket filling of the waste rock. Records of batch quantities for all batches are kept.

Emergency dump and wash-out areas are located underground should a load of batched cement need to be disposed of before curing in the agitator bowl occurs.

Once mixed, the CRF is trammed to the fill point of the open stope using a Toro 151 (or equivalent) loader. A bund is placed at an appropriate distance from the top of the stope to minimise the potential for a loader to overbalance into the stope void. During placement of the CRF, care is taken that the rock-tube, which is secured by chains during the filling process, is not displaced.

The nominal curing time is 12 hours and after approximately eight hours, the rock bund placed at the brow of the stope is removed.

CRF has proved effective in minimising dilution during subsequent panel extraction; in addition, CRF provides better ground stability and has eliminated the requirement for rib pillars.



16.7 Mineral Reserve Schedule

16.1.12 Development Schedule During 2016, ore development has transitioned from hand-held development to mechanised development boring with hand-held bolting. This involved a successful trial of a Resemin Muki drill owned by Pybar and the ongoing use of UD09 (Mandalay development drill) for approximately 80% of ore development faces. The Muki rig has since been altered to allow for boring and bolting within a 1.8 m wide ore development drive. The Muki rig is expected to begin operation at Costerfield underground during April 2017. The rate of mechanised boring is higher than that of hand-held boring, and requires fewer production crews to meet the required target. This has been factored into development cost and rates.

Development rates are based upon the assumptions as shown in Table 16-3. The production development rate is less than the waste development rate as it allows time for face sampling and poorer ground conditions.

Table 16-3: Development Rates

Description Value

Production crews (#) 3

Advance/cut (m) 1.8

Cuts/face/week (#) 5

Advance/face/week (m) 9

Available faces/crew/week (#) 3

m/week/crew (m) 27

Total m/week (m) 81.0

Waste crews (#) 1

Advance/cut (m) 1.8

Cuts/week (#) 14

Total m/week (m) 25.2

Weeks/mth (#) 4.3

m/Mth (production) (m) 348

Waste/mth (m) 108

Total m/mth (m) 457

16.1.13 Equipment Requirements The existing development, production and auxiliary underground equipment fleet will continue to be used (where applicable), with additional equipment purchased as required to meet the planned replacement schedule or meet increased production demands. The single-boom Muki drill with capacity of bolting and boring headings has been included in the Reserve assumptions. The existing mobile equipment fleet is summarised in Table 16-4.



Table 16-4: Underground Mobile Equipment Fleet

Equipment Type Existing Fleet

Single-Boom Jumbo 1

Production Drill 2

LHD - 1700 Loader 1

LHD - 151 Loader 7

Haulage Trucks 2

Charge Rig 1

Tele-handler 1

Service Tractor 4

Light Vehicles 10

Total 27

16.1.14 Personnel An existing core group of management, environmental, technical services (engineering/ geology), administration, maintenance, supervisory, and production personnel will continue to operate the Augusta site. As a residential operation, all employees commute daily from their place of residence.

Shift Schedule Costerfield operates a continuous mining operation, 24 hours a day, 365 days per year.

Operators and maintenance personnel work seven days on, seven days off, on 11-hour shifts, alternating between dayshift and nightshift.

Augusta support staff work a standard Australian working week of five days on, two days off, eight hours per work day.

All on-costs for annual/ sick leave and training have been estimated in the direct and indirect operating costs respectively.

Personnel Levels All equipment has been assigned with one operator per crew per machine. It is assumed that cross-training will occur for all operators, ensuring that each shift panel is adequately multi-skilled so as to relieve for sickness, annual leave and general absenteeism.

The current total workforce is 200 employees, 139 of whom work in the mining department.

16.8 Schedule Summary A summary for the key physicals in the Mineral Reserve schedule is presented in Table 16-5.

Table 16-5: Summary of design physicals

Description Units Quantity

Operating Development (Waste) m 7,829

Handheld Operating Development (Ore) m 5,778

Development Ore Tonnes t 87,071

Development Ore Grade Au g/t 5.4

Development Ore Grade Sb % 2.2

Stoping Ore Tonnes tonnes 509,687




Stoping Ore Grade Au g/t 6.7

Stoping Ore Grade Sb % 3.0

Total Ore Tonnes t 596,758

Total Ore Grade Au g/t 6.5

Total Contained Au oz 124,710

Total Ore Grade Sb % 2.9

Contained Sb t 17,305

Opening Stocks

ROM Ore Tonnes t 24,508

ROM Ore Grade Au g/t 5.98

ROM Ore Grade Sb % 2.13



17 Recovery Methods 17.1 Brunswick Processing Plant

The Brunswick Processing Plant processes a sulphide gold-antimony ore through a conventional comminution and flotation style concentrator. It has been operating since 2007, and has been run by Mandalay since late 2009. Since late 2009, several processing plant upgrades have seen production increase from approximately 2,000 t/mth to the current average of +12,800 t/mth over the 2015 and 2016 calendar years. The concentrator operates 24 hours per day, 7 days per week. Crushing operates under noise restriction guidelines.

The surface crushing and screening system processes underground feed down to a particle size range suitable for milling through a two stage, closed circuit ball milling circuit. Centrifugal style gravity concentrators are used on the combined primary milling product and secondary mill discharge, to recover a gold-rich gravity concentrate that is tabled and sold as a separate gold concentrate product, which is sent to a local refinery. Secondary milled products are classified by size and processed through a simple flotation circuit comprising a single stage of rougher, scavenger and cleaning. The concentrate is dewatered through thickening and filtration, to produce a final antimony/ gold concentrate product that is then bagged, packed into shipping containers and shipped to customers in China. The tailings are thickened before being pumped into a TSF to the east of the Brunswick Processing Plant. The flowsheet is simple and conventional, and is suited to processing the Costerfield ores in the LOM plan. With the addition of a mobile crushing unit in 2012, and gradual production creep since then, the capacity of the plant has been successfully upgraded to over 13,000 t/mth. A summary processing flowsheet is provided in Figure 17-1. A more detailed processing description is provided below.

Crushing and Screening Circuit The crushing and screening plant consists of a primary crushing circuit in closed circuit with a 15 mm vibrating screen. It uses one duty and one standby diesel-powered Finlay I-130RS mobile impact crusher. The crushed ore is conveyed to two x 120 t fine ore bins in parallel. The old conveyor system through the fixed crushers was bypassed in 2016, with the installation of a new conveyor system from the ROM bin directly to the fine ore bins.

Milling Circuit The fine ore is reclaimed from the crushed ore bins, which both discharge onto the primary mill feed conveyor as feed to the milling circuit. The milling circuit comprises two ball mills in series, both in closed circuit. The primary mill operates in closed circuit with a “DSM” screen, with screen oversize returning to the primary mill and undersize being fed to a centrifugal style gravity concentrator. This recovers a small mass of high grade gold concentrate that is sent to the gold room for gravity tabling. The final gravity concentrate is sent directly to a refinery as a separate saleable product. Gravity production varies, but typically recovers around 30 - 40% of the gold in the feed. The gravity tailing is pumped to classifying cyclones, the overflow of which becomes the flotation plant feed, while the underflow is returned to the secondary ball mill for further size reduction. The secondary ball mill discharge is combined with the DSM screen undersize as feed to the centrifugal gravity concentrator. The current milling circuit capacity is capable of handling in excess of 13,500 t/mth in some months.

Flotation Circuit The flotation circuit is designed to recover an antimony/ gold concentrate.



The flotation circuit is fed from the secondary ball mill cyclone overflow. The cyclone overflow is fed to a conditioning tank where lead nitrate, an activator and potassium amyl xanthate (PAX), a collector, are added. The conditioning tank feeds two site-fabricated ‘rougher tank’ flotation units in series. The two rougher tank cell concentrates combine with the final cleaner concentrate in the concentrate thickener as the final product. The rougher tank cell tailings flow to the Denver rougher flotation cells.

The rest of the flotation circuit consists of eight Denver square DR100 cells for the remaining rougher and scavenging duties, and six Denver square DR15 cells for cleaning duties. The concentrate from the rougher flotation cells is pumped to the cleaner flotation cells while the tailing becomes feed for the scavenger flotation cells. The concentrate from the scavenger flotation cells is recycled to the feed of the rougher flotation cells, while the tailing is pumped to the tailings thickener. The concentrate from the cleaner flotation cells is pumped to the concentrate thickener, while the tailing is recycled to the rougher flotation cells.

The flotation circuit effectively recovers gold not collected in the gravity gold circuit, i.e. gold recovery is robust even with variability in gravity recovery.

Concentrate Thickening and Filtration The final concentrate is then pumped from the concentrate thickener directly to a plate and frame filter. The moist filter cake is discharged directly into concentrate bags. The filtrate is recycled to the concentrate thickener, while the concentrate thickener overflow is recycled through the plant as process water to maximise water re-use and minimise concentrate losses.

Tailings Circuit The flotation tailings are settled in a thickener. Overflow is recycled through the plant as process water and the thickened solids are pumped into a TSF. The tailings discharge to the TSF is managed via a conventional spigot system. Additional water from the tailings is decanted and pumped back to the plant for use as process water.

Throughput The Brunswick mill capacity is typically between 12,000 and 13,500 dry t/mth. The plant has demonstrated ongoing production throughput creep through continual improvements over the last several years, increasing from around 5,000 t/mth in January 2012.

Plant throughput budgeted for 2017 is 12,833 t/mth. This is based on the 2015 throughput of 12,822 t/mth and the 2016 throughput of 12,867 t/mth. The historical and forecast LOM plant throughput on the current ore feed blend is discussed in more detail in Section 13.

Recovery The 2016 reconciled plant recoveries were 95.4% Sb and 90.1% Au. The forecast LOM model recoveries are 92.9% Sb and 81.2% Au for the LOM to 2021. Both are considered to be reasonable recovery estimates given that both are supported by historic recoveries at a similar feed grades and based on grade/ recovery relationships. Further confidence is provided by the consistent recoveries of both Sb and Au across a range of feed grades. The historical and forecast plant LOM recoveries on the LOM feed blend are discussed in more detail in Section 13.

Concentrate Grade The current concentrate antimony grades have remained stable at 54% Sb. Supporting historical plant throughput and recovery data is provided in Section 13.



17.2 Services

Water The water services at the Brunswick Processing Plant consist of the raw water, process water and excess water disposal systems. The process water supply consists of concentrate thickener overflow, tailings thickener overflow and TSF decant return water. Most of the raw make-up water is provided by dewatering of the underground operations (approximately 1.3 ML/day). The plant operates with a positive water balance, with excess water being disposed of. Mandalay constructed a 2 ML/day permeate RO plant at the Brunswick mill in 2014. The 2 ML/day plant has remained in operation as per regulatory approvals. Construction of the Splitters Creek evaporation facility started in the fourth quarter of 2014 and was completed in 2015. This has the capacity to treat 104 ML/year net (evaporation minus rainfall).

The TSF and process water is stored in and distributed from a dedicated tank system. As the site is in a positive water balance, adequate process water supplies are available to meet the LOM requirements.

Air The Brunswick Process Plant requires both low and high pressure air supplies. Currently, three separate low pressure blowers supply the rougher, scavenger and cleaner cells, with existing tank cells running off high-pressure air. The high-pressure air supply was upgraded in 2013, to increase the capacity and availability of high pressure air. The pressure filter runs off high pressure air. The processing facility has adequate air to meet the LOM requirements.

Power Electrical power is currently supplied to the Brunswick Processing Plant and associated mining infrastructure from the Powercor Bendigo Zone Substation. The Brunswick supply consists of 1,000 kVA at 240/415 V and is sufficient for the operational requirements. The current calculated peak load is estimated to be under 1,000 kVA. The main Brunswick Processing Plant MCC was replaced in 2016 and now has the capability to be partly or fully supplemented with generator power if this is required in the future.

The RO plant is powered via diesel generators. These generators will continue to be its power source in the future. The processing facility has adequate electrical power supplies to meet the LOM requirements. It can be simply expanded by adding additional generators in the future, if required.


FAIR/KENT/WALS/EBBE/swee PLI022_Costerfield_Operations_NI_43 101_Dec 2016_Rev1 17 March 2017

Figure 17-1: Brunswick Processing Plant summary flowsheet



17.3 Plant Upgrades No major capital expansion projects are planned. Specific process plant sustaining capital project costs and a TSF lift have been assigned to the financial model in 2017, and an appropriate allowance for ongoing sustaining capital for the plant and tailings have been assigned for the following years of the LOM.

Crushing and Screening Circuit The initial trial of a mobile crusher in September 2012 significantly improved the capacity of the plant. A larger portable crushing unit is now a permanent part of the process flowsheet configuration, and enables an average throughput of over 12,800 t/mth. Another mobile crusher was purchased in 2015 to allow for a duty and standby arrangement to further increase the availability of the crushing circuit.

By utilising the existing crushing plant conveyors and fine ore bins, the crushed product from the portable crushing plant can be transferred at a finer feed size into the ore bins twice a shift, to provide consistent feed to the existing milling circuit. This system was simplified in Q4 2016 with the installation of a single conveyor to directly fill the fine ore bins, rendering four old conveyor belts redundant.

Milling Circuit The milling circuit remains unchanged. The finer crushed ore feed size allows the target average throughput of over 12,800 t/mth to be achieved. No further major work is planned at this time.

Flotation Circuit Ongoing optimisation of this circuit via reagent testing is ongoing. No major upgrade work is currently planned.

Concentrate Thickening and Filtration The existing concentrate thickening and filtration circuit is adequate for the target throughput. No further major work is planned at this time.

Tailings Circuit The tailings thickener has sufficient capacity to handle the current throughput. The average tails thickener underflow solids density has been maintained at around 50% (+/-10%).

The currently utilised Brunswick TSF will provide sufficient tailings storage capacity until June/ July 2017. Additional TSF storage capacity (150,000 m3 of volume) to be provided during 2017, with construction on a lift on the existing Brunswick TSF to be completed by Q2 2017. It is expected this capacity will be fully utilised by December 2018.

Planning approval has been granted for the Bombay facility for a 327,000 m3 total increase (comprising a 183,600 m3 lift and a separate 143,400 m3 lift).

The total approved increase in tailings storage (Brunswick and Bombay) allows for over three years’ capacity at current milling rates.

The Bombay TSF lift of 183,600 m3 is the next to be built. This will provide an additional 22 months of capacity.



Recovery Projects No major recovery projects are planned. Ongoing optimisation work will be undertaken as normal practice. All improvements planned for the existing circuit will be to maintain current levels of recovery or marginally improve it.

Reagent Mixing and Storage No upgrade work is required for the reagent mixing and storage area.



18 Project Infrastructure 18.1 Surface Infrastructure

The Costerfield Operation’s surface facilities are representative of a modern gold-antimony mining operation. The Augusta Mine site (Figure 18-1) comprises the following:

• Office and administration complex, including change house

• Store and laydown facilities

• Heavy underground equipment workshop

• Evaporation and storage dams

• Temporary surface ore stockpiles and waste stockpile area

• Augusta Mine box-cut and portal

• Cement silos

• Ventilation exhaust raise

• Ventilation intake raises

• Augusta Mine dam to manage rainfall runoff.

Figure 18-1: Augusta Mine Site



The Brunswick site (Figure 18-2) comprises the following:

• Gold-antimony processing plant and associated facilities

• Central administration complex

• Process plant workshop

• Tailings storage facilities

• ROM stockpiles

• RO Plant capable of producing 2 ML of treated water per day

• Previously mined Brunswick Open Pit

• Core farm and core processing facility.

Figure 18-2: Brunswick Site Area

The Splitters Creek Evaporation Facility is situated on a 30 ha parcel of land located approximately 3 km from the Augusta site. The facility exists on a new Mining Licence (MIN 5567). The facility evaporates groundwater extracted from the Costerfield Operations, enabling underground dewatering rates to be maintained. The site comprises the following:

• 150 ML Storage Dam

• 40 ML Evaporation Terraces



• Recirculation pump system that directs water from the Storage Dam to the Evaporation Terraces

• Splitters Creek pipeline that feeds water from the Augusta Mine Dam to the Evaporation Terraces

• Leakage detection system on the Splitters Creek pipeline.

18.2 Underground Infrastructure

Secondary Means of Egress The secondary means of egress consists of a ladderway system that extends from surface to the 900 mRL within the Augusta underground workings and a ladderway system in the Cuffley workings that extends from the Cuffley Incline, Cuffley Decline and 4800 Decline to the 945 mRL level. From the 945 mRL level, extraction is performed via an Emergency Gig. The Emergency Gig attaches to a standard crane hook and hoists personnel, in an emergency, up and down the Cuffley RAR using a 200 t mobile crane as the hoist. The Emergency Gig is capable of evacuating five persons or 600 kg (120 kg per person) at a time.

Refuge Chambers Underground refuge chambers are installed in response to hazards posed by irrespirable atmospheres. The Augusta decline has fresh air bases located at 3 Level and 1040 Level. The Cuffley workings have refuge chambers at the 970 mRL, 864 mRL, 1105 mRL, 1015 mRL and 909 mRL. The refuge chambers are configured as either 4-, 16- or 20-man capacity. The capacity of the refuge chamber required is dictated by the number of personnel planned to be working in the immediate vicinity serviced by the refuge chamber.

The position of the refuge chamber facilities enables all personnel to be within 750 m of a refuge chamber, as recommended in the Western Australian ‘Refuge Chambers in Underground Metalliferous Mines’ Guideline (Department of Consumer and Employer Protection, 2008).

It is not intended for refuge chambers to substitute a secondary means of egress, but to provide refuge during fire or containment when ladderways may be inoperable or inaccessible.

Compressed Air The existing compressed air plant comprises three 593 cfm compressors. The overall plant capacity is 840 L/s (1,779 cfm).

Compressed air is delivered underground via a 4-inch HDPE poly pipe. Each level is then supplied from the decline via 2-inch HDPE poly piping. Air receivers have been placed at 909 mRL, 1105 mRL, 864 mRL and 1015 mRL to increase the system efficiency. Compressed air is used to power pneumatic equipment and/ or activities, including:

• Airleg drills

• Kempe exploration drill rig

• Pneumatic ammonium nitrate-fuel oil (ANFO) loaders

• Blast hole cleaning for development rounds

• Diaphragm air pumps

• Pneumatic longhole drills

• Longhole cleaning.



Ventilation System The primary ventilation infrastructure consists of two exhaust shafts known as the Augusta RAR and Cuffley RAR. These two shafts are 3 m in diameter and have been supported by the use of shotcrete. The Augusta RAR extends from surface to the 1020 mRL, with a series of short raises from the 1020 mRL to the 900 mRL. The Cuffley RAR extends from surface to the 945 mRL, with a series of short 3.5 x 3.5 m raises extending up to the 995 mRL in the Cuffley Incline and down to the 864 mRL in the 4800 Decline. A small diameter FAR extends from surface to the 1020 mRL.

The entire underground workings are ventilated via three single stage 110 kW axial fans located in a bulkhead at the 945 mRL.

Fresh air is supplied via the decline portal the small diameter Augusta Fresh Air Intake (FAI) and the Augusta FAR. This primary ventilation flow is distributed through the mine using secondary fans positioned in the decline/ incline that force ventilation through the development levels.

The 1020 Magazine is ventilated by a single stage 45 kW fan which forces return air from the Magazine to surface via ducting located in the Augusta FAR.

Dewatering System The process of dewatering in advance of the mining levels is achieved by leaving diamond drill holes drilled from underground open to drain. Due to the fractured nature of the aquifer, the groundwater inflows are unpredictable. Flow rates have continued to increase as the decline advances and more drives are developed.

Removal of the groundwater from the Augusta decline is achieved by a staged pump station system comprising Weartuff WT084 mono pumps located at 890 mRL.

Removal of the groundwater from the Cuffley workings is achieved by a staged pump station system comprising Weartuff WT084 mono pumps located at 849 mRL.

The lower working levels utilise electric submersible pumps to direct water to the mono pumps. Both the 890 mRL and 849 mRL mono pumps bring water to the 945 mRL permanent pump station which pumps water to surface via the Cuffley Rising Main.

This Rising Main extends to the Brunswick Pit which is as a location for settling out of the solids before the clean water is directed to the Augusta Mine Dam for distribution to the Splitters Creek Evaporation Facility, as well as for underground mine water supply.

Infrastructure There is an underground crib room at the 1085 mRL and an underground magazine at the 1020 mRL. In addition to the fixed plant, Mandalay owns, operates and maintains all of the mobile mining equipment, including production drills, loaders, trucks, and ancillary equipment required to undertake ore development and production operations.

18.3 Tailings Storage Since operations began in the 1970s, two tailings dams have been constructed and operated – Bombay Tailings Storage Facility (TSF) and the Brunswick TSF which is currently operational.

Both TSFs were constructed based on a conventional paddock-style/ turkey’s nest design with earthen embankments.

All tailings are currently deposited in the Brunswick TSF. A 2.5 m raise is under construction on the Brunswick TSF, due to be completed by Q2, 2017, which will provide an additional 150,000 m3 of



storage. This capacity will enable tailings to be deposited in the Brunswick TSF until December 2018.

Mandalay has regulatory approval to raise the current Bombay TSF embankment. At current production rates, this embankment raise would provide sufficient tailings storage capacity until the commencement of calendar year 2022.

18.4 Power Supply The Costerfield Operation purchases electricity directly from the main national electricity grid and has connections at the Brunswick Processing Plant, Cuffley Mine and Augusta Mine.

The Costerfield Operation has an existing arrangement for high voltage electrical supply of 2,222 kVA shared between the Augusta Mine and Cuffley Mine and a 1,000 kVA supply at the Brunswick Processing Plant.

Powercor owns and manages the central Victorian electrical distribution infrastructure. ERM Business Energy is the current electrical retailer.

A 1 MVA 4.5 V/ 11 kV step-up transformer is located on Cuffley site to act as an emergency generator backup.

Power Reticulation Power for the Augusta underground operations comes from a high voltage 22 kV Powercor supply that feeds:

• The Augusta surface switch room via a 1 MVA transformer that steps the voltage down to 415 V

• The surface switchyard via a 2.2 MVA transformer that steps the voltage down to 11 kV.

The 415 V supply is stepped up to 1,000 volts via a 750 kVA transformer that feeds the heavy equipment workshop on surface and is also fed underground via a service hole to 3 Level to service all 1,000 volt requirements in the upper levels of the Augusta Mine.

The 11 kV supply is fed underground via a service hole to a 1.5 MVA substation located at the 1020 mRL. From this location, the substation steps the voltage down to 1,000 volts and the power is reticulated to the Augusta Decline working areas via cable and distribution boxes.

The Augusta Mine has low voltage (415 V) power delivered underground from surface to 3 Level through a cased service hole. The 415 V feed also supplies surface facilities such as air compressors, workshop and all office amenities.

Power for the Cuffley underground operations is fed from a high voltage 22 kV Powercor supply that feeds a surface switchyard via a 2.5 MVA transformer that steps the voltage down to 11 kV. This 11 kV is directed underground via a 70 mm cable that extends down the Cuffley RAR and feeds a 1.5 MVA 11 kV substation as well as an adjacent 1.0 MVA 11 kV substation located at the 950 mRL. The 1.0 MVA substation feeds the permanent loads within Cuffley (i.e. primary fans and 945 Pump Station), while the 1.5 MVA substation supplies the secondary ventilation fans, submersible pumps and mobile electrical equipment located within the vicinity of the 950 mRL. A 1.5 MVA substation at the 849 mRL in the 4800 decline is fed via a Ring Main Unit from the 1.0 MVA substation located at the 950 mRL.

From these underground substations, the power (1 kV) is reticulated to the working areas via cable and distribution boxes. To the extent practical, power is reticulated between each level through purpose-drilled service holes as opposed to the decline, thereby reducing the length of cable runs throughout the mine.

A Power Factor Correction Unit (PFCU) installed in 2013 on the 11 kV Augusta mine supply was



relocated to the Cuffley Substation in 2015 to obtain the most benefit from this infrastructure as this substation now supplies the majority of the underground power. The PFCU increases the effective use and availability of power to the Cuffley Mine.

18.5 Water Supply Water for the Augusta underground and surface operations is sourced from the Augusta Mine Dam which is fed from the Brunswick Pit. The Brunswick Pit is supplied with mine water (groundwater inflows and mine process water) via a Rising Main that extends from the Cuffley 945 Pump Station into which all mine water is fed from the underground dewatering system.

An underground recirculation system is used; allowing groundwater inflows from the Augusta mine to be pumped to a sump at 1030 mRL to be circulated through Cuffley workings via existing systems.

The Brunswick Processing Facility sources water from a number of sources, including recycled process water from the Brunswick TSF and water in the Brunswick Pit.

Potable water is trucked to site by a private contractor and is placed in surface holding tanks for use in the change house and office amenities. Potable water for drinking is provided in 15 L containers.

Water disposal is discussed in Section 20.1.2.

18.6 Water Management Groundwater is currently pumped from the underground workings to the Brunswick Pit at a rate of approximately 1.6 ML per day. The water then undergoes a period of settling prior to being distributed for operational needs as well as disposal.

The Brunswick Pit consists of a disused open cut mine located to the western side of the Brunswick Processing Facility. It is used for water storage and has a capacity at freeboard of approximately 160 ML.

The Augusta Evaporation Facility comprises three dams with a total storage capacity of 150 ML. When some smaller catchment and operational dams such as the Mine Dam are included, total site storage capacity is calculated at approximately 318 ML.

The water services at the Brunswick Processing Plant consist of the raw water, process water and excess water disposal systems. The process water supply consists of concentrate thickener overflow, tailings thickener overflow and Brunswick TSF decant return water. The majority of the raw make-up water is sourced from the Brunswick Pit. While the process plant utilises water from a closed circuit, make-up process water is required to supplement water evaporated at the Brunswick TSF.

Total evaporation/ water disposal capacity including discharge of RO treated water and Splitters Creek Evaporation Facility is currently estimated at 713 ML per year, assuming long term average Heathcote climatic conditions.

Total inflows are currently estimated from water balance data to be approximately 660 ML per year plus 52 ML per year of rainfall.

18.7 Waste Rock Storage The 2015 expansion of the Augusta Waste Rock Storage Facility to increase in height by a further 3.0 m provided approximately 70,000 m3 of additional storage capacity. Waste from underground capital development was hauled to the Augusta Waste Rock Storage Facility until August 2015 at which time capital development ceased. Waste rock generated underground from operating development is now being used underground for stope backfilling. Further detail is provided in Section 20.1.3.



18.8 Augusta to Brunswick ROM Pad Transport As the Augusta Mine and Brunswick Processing Plant are divided by the Heathcote–Nagambie Road, all ROM products must be trucked between the two sites by an independent contractor. The road distance between the two sites is approximately 3 km and trucking is undertaken using a fleet of single body road trucks, each of 13 t capacity. Correct load weight is achieved via the use of a load cell system on the contractor’s surface loader.

18.9 Diesel Storage A 60,000 L capacity, self-bunded diesel storage tank exists at the Augusta Mine site. This diesel storage caters for all underground and surface diesel needs at both Augusta and Brunswick.

18.10 Explosives Storage All storage, import, transport and use of explosives is conducted in accordance with the WorkSafe Dangerous Goods (Explosives) Regulations 2011.

Mandalay utilises its own licenced personnel and equipment to handle, store, transport and use explosives on the Augusta site. The designated explosives supplier produces all the explosives products off site. The ANFO is supplied in 20 kg bags, while the emulsion is supplied as a packaged product. ANFO is primarily used for development and production purposes, with emulsion used when wet conditions are encountered.

The current Augusta Magazine is located at the 1020 mRL and is operated under the control of the designated black ticket holder on behalf of Mandalay who is the licensee. Table 18-1 states the current Augusta Magazine licence allowances.

Table 18-1: Current Augusta Licence - Maximum quantities and types of explosives

Class Code Type of Explosive Maximum Quantity

1.1D Blasting Explosives 40,000 kg

1.1D Detonating Cord 10,000 m

1.1B Detonators 21,000 items

18.11 Maintenance Facilities A surface maintenance workshop facility is located adjacent to the box-cut at Augusta. At present, all servicing and maintenance activities are undertaken on surface as no facility exists underground at the Augusta Mine. There is a small maintenance/ boilermaker workshop at Brunswick to assist with undertaking processing plant maintenance activities.

18.12 Housing and Land Mandalay owns six land allotments surrounding the Augusta and Brunswick sites. Of these properties, five have residential dwellings with the remaining five consisting of vacant land. The residential dwellings are used as temporary housing for company employees.

The land allotment located on Peels Lane, Costerfield South, acts an offset area for the Mandalay’s mining and processing activities. It has been identified that the Peels Lane Offset has ‘the potential to generate a total of 4.35 habitat hectares’ and associated large trees (Biosis Research, 2005).

The Peels Lane Offset was purchased as part of the Work Plan for MIN 4644 and acted as an offset for the vegetation loss due to the construction of the Augusta Mine Site. The offset site will also be used to meet the offset requirements for Brunswick.



19 Market Studies and Contracts 19.1 Concentrate Transport

The concentrate is discharged directly into 1.5 t capacity bulk bags ready for transportation by road-train to the Port of Melbourne, for shipping to overseas markets. The average payload of each B-double3 road train is approximately 42 t, and sea shipments are normally scheduled at least once per month on a Cartage, Insurance, Freight (CIF) basis to the destination port.

A third-party trucking company collects the concentrate from the Brunswick site, then transports, stores and loads the concentrate. Logistics and shipping documentation services are provided by Minalysis Pty Ltd.

19.2 Marketing Costerfield is a combined gold and antimony mine, with the business being sensitive to the price of both metals and the AUD:USD exchange rate, as the sale currency is USD. Antimony is not traded on international metal exchanges, with terms and prices being agreed upon between producer and consumer. Pricing is dependent on the quality and form of antimony product sold.

The comments in this section are based on Mandalay’s review of market reports by Roskill, the United States Geological Survey (USGS) and public comments by major consumers such as Campine.

World antimony mine production in 2016 was estimated to have been between 140,000 and 150,000 t of contained antimony. China is the world’s largest producer of antimony, accounting for approximately 75 - 80% of world mine production4,5. Primary antimony mines with no precious metal credits are increasingly becoming uneconomic, including those in China, such that global antimony mine output is now shrinking. Recovery of prices in 2017 has incentivised studies into restarting historic mine production and greenfield exploration globally, but no major new antimony production is expected in the next one to three years.

Antimony is primarily used as a flame retardant and in the production of lead acid batteries, these markets together accounting for nearly 90% of antimony consumption worldwide (refer to Figure 19-1). Antimony consumption began to recover in 2016 following years of weak global economic growth and the substitution of antimony in flame retardant formulations in response to price peaks in the previous cycle. Prices sharply recovered in 2016 and early 2017 (refer to Figure 19-2) in response to both a positive demand environment and shrinking availability of primary feedstocks.

3 A B-double road train consists of a prime mover towing a specialised lead trailer that has a fifth wheel

mounted on the rear, towing another semi-trailer, resulting in two articulation points. 4 Antimony: U.S. Geological Survey, Mineral Commodity Summaries, January 2016,

http://minerals.usgs.gov/minerals/pubs/commodity/antimony/mcs-2016-antim.pdf. 5 China’s 2016 Nonferrous Industrial Output Production Summary, China Ministry of Industry and Information Technology (MIIT), February 4, 2017, www.miit.gov.cn/n1146290/n1146402/n1146455/c5479645/content.html

http://www.miit.gov.cn/n1146290/n1146402/n1146455/c5479645/content.html



Figure 19-1: Estimate of global antimony demand by end-use segment Source: Roskill6, USGS and industry reports.

Figure 19-2: Antimony metal prices 2010 - 2015

The emergence of a supply deficit is forecast in the 2018 - 2022 period, largely due to a further reduction in mine output in China in response to the lack of new discoveries, a determined effort by the Chinese Government to limit environmental damage from smaller producers and continue the crackdown on illegal mining and smuggling (Figure 19-3). Recent closure or downsizing of antimony mines outside China, combined with few new antimony mines being built in the foreseeable future, has positioned the market for a significant shortage moving forward.

The size of the forecast deficit will be primarily dependent on the scale of investment in new mine production worldwide, the pace of metal stock drawdown and global industrial production growth.

6 https://roskill.com/product/antimony-world-market-for-antimony-to-2025-12th-edition/



These market dynamics are likely to result in the antimony price remaining in the range of USD7,000 to USD10,000/t in the coming year due to growing shortages.

Figure 19-3: World antimony production and consumption, 2000 - 2020 (t antimony) Source: 2000 - 2012 Roskill, 2013 - 2022 Penfold.

Critical/Strategic Supply Reviews Various departments within the relevant government authorities in the United States, Britain, the European Union, and Japan have in recent years published analytical reports identifying supply risks to metals deemed as strategic. The thrust of these reports is to focus on metals where China is a dominant producer and as such presents a potential threat to those countries dependent on such supply. Such metals include antimony, tungsten, and the so-called Rare Earth Elements (REE).

• USA: The “2015 Department of Defence Strategic and Critical Materials 2013 Report” foresees a shortfall of 13.118 short tons of antimony in a four-year period, and recommends mitigation options to address this shortfall including strategic stockpiling of ~13,000 short tons.7

• British Geological Survey: Ranked antimony #2 on the “2015 Risk List” given China’s predominance in terms of supply and antimony’s low recycling rates and limited substitutes.8

• European Union: An EU report updated an earlier 2010 report on critical materials, maintaining antimony as one of the key metals with a supply risk and economic importance. The 2013 update forecasts a “large deficit” emerging for antimony by 2015 and continuing in 2020, with antimony the only metal among all critical raw materials to have a large deficit, defined as >10% of the market.9

7 “Strategic and Critical Materials 2015 Report on Stockpile Requirements”, U.S. Department of Defense, Office

of the Under Secretary of Defense for Acquisition, Technology and Logistics, January 2015 (http://www.dla.mil/Portals/104/Documents/StrategicMaterials/Reports/Operations%20Report/2015%20Operations%20Report.pdf)

8 “Risk List 2015- Current Supply Risk Index for Chemical Elements or Elements Groups which are of Economic Value," British Geological Survey (https://www.bgs.ac.uk/downloads/directDownload.cfm?id=3075&noexcl=true&t=Risk%20list%202015)

9 “Report on Critical Raw Materials for the EU”, May 26, 2014 (http://ec.europa.eu/enterprise/policies/raw-materials/files/docs/crm-report-on-critical-raw-materials_en.pdf) and “Critical Materials for the EU,” The European Commission Enterprise and Industry, 30 July 2010 (http://ec.europa.eu/enterprise/policies/rawmaterials/documents/index_en.htm)



19.3 Contracts The antimony-gold concentrate produced from the Costerfield Operations is sold directly to smelters capable of recovering both the gold and antimony from the concentrates, such that Mandalay receives payment based on the concentration of the antimony and gold within the concentrate. The terms and conditions of commercial sale are not disclosed, pursuant to confidentiality requirements. The marketing of the concentrate is conducted through the agency of Penfold Limited.



20 Environmental Studies, Permitting, and Social or Community Impact

20.1 Environment and Social Aspects

Mine Ventilation A ventilation shaft has been installed in the Cuffley mine to maintain suitable air quality and volumes within the expanded underground mine. The Cuffley ventilation shaft is located on private land owned by Mandalay, and acts as the primary exhaust for the Cuffley area. The facility includes a transformer and connection to an adjacent power supply line.

Water Disposal The disposal of groundwater extracted from the mine workings is a critical aspect of the Costerfield mining operations. The current approved Work Plan does not allow for off-site disposal of groundwater or surface water.

The climate in Central Victoria enables water to be removed through evaporation. Average pan evaporation is 1,400 mm per year according to the nearest Bureau of Meteorology monitoring station at Tatura (65 km northwest of Costerfield). Mean rainfall in the area is 576 mm per year according to the Bureau of Meteorology monitoring station at Heathcote, with the highest annual rainfall recorded in 1973, at 1,048 mm. Rainfall recorded in Heathcote in 2010 and 2011 was higher than average, with 979 mm and 759 mm recorded respectively. Rainfall in 2013, 2014 and 2015 was below average, at 554 mm, 510 mm and 299 mm respectively. Rainfall for 2016 was above average, at 687 mm.

The Costerfield Operation currently operates a series of water storage and evaporation dams, including the following major storage facilities:

• Splitters Creek Evaporation facility: 20 terraces and an HDPE-lined storage dam

• Brunswick Open Pit

• Three HDPE-lined evaporation and storage dams at the Augusta site.

In 2014, Mandalay constructed an RO plant at the Brunswick processing plant, to treat the dewatered groundwater that is piped into the Brunswick pit from the underground workings via a rising main. The RO plant initially uses multi-media filters (MMFs) to remove solid material from the groundwater; flocculent is then added to further separate fine particles, before the water is passed through high pressure membranes to remove the majority of salt and other elements such as antimony, arsenic, lead and zinc.

The treated water is licenced to be discharged into a neighbouring waterway, to be provided to local community members for stock watering or gardening. Otherwise it is used for dust suppression purposes on roads around site. The creek discharge is licenced by the Environmental Protection Authority (EPA), and permits up to 360 ML/year of RO-treated permeate to be discharged into the Mountain Creek South diversion, which feeds into the Wappentake creek at a maximum rate of 2.0 ML/day.

The waste product from the RO plant, known as brine, contains concentrated levels of salt, antimony and other elements removed from the groundwater. The RO plant brine is stored in the plastic-lined evaporation dams at Augusta, reused in the Brunswick processing plant or evaporated on the Bombay TSF.

In 2013, Mandalay purchased 30 ha of land, and a new Mining Licence was granted with the intention of constructing an evaporation facility on the land. Approval was granted in May 2014 for the



evaporation facility known as the Splitters Creek Evaporation Facility. This facility was completed in 2015 and has the capacity to treat 104 ML/year net (evaporation minus rainfall).

The purpose of the facility is to evaporate groundwater extracted from the Costerfield Operations, and thereby maintain dewatering rates from the underground workings. The facility consists of a series of shallow evaporation terraces that follow the natural topographic contours. Groundwater is pumped from the Augusta mine site and discharged at the highest point of the facility; this water cascades down the slope via the evaporation terraces to a deeper Storage Dam at the lowest point. A water pump reticulates water from the Storage Dam to the highest point, to enable the evaporation terraces to be filled from the Storage Dam as evaporation rates allow.

Current evaporation, RO plant processing and re-use capacity are calculated to be approximately equivalent to the current dewatering rates; however, additional complementary treatment options are being investigated to ensure adequate capacity in the future.

Waste Rock Waste rock that is surplus to underground backfilling requirements is stockpiled on the surface in various locations. Testing of the waste rock has confirmed that the material is non-acid generating and therefore does not pose any risk associated with acid mine drainage.

Waste rock is currently stockpiled next to the Augusta Mine box-cut, with the maximum height and shape of the stockpile prescribed in the approved Work Plan. The approved Work Plan requires that this stockpile will be removed on closure, in order to return the land to its prior use as grazing pasture. The waste rock will ultimately be used to fill the box-cut and cap the TSFs. Waste rock has also been transported to both the Bombay and Brunswick TSFs to increase the height of the TSFs, and was used for construction of the Splitters Creek evaporation facility.

A portion of waste rock is screened then utilised in backfilling the underground stopes; however, sufficient waste rock will need to be retained to fulfil rehabilitation and TSF expansion requirements.

Tailings Disposal In May 2015, Mandalay completed construction of a raise of the old Brunswick TSF facility. This facility is currently in use and will provide sufficient tailings storage capacity until the second quarter of 2017. Additional TSF storage capacity is to be provided during 2017 with a lift on the existing Brunswick TSF. Construction of the lift on Brunswick started in January 2017 and will be completed by Q2 2017. This lift will provide an additional 150,000 m3 of volume and the TSF is expected to reach capacity by December 2018.

Planning approval has been granted for a 327,000 m3 total increase in the Bombay facility capacity (comprising a 183,600 m3 lift and a separate 143,400 m3 lift).

The total Brunswick and Bombay approved tailings storage increases allow for an additional three years’ capacity at current milling rates.

The Bombay TSF’s 183,600 m3 lift is the next to be built. This will provide approximately 22 months of capacity.

Air Quality The approved Environmental Monitoring Plan for the Augusta Mine includes an air quality monitoring program based on dust deposition gauges at various locations surrounding the Costerfield Operation, and five dust deposition gauges at the Splitters Creek evaporation facility. Monitoring data is provided to regulatory authorities and community representatives through quarterly Environmental Review Committee (ERC) meetings.



Control measures in place to manage dust emissions from the operations are as follows:

• Road watering program with treated groundwater

• Proactive monitoring of dust with portable Dust Trak monitors

• Moisture control of mill feed during processing

• Wheel washes for vehicles to pass through before leaving site

• Sealing of sections of haul roads

• Maintaining moisture on TSFs and waste rock stockpiles.

The operation of exhaust ventilation shafts is monitored annually, and indicates that they are not a significant source of dust emissions. These results are communicated quarterly at the ERC meetings.

Groundwater Dewatering rates from the mine decreased in 2016, to 468 ML, due to mine development not progressing deeper and a concerted program of capping or grouting exploration holes once drilled. The current groundwater extraction licence of 700 ML/year has been granted by Goulburn-Murray Water.

A conceptual hydrogeological model has been developed for the site based on current groundwater monitoring data, and indicates that the Augusta and Cuffley deposits are located in the regional groundwater aquifer. The model shows a cone of depression in the bedrock aquifer trending in a north to south orientation parallel to the deposits, and indicates that some dewatering has already occurred along the line of the Cuffley Lode, as shown in Figure 20-1.

The regional groundwater aquifer is confined to semi-confined and consists of Silurian siltstones and mudstones. Groundwater flow within this regional aquifer is through fractures and fissures within the rock. This is overlain by a perched alluvial aquifer comprising recent gravels, sands and silt. The perched alluvial aquifer is connected to the surface water system.

Based on the monitoring data and the conceptual hydrogeological model, the dewatering activities at Augusta/Cuffley do not affect the alluvial aquifer. There are no beneficial users of groundwater in the area due to the poor water quality.



Figure 20-1: Groundwater elevation contour map of the areas surrounding the Augusta Mine as at December 2016

Noise The approved Environmental Monitoring Plan for the Costerfield Operation includes a noise monitoring program which comprises routine attended and unattended noise monitoring at six locations, and reactive monitoring at sensitive receptors in the event of complaints or enquiries. Monitoring is carried out in accordance with EPA Victoria’s SEPP N1 policy.

Noise from the operation is a sensitive issue for near neighbours, and the company operates a 24-hour, 7-day complaints line to deal with noise complaints or any other issues from members of the public. The Company’s complaints procedure includes processes to record complaints, and identify and implement immediate and longer-term actions. All complaints are discussed at the quarterly ERC meetings.

The current Costerfield Operation is not expected to significantly change the nature of noise emissions from the site. Construction of new facilities may require some additional noise monitoring, which will be identified as part of the WPV approval process. During construction, an additional 10 dBA of noise is permitted to be generated. Existing resources and procedures are adequate to accommodate any required modifications to the noise monitoring program.

Blasting and Vibration The Department of State Development and Business Innovation (DEDJTR) prescribes blast vibration limits for the protection of buildings and public amenity.

Native Vegetation The Costerfield Operation has been developed and operated with the aim of avoiding and minimising impacts on native vegetation. Where native vegetation has been impacted, the Company has obligations to secure native vegetation offsets.



Mandalay purchased approved native vegetation offset at Peels Lane in Costerfield to fulfil obligations relating to Victoria’s “Native Vegetation Management – A Framework for Action” associated with the original clearing of native vegetation at the Augusta Mine site and the Bombay TSF.

The Peels Lane offset site has been assessed as containing 4.35 habitat hectares of various Ecological Vegetation Classes (EVCs) and associated large trees, in accordance with the framework guidelines.

Expansion of the Costerfield Operation, through construction of the Splitters Creek Evaporation Facility and the Brunswick and Bombay TSFs, has had a minimal impact on native vegetation, and the Peel Lane site has sufficient offset credits to meet the site’s foreseeable future needs.

Visual Amenity The key aspect of the Costerfield Operation that may affect visual amenity was the recent construction of the Splitters Creek evaporation facilities.

Community consultation took place as part of the planning for the facilities, and mitigation measures were implemented where appropriate. Screening vegetation was planted, in consultation with the relevant land manager and near neighbours.

Heritage A heritage survey of the South Costerfield Shaft, Alison and New Alison surface workings was completed by LRGM Consultants in the first quarter of 2012. The purpose of this survey was to identify and record cultural heritage features in the areas of interest that exist within the current Mining Licence (MIN4644). The Taungurung Clans Aboriginal Corporation is the Registered Aboriginal Party designated as the traditional owners of the land on which Mining Licence MIN4644 is located.

The survey identified that no features of higher than local cultural heritage significance were identified, with the following features of local cultural heritage significance being noted:

• South Costerfield (Tait’s) Mine Shaft

• Old Alison Mine Shaft

• New Alison Mine Shaft.

The expansion of the mining operations did not result in any disturbance of historic mine workings or other heritage features.

Community The Costerfield Operation is one of the largest employers in the region and is a significant contributor to the local economy. Mandalay preferentially employs appropriately skilled personnel from the local community and sources goods and services from local suppliers wherever possible.

Mandalay has developed and implemented the Costerfield Operations Community Engagement Plan, which has been approved by the DEDJTR in accordance with the requirements of the MRSD Act. This Plan sets the framework for communication with all of the business’ stakeholders to ensure transparent and ongoing consultative relationships are developed and maintained.

The Community Engagement Plan includes processes to manage community inquiries and complaints to ensure timely and effective responses to issues affecting members of the community.

The current Community Engagement Plan is considered an appropriate framework to address the needs of stakeholders through the planning and implementation of the proposed mine expansion.



In early 2016, Mandalay initiated regular community reference meetings under the auspices of the ERC. This forum, the Community Reference Sub-Committee, gives community members the opportunity to find out about current and future issues at the mine, to provide their input and ask questions.

Mine Closure and Revegetation The MRSD Act requires proponents to identify rehabilitation requirements as part of the Work Plan approvals process, and ensures that rehabilitation bonds are lodged in the form of a bank guarantee to cover the full cost of rehabilitation upfront, prior to commencing work. Rehabilitation bonds are also reviewed on a regular basis to ensure that unit cost assumptions and the scope of work is kept up to date. WPVs also trigger a review of the rehabilitation bond if the work to be carried out affects final rehabilitation.

Mandalay has developed a Mine Closure Plan, which provides an overview of the various aspects of closure and rehabilitation that have been included in the rehabilitation bond calculation and reflects the rehabilitation requirements described in the approved Work Plans and Variations.

The Mine Closure Plan describes how the Augusta site, including the box-cut, waste rock storage, office area and evaporation dams, will be rehabilitated back to its former land use as grazing pasture. The mine decline will be blocked and the portal backfilled with waste rock, with the box-cut being levelled back to its original surface contours. Topsoil and subsoil have been stored on site to facilitate the final vegetation.

The rehabilitation plan for the Brunswick site includes removing all plant and rehabilitating the disturbed area to native forest, to create a safe and stable landform that can be used for passive recreation. The TSFs will be dried out, capped with waste rock and topsoil and planted with native vegetation. The plan includes provisions for monitoring the TSFs post closure.

The rehabilitation plan for the Splitters Creek evaporation facility includes evaporating the remaining stored groundwater and removing the clay lining from the terraces, then placing it back into the HDPE-lined storage dam. The liner in the storage dam is then folded back over the clay and capped with waste rock, clay and topsoil, and planted with grasses. Topsoil and subsoil have been stored on site to enable this final vegetation.

20.2 Regulatory Approvals

Work Plan Variation Future changes to mining activities, such as potential changes to waste rock storage facilities, will require a risk-based WPV to be approved. The DEDJTR facilitates this approval process and will engage with relevant referral authorities as required. The DEDJTR may prescribe certain conditions on the approval, which may include amendments to the environmental monitoring program. The Work Plan approval process involves a thorough consultation process with regulatory authorities, and any conditions or proposed amendments requested to the WPV are generally negotiated to the satisfaction of both parties. All onsite and offsite risks must be assessed in the new Work Plan review process and adequate controls and monitoring programs implemented to mitigate any negative impacts.

Other Permitting In addition to the approval of a WPV, any future expansion of the current Costerfield Operation will require a number of other potential consents, approvals and permits, as listed in Table 20-1.



Table 20-1: Permit requirements

Stakeholder Instrument

Private Landholders Consent/ compensation agreement with owner of land on which the mine is located

City of Greater Bendigo

Planning Permit required for new groundwater evaporation facility and modification to existing TSFs

Department of Environment, Land, Water and Planning

Compliance with Native Vegetation Management Framework for removal of native vegetation associated with the power supply, evaporation facility and expansion of TSF footprints

EPA Consent to discharge RO-treated water to a local waterway



21 Capital and Operating Costs The capital and operating cost estimates for the Project described in the following section have been derived from a variety of sources, including:

• Historic production from the Costerfield Operation

• Lower development costs, due to a change in the mining method to take place in 2017; costs have been estimated from a mining method trial in 2016, allowing for reduced labour costs

• Manufacturers and suppliers

• First principles calculations (based on historic production values)

• Costs including allowances for power, consumables and maintenance.

All cost estimates are provided in 2017 Australian dollars (AUD) and are to a level of accuracy of ±10%. Escalation, taxes, import duties and customs fees have been excluded from the cost estimates.

For reporting purposes, summary tables provide estimates in AUD and US dollars (USD). Costs in USD have been estimated using the AUD:USD exchange rate of 0.75.

21.1 Capital Costs Table 21-1 summarises the estimated total capital requirements for the Costerfield Operation. A detailed breakdown of the individual capital items included in the Economic Model was sourced from the 2016 budget document Version1.4 and updated in January 2016.

Table 21-1: Costerfield Operation – Capital Cost Estimate

Item Total (AUD M)

CY 17 (AUD M)

CY 18 (AUD M)

CY 19 (AUD M)

CY20 (AUD M)

CY21 (AUD M)

Plant 7.3 1.8 1.2 2.5 1 0.7

Admin 0.3 0.1 0.1 0.0 0.1 0.0

Environmental 6.0 2.7 1.0 1.0 0.8 0.4

OH & S 0.2 0.1 0.0 0.0 0.1 0.0

Exploration 0.1 0.1 0.0 0.0 0.0 0.0

Mining 7.4 3.9 1.1 2.4 0.0 0.0

Total - Plant and Equipment 21.2 8.7 3.5 5.9 1.9 1.2

Capital Development 2.3 2.3 0.0 0.0 0.0 0.0

Total Capital Cost 23.5 11 3.5 5.9 1.9 1.2

Note: Total may not add up due to rounding.

Processing Plant Mandalay has identified and estimated the capital costs associated with the maintenance of the Brunswick Processing Plant and other mill site related initiatives, including:

• Bombay and Brunswick TSF Embankment Raise

• Refurbishment of existing plant and key components

• Purchase of critical spares

• Miscellaneous upgrades to surface facilities.

The main Processing Plant infrastructure cost item is the planned embankment raise on the Bombay TSF. All associated costs are based on tendered unit rates for the specific design of this major infrastructure project.



Administration Administration-related capital costs include the replacement of mobile surface plant associated with material handling on surface.

Environmental Environmental-related capital costs include sustaining capital related to ongoing operation of the RO plant, as well as further investigations regarding water management and mine closure.

Mining Mining-related capital costs consist of mainly sustaining capital to ensure the current production rate continues to be maintained, and project capital that further improves the efficiency of the mining process. In particular, the sustaining capital includes mobile plant purchases, including an underground tele-handler and two new development drills. Sustaining capital, in the form of an expansion of the underground electrical distribution system and improved Mines Rescue capability, has been included.

Project capital includes the purchase of a surface tele-remote system and an upgrade of the radio system. The cost estimates are based on recent quotations or agreements with appropriate suppliers.

Capital Development Decline development quantities have been based on the mine designs prepared for the Project. The lateral development quantities are based on each production level in the mine being accessed by the decline system with allowance for stockpiles, level access, sumps, truck tips and CRF mixing bays.

The unit cost for lateral development is based on the development rates agreed with a Mining Contractor undertaking the capital development and Cuffley historical costs for consumables, services and explosives. The Contractor development rates include an allowance for the haulage of waste rock to surface.

Closure Closure costs are estimated using a calculation tool to estimate rehabilitation bonds. Bond amounts are reviewed when major changes are made to the operation for example construction of a TSF. Closure costs are expected to be refunded by the current rehabilitation bonds held by the regulatory authorities; hence no additional closure costs have been included.

21.2 Operating Costs The operating cost estimates applied in this Technical Report summarised in Table 21-2 are based on the production physicals outlined in Section 16 and the unit costs summarised in Table 21-3. The basis of estimate is described further in the following sections.



Table 21-2: Costerfield Operation – Operating Cost Estimate

Description Operating Cost (per t milled)

AUD M AUD/t USD M USD/t

Mining 108 171 81 129

Processing 30 48 23 36

Site Services, General and Administration 47 74 35 56

Total 184 294 139 221

Table 21-3: Operating Cost Inputs


Mining

Jumbo Lateral Development AUD/m 1,371

Stoping AUD/t 90

Mining Admin AUD/day 16,959

Geology AUD/day 5,679

ROM Haulage AUD/t 8

Processing Plant AUD/t milled 51

Site Services AUD/day 13,229

General and Administration AUD/day 14,741

Selling Expenses including Royalty AUD/t con 164

Lateral Development The estimated unit cost for lateral development has been developed from Augusta and Cuffley historical costs for labour, equipment, consumables, services and achieved productivities. An allowance for haulage to surface has also been included.

The lateral airleg development (operating) for Augusta and Cuffley will continue to be undertaken as an ‘owner operator’.

The required lateral development is summarised in Table 21-3.

Table 21-4: Summary of Development requirements


Jumbo Development m 253

Airleg Development (Waste) m 7,829

Airleg Development (Ore) m 5,778

The direct operating costs related to lateral development include:

• Direct labour (includes superannuation, workers compensation, payroll tax and partial allowances for leave accrual)

• Drilling consumables (drill steel, bits, hammers, etc.)

• Explosives

• Ground support supplies

• Direct mobile plant operating costs (fuel and lubricants, tyres and spare parts)

• Services materials including poly pipe, ventilation bags and electrical cables



• Reallocation of costs associated with maintenance, ventilation, power supply, compressed air supply, dewatering, water supply and underground communications

• Miscellaneous materials required to support development activities.

Production Stoping The direct costs for production stoping have been developed from Augusta and Cuffley historical costs for direct labour, consumable materials, equipment operating and maintenance as well as achieved productivities associated with the following:

• Installation of secondary ground support

• Drilling, loading, and blasting longholes by Mandalay employees

• Production from the stope with an underground loader and tramming to a stockpile or truck loading area

• Loading haul trucks from stockpile (if required)

• Backfill preparation and CRF placement

• Reallocation of costs associated with maintenance, ventilation, power supply, compressed air supply, dewatering, water supply and underground communications.

Mining Administration Mining administration includes costs associated with mining management, supervision and technical services (mining engineering, survey and geotechnical engineering). These costs have been estimated from Mandalay’s actual 2016 mining administration costs.

Geology Geology includes costs associated with resource estimation, resource definition drilling, production geology expenses, sampling, assaying, laboratory expenses and associated management and labour. These costs have been estimated from Mandalay’s actual 2016 geology costs.

ROM Haulage The cost of trucking from the Augusta box-cut to the Brunswick ROM pad has been calculated based on historical Augusta cost information and includes private contractor labour, haul truck operating and maintenance costs, including indirect costs and profit. The average cost of the trucking has been calculated at AUD8.10/t delivered to the Brunswick ROM pad.

21.3 Processing Plant Processing Plant costs include tailings disposal, ROM management, ball mill crushing and grinding, general operating and maintenance, reagent mixing, thickening, flotation, gold room expenses, all flocculants and reagent chemicals, plant maintenance and reallocated electrical costs associated with Plant operation. Brunswick processing costs have been estimated from Mandalay’s actual 2016 processing costs.

21.4 Site Services Site services costs refer to indirect costs related to Health and Safety, Environment and Community Relations, as well as costs related to the water treatment plant, water disposal and RO plant. Compensation expenses are also included in this cost item. These costs have been estimated from Mandalay’s actual 2016 site services costs.



21.5 General and Administration General and administration costs refer to site wide operational costs rather than costs directly associated with operational departments. These costs include General Site Management (including all staff costs), Human Resources, Finance and Administration. These costs have been sourced from Mandalay’s actual 2016 general and administration costs.

21.6 Selling Expenses Mandalay utilises a third party company to arrange the sale and transport of concentrate from the Brunswick Processing Plant to the smelter in China. The Mandalay portion of the selling expenses is calculated from historical costs and comprises road transport from the Brunswick Processing Plant to the Port of Melbourne, ship transportation from Melbourne to China, shipment documentation, freight administration and assay exchange/ returns.



22 Economic Analysis This section is not required as the property is currently in production. Mandalay is a producing issuer and there is no planned material expansion of the current production. SRK has verified the economic viability of the Mineral Reserves via cash flow modelling, using the inputs discussed in this report.



23 Adjacent Properties 23.1 General Statement about Adjacent Properties

The Costerfield Operation Mining Lease (MIN4644) is completely enveloped by exploration leases held by Mandalay. In the immediate area of the Augusta Mine there are no advanced projects, and there are no other Augusta-style antimony-gold operations in production in the Costerfield district.

Exploration on adjacent prospects (EL5406, EL5490, EL3316), as shown in Figure 23-1, is at an early stage and is not relevant for further discussion in this Technical Report. Three other tenements are under application and awaiting approval (EL5546, EL06280 and EL5548).

The ownership and status of each of the surrounding exploration leases is shown in Table 23-1.

Figure 23-1: Augusta Mine adjacent properties Source: DEDJTR Geovic, 2016.

Table 23-1: Ownership of Augusta Mine adjacent properties

Title Owner Status First granted Expiry

EL5406 Victoria Mining Exploration Pty Ltd Granted 11/01/2012 31/10/2017

EL5490 Golden Camel Mining Pty Ltd Granted 23/08/2013 05/12/2018

EL3316 Nagambie Mining Granted 19/12/1992 19/12/2014

EL5546 Nagambie Mining Under application (submitted 29/01/2015)

EL006280 Mercator Gold Australia Pty Ltd Under application (submitted 19/09/2016)

EL5548 Golden Camel Mining Pty Ltd Under application (submitted 16/02/2015)

Source: DEDJTR Geovic, 2016.



The Costerfield Operation is situated 35 - 51 km from other significant central Victorian mining operations; Table 23-2 gives distances from the Augusta Mine site to mines in Central Victoria.

Table 23-2: Distance from the Augusta Mine site to significant mining projects

Mine Owner Distance (km) General direction

Nagambie Mine Nagambie Mining Ltd 40 East-northeast

Fosterville Mine Kirkland Lake Gold 35 Northwest

Kangaroo Flat Mine GBM Gold Ltd 51 West-northwest



24 Other Relevant Data and Information Additional information that is deemed relevant to ensure this Technical Report is understandable and not misleading is discussed in the sections below.

24.1 Remnant Mining Remnant mineralisation exists above the 1049 mRL, primarily in E Lode and to a lesser extent W Lode. This remnant mineralisation has remained in situ as pillars to support the local ground stability when difficult mining conditions were encountered or the mining shapes did not prove to be economically viable when mining was conducted.

There is an opportunity to recover this material and realise its value; however, at this stage of assessment sound methodologies to identify the hazards associated with remnant mining to ensure higher levels of safety and improved extraction efficiency have not yet been determined. For this reason, only areas of high confidence, to be extracted with our current mining methods have been included in this report.

However, Mandalay expects to progressively target remnant areas as a source of higher grade mineralisation and to support increased mining production rates as sound methodologies for planning are devised and implemented.



25 Interpretation and Conclusions It is the QP’s opinion that the positive results of the financial analysis demonstrates Mandalay’s ability to maintain and manage the economics of the project. The work has bene undertaken at an appropriate level to support the release of the Mineral Resource and Ore Reserve estimates.

Reconciliation results show good precision and reasonable accuracy between the resource block model data and the processing plant data. Unquantified errors such as stockpiling, ore-waste misallocation, and unplanned dilution influenced the reconciliation data. Over the period, the ounces of gold predicted by the model were 9% higher than produced by the plant. The tonnes of Antimony predicted by the model were 1% lower than produced by the plant. Minor adjustments to the Resource estimation process for gold grades have been made to attempt to improve the reconciliation for 2017.

The reconciliation results give confidence to the sample collection procedures, the quality of the assays and the resource estimation methodology.

Overall the measured and indicated resource tonnage has remained similar to 2015 with a contained metal decrease of 40,000 oz gold (16%) and 6.5 tonnes (16%) antimony. These decreases equate to an overall drop in gold grade from 8.6 g/t to 6.9 g/t and antimony grade from 3.7% to 2.9%.

SRK makes the following observations:

• Inferred Resources have not been included in the economic evaluation.

• There has been a history of conversion of Inferred to Indicated Resources resulting in additional Resources from outside the Mineral Reserve being included into the life of mine plans that have the potential to improve the project economics.

• Mandalay has demonstrated an ability to improve the mining method and productivity based on improved geological information and thus mine designs and planning.



26 Recommendations The Costerfield Property is an advanced property and Mandalay has a history of successful exploration and mining on the Property. SRK has observed that the degree of technical competency evident in the work performed by Mandalay geologists is high, particularly in the structural analysis of the local geology. Therefore, there is no requirement for additional work programs over and above the existing operational plans.

Based on the findings of this Technical Report, SRK notes that additional Resources below the current cut-off grade and in the Inferred Resource classification have the potential to be included in the life of mine plans. SRK supports ongoing technical work and assessment of the Brunswick deposit.

SRK recommends that Mandalay continually review the cut-off grade based on changes to the cost profile and commodity prices.

Application of a variable cut-off grade ie for each deposit (Cuffley and Augusta) should continue to be explored and applied by site operational personnel.



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VandenBerg AHM, Willman CE, Maher S, Simons BA, Cayley RA, Taylor DH, Morand VJ, Moore DH & Radojkovic A, 2000. The Tasman Fold Belt System in Victoria. Geological Survey of Victoria Special Publication.

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