Costerfield Operation, Victoria, Australia NI 43-101 · PDF fileCosterfield Operation,...

Costerfield Operation, Victoria, Australia NI 43-101 Report

Prepared for:

Mandalay Resources Corporation 76 Richmond Street, Toronto ONTARIO MSC1P1 Canada

Prepared by:

SRK Project Number: PLI016

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

Simon Walsh, BSc (Extractive Metallurgy), MBA Hons, CP, GAICD, Associate Princpal Consultant

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

Date of Report: 30 March 2015 Effective Date: 31 December 2014

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Date and Signature Page

SRK Project Number: PLI016

SRK Consulting (Australasia) Pty Ltd Level 8, 365 Queen Street Melbourne Victoria 3000 Australia

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

Project Manager: Peter Fairfield

Date of Report: 30 March 2015

Effective Date: 31 December 2014

Peter Fairfield, BEng (Mining), FAusIMM, Principal Consultant (Project Evaluation)

Qualified Person (QP)

Authors:

Peter Fairfield, Danny Kentwell, Simon Walsh.

FAIR\EBBE\mota PLI016_Costerfield_Operations_NI_43 101_Master_Rev0.docx 30 March 2015

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Important Notice This report was prepared as a National Instrument 43-101 Technical Report for Mandalay Resources Corporation (Mandalay) by SRK Consulting (Australasia) Pty Ltd (SRK). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in SRKs services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, including without limitation, those sources listed in Section 3 - Reliance on Other Experts, and iii) the assumptions, conditions, and qualifications set forth in this report.

This report is prepared for the investing public and is intended for use by Mandalay subject to the terms and conditions of its contract with SRK and relevant securities legislation. The contract permits Mandalay to file this report as a Technical Report with Canadian securities regulatory authorities pursuant to National Instrument 43-101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk. SRK accepts no responsibility with respect to the opinions of those experts listed in Section 3 - Reliance on Other Experts nor determinations made by Mandalay with respect its obligation to file this Technical Report, or subsequent technical reports, nor any determinations as to the materiality of a mineral project to Mandalay, and SRK is under no obligation to update this Technical Report, except as may be agreed to between Mandalay and SRK by contract from time to time. The user of this document should ensure that this is the most recent Technical Report for the property as it is not valid if a new Technical Report has been issued.

Copyright

This report is protected by copyright vested in SRK. It may not be reproduced or transmitted in any form or by any means whatsoever to any person without the written permission of the copyright holder, SRK, except for the purpose as set out in this Technical Report Section 2.3.


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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 April Westcott

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

8 Deposit Types Danny Kentwell April Westcott

9 Exploration Danny Kentwell April Westcott

10 Drilling Danny Kentwell April Westcott

11 Sampling Preparation, Analyses and Security Danny Kentwell Chris Davis

12 Data Verification Danny Kentwell Chris Davis

13 Mineral Processing and Metallurgical Testing Simon Walsh Damon Buchanan

14 Mineral Resource Estimates Danny Kentwell Chris Davis

15 Mineral Reserve Estimates Peter Fairfield Shannon Green

16 Mining Methods Peter Fairfield Shannon Green

17 Recovery Methods Simon Walsh Damon Buchanan

18 Project Infrastructure Peter Fairfield Shannon Green

19 Market Studies and Contracts Peter Fairfield Andre Booyzen

20 Environmental Studies, Permitting and Social, or Community Impact Peter Fairfield April Westcott

21 Capital and Operating Costs Peter Fairfield Shannon Green

22 Economic Analysis Peter Fairfield Shannon Green

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 DSDBI Department of State Development, Business and Innovation DEPI Department of Environment and Primary Industries 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/LoM life of mine L/s litres per second LRGM LRGM Consultants M million Ma million years Mandalay Mandalay Resources Corporation 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 Mtpa million tonnes per annum


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Abbreviation Meaning

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 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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Executive Summary 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 28 March 2014.

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 has been the sole ore source for the Brunswick Processing Plant until December 2013. At this time ore production started from the Cuffley Deposit located approximately 500 m to the north of the Augusta mine workings. The Cuffley deposit will be mined in conjunction with remanent ore from the Augusta deposit.

Property Description and Location The Costerfield Operation encompasses the underground Augusta Mine, the Brunswick Processing Plant, the Cuffley deposit and associated infrastructure, all of which is located within mining licence MIN 4644. The mining licence is located with exploration lease EL 3310 which is held 100% 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.

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, as well as in 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 site is located on unrestricted Crown land. The surrounding land is largely rocky, rugged hill country administered by the Department of Environment and Primary Industries (DEPI) as State Forest. The Puckapunyal Military Area is located on the eastern boundary of the Project Area.


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The area has a Mediterranean climate with temperature ranges 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 to 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.

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 and 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.

It was not until 2001 that 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. Between November 2006 and May 2012, 249,751 t at 7.7 g/t gold and 4.2% antimony was processed from the Augusta Mine yielding 10,459 t of antimony and 61,833 oz of gold.

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 this Inferred Resource to Measured and Indicated in 2013.

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 nine lodes: E Lode, W Lode, N Main Lode, NE Lode, NSW Lode, P1 Lode, P2 Lode, K Lode, and B Lode. The Cuffley Deposit is comprised of four lode structures, Cuffley Main (CM Lode), Cuffley East (CE Lode) Cuffley Deeps (CD Lode) and Alison South (AS Lode). All the lodes are located in the west-dipping Costerfield Formation (as defined by Talent, 1965) is a series of thickly bedded mudstones and siltstones featuring heavy bioturbation. Augusta and Cuffley deposits are bounded between two large, low angle west-dipping parallel thrust


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faults named the Adder Fault (upper) and the King Cobra Fault (lower) that are typically in the range of 250 m apart.

Mining has demonstrated that mineralised splay veins and that 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. The Costerfield Operation is currently extracting ore from the W Lode, E Lode, N Lode CM Lode and CE Lode.

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 occurs 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 at Costerfield consists of fault-hosted veins that are mostly less than 1.5 m in width and that have been formed in multiple phases. The earliest phase consists of bedding parallel laminated quartz veins, which are barren. The laminated quartz phase is followed by a quartz-pyrite-arsenopyrite phase that contains erratic coarse gold. The last phase consists of massive stibnite, which contains evenly distributed fine-grained gold. The Costerfield ‘lodes’ or ‘reefs’ are typical anastomosing, en echelon-style narrow vein systems dipping from 25° to 70° west to steep east (70° to 90° east). Mineralised shoots plunge steeply north at the southern end of the field.

Mineral Resource Estimates The Mineral Resources are stated here for the Augusta, Cuffley and Brunswick deposits with an effective date of 31 December 2014. The Mineral Resource is depleted for mining up to 31 December 2014. The figures also exclude remnant mineralisation that remains in pillars.

The Augusta, Cuffley and Brunswick deposits consist of a combined Measured and Indicated Mineral Resource of 999,000 tonnes at 7.5 g/t gold and 3.6% antimony, and an Inferred Mineral Resource of 519,000 tonnes at 5.3 g/t gold and 2.6% antimony.

This Mineral Resource includes a ROM stockpile comprised of 17,000 tonnes at 6.0 g/t gold and 2.6% antimony.

The Mineral Resources are reported at a cut-off grade of 3.8 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 USD12,000 per tonne as follows:

• AuEq =Au (g/t) + 2.03 x Sb (%)

The cut-off grade has decreased from, 3.9 g/t AuEq used in the previous resource report effective 31 December 2013, and the relative value attributed to antimony has increased slightly from 1.99 over the year. The changes in the cut-off grade and the gold equivalent (AuEq) formula are based on updated costs, recoveries and assumptions.


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All relevant diamond drillhole and underground face samples in the Costerfield Property, available as of 31 November 2014 for the Augusta and Cuffley deposits were used to inform the estimate. The estimation methodology followed the approach used by SRK in the December 2013 Resource Estimate reported in an NI 43-101 Technical Report on the Property released on 28 March 2014 (SRK, 2014).

Mineral Resources for Brunswick have not been re-estimated since they were last reported in 2009 (Fredericksen, 2009). No drilling or mining has occurred on these structures since then. The Brunswick Mineral Resources are restated here for consistency with the other Costerfield lodes with an updated cut-off grade of 3.8 g/t AuEq and a 1.2 m minimum mining width criterion.

The effective date for the Brunswick Mineral Resources is 31 December 2014. The reporting of the Brunswick model in previous years has not been in line with the reporting methodology of Augusta and Cuffley Resources. This has been rectified in the 2015 estimation and has resulted in a drop in tonnes and grade of the Brunswick deposit.

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


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Table ES-2-1: Summary of the Augusta, Cuffley and Brunswick Mineral Resource, inclusive of Mineral Reserve

Lode Name Resource Category

Inventory (Tonnes)

Gold Grade (g/t)

Antimony Grade

(%)

Contained Gold (Oz)

Contained Antimony

(t)

Aug

usta

Dep

osit

E Lode Measured 55,000 5.4 2.9 10,000 1,600

Indicated 53,000 4.1 2.4 7,000 1,300

Inferred 86,000 3.1 2.7 9,000 2,400

B Lode Indicated 33,000 4.4 2.6 5,000 800

W Lode Measured 20,000 6.0 2.7 4,000 500

Indicated 25,000 5.3 2.8 4,000 700

Inferred 29,000 5.1 2.3 5,000 700

NM Lode Measured 54,000 11.5 6.0 20,000 3,300

Indicated 242,000 7.3 3.8 57,000 9,100

Inferred 121,000 4.6 2.3 18,000 2,800

NE Lode Measured 1,000 3.6 2.1 100 30

Indicated 52,000 3.4 1.6 6,000 800

Inferred 68,000 3.9 1.1 8,000 700

NSW Lode Indicated 9,000 3.7 2.4 1,000 200

P1 Lode Measured 14,000 9.8 2.5 4,000 300

Indicated 10,000 10.0 2.7 3,000 300

P2 Lode Indicated 12,000 6.9 3.2 3,000 400

Inferred 6,000 3.1 1.6 1,000 100

K Lode Indicated 29,000 2.9 2.0 3,000 600

Cuf

fley

Dep

osit

CM Lode Measured 40,000 17.7 6.9 23,000 2,800

Indicated 274,000 8.7 3.8 76,000 10,300

CE Lode Measured 11,000 16.9 5.2 6,000 600

Indicated 20,000 10.2 3.6 7,000 700

Inferred 6,000 13.7 6.5 2,000 400

CD Lode Inferred 56,000 8.9 3.4 16,000 1,900

AS Lode Indicated 28,000 4.8 3.7 4,000 1,000

Inferred 9,000 0.8 2.3 200 200

Brunswick Deposit Inferred 139,000 6.7 3.3 30,000 4,600

Stockpiles Measured 17,000 6.0 2.6 3,000 400

Total Measured + Indicated 999,000 7.5 3.6 242,000 35,900

Total Inferred 519,000 5.3 2.6 89,000 13,700

Notes:

1). Mineral Resources estimated as of 31 December 31 2014, and depleted for production through 31 December 2014.

2). Mineral Resources stated according to CIM guidelines.

3). Mineral Resources are inclusive of Mineral Reserves.

4). Tonnes and contained gold (oz) are rounded to the nearest thousand; contained antimony (t) rounded to nearest hundred.

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


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6). A 3.8 g/t Au Equivalent (AuEq) cut-off grade over a minimum mining width of 1.2 m is applied where AuEq is calculated at a gold price of $1,400/oz, antimony price of $12,000/t and exchange rate AUD:USD of 0.85.

7). The gold Equivalent value (AuEq) is calculated using the formula: AuEq = Au g/t + 2.03 * Sb%.

8). The Brunswick Mineral Resource has not been re-estimated since it was reported in Frederickson, 2009, Costerfield Gold and Antimony Project, Augusta and Brunswick Deposits. Frederickson Geological Solutions Pty Ltd.

9). The Brunswick resource reporting methodology has been reviewed and is now consistent with that of the Augusta and Cuffley Deposits.

10). The Mineral Resource estimation for Augusta and Cuffley deposits was performed by Chris Davis BSc, MAusIMM, who is a full-time employee of Mandalay Resources and was independently verified by Danny Kentwell MSc, BAppSc, FAusIMM, a full-time employee of SRK Consulting who is a qualified person under NI 43-101 and is the Competent Person for the Augusta and Cuffley Mineral Resource Estimates.

11). Danny Kentwell MSc, BAppSc, FAusIMM, full-time employee of SRK Consulting is a qualified person under NI 43-101 and is the Competent Person for the Brunswick Mineral Resource Estimate.

Mining Methods Mandalay plans to mine the extension of the Augusta Mineral Resources to the 829 mRL and the Cuffley Lode to the 814 mRL. The proposed stope outlines are presented in Figure ES- 1.

Figure ES- 1: Schematic of Augusta and Proposed Cuffley Underground Design

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 is serviced by a decline haulage system. 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.

The Augusta Mine employs predominantly air leg long-hole stoping methods as well as longitudinal uphole retreat working a bottom up sequence. These mining methods have been utilised throughout 2010 to 2014. 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

North


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single decline that connects to the existing Augusta decline at the 1,030 mRL. The Augusta Mine is scheduled to produce mill feed from three different mining methods: full face development, long-hole CRF stoping and half upper 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.

Mineral Reserve Estimate From the Mineral Resource, a mine plan was designed based only on Measured and Indicated Resource blocks using predominantly the cemented rock fill blast hole stoping method. A cut-off grade of 5.0 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 USD10,000/t Sb prices. The Mineral Reserve estimate is presented in Table ES- 2.

Table ES-2: Mineral Reserves at Costerfield, as of 31 December 2014

Category Inventory (Tonnes)

Gold Grade (g/t)

Antimony Grade

(%)

Contained Gold (oz)

Contained Antimony

(t)

Proven 98,000 10.4 4.5 32,000 4,400

Probable 336,000 7.4 3.3 80,000 11,200

Proven + Probable 434,000 8.1 3.6 112,000 15,600

Notes:

1). CIM definitions followed for classification of Proven and Probable Reserves.

2). Mineral Reserve estimated as of 31 December 2014, and depleted for production through to 31 December 2014.

3). Tonnes and ounces are rounded to the nearest thousand; contained antimony rounded to nearest hundred.


5). Lodes have been diluted to a minimum mining width of 1.2 m for stoping and 1.8 m for ore development.

6). A 5.0 g/t Au Equivalent (AuEq) cut-off grade has been applied.

7). Commodity prices applied are gold price of USD1,200/oz, antimony price of USD10,000/t and exchange rate AUD:USD of 0.85.

8). The Au Equivalent value (AuEq) is calculated using the formula: AuEq = Au g/t + 1.97 * Sb%.

9). 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. The Mineral Reserve estimate was prepared by Shannon Green P.Eng., BEng, MAusIMM, a full-time employee of Mandalay Resources and was independently verified by Peter Fairfield, FAusIMM, a full-time employee of SRK Consulting and a qualified person under NI 43-101.

The Mineral Reserve does not include any portion of the 519,000 tonnes of Inferred Resources.

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 north of the Brunswick Processing Plant – the Bombay tailings storage facility.


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Ore from the Augusta underground mine, including 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 ore has been demonstrated over this period of operation and has delivered stable and consistent recoveries. The Brunswick Processing Plant performance on the currently mined and processed Augusta orebody from February 2012 to December 2013 has increased, with the mobile crushing plant increasing throughput. Throughput for the 2014 calendar year was 12,445 t/mth, with a maximum of 13,618 tonnes on a blend of 44% Cuffley and 56% Augusta ore which yielded metallurgical recoveries of 94.5% Sb and 89.6% Au.

Figure ES- 2 shows the mill performance before and after the transition to the Cuffley ore.

Figure ES- 2: Mill performance on Augusta ore and Cuffley/Augusta mix

Mineralogical and metallurgical test work has been completed on three different representative samples of the Cuffley mineralisation. These tests show that the Cuffley mineralised material will perform through the existing Brunswick Processing Plant similarly to the Augusta Mill Feed. Actual processing of the Cuffley mineralised material has also confirmed the similar performance to the Augusta material.

Grindability of the Cuffley material tested to date has also been the same as for Augusta.

With the inclusion of a mobile crushing unit, the capacity of the plant has been successfully upgraded to over 12,000 tonnes per month (t/mth)

The improvements in the gold room capacity have enabled an increased proportion of the gold to be recovered into a separate gold concentrate. As the gold concentrate has superior payment terms compared to the gold credit received for gold in the antimony concentrate, this has resulted in increased value for the project.


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Project Infrastructure The Costerfield Operations surface facilities are representative of a modern gold-antimony mining operation. The Augusta Mine site, refer to Figure 18-1, comprises the following:

• Office and administration complex, including change facilities;

• 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;

• Ventilation exhaust raises;

• Ventilation intake raise; and

• Augusta Mine dam to manage rainfall run-off and supply the mine with water.

Figure ES- 3: Augusta Mine Site


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Figure ES- 4: Brunswick Site Area

The Brunswick site, refer to Figure 18-2, comprises the following:

• Gold-antimony processing plant and associated facilities;

• Central administration complex;

• Process plant workshop;

• Tailings storage facilities;

• Run of Mine stockpiles;

• Reverse Osmosis Plant;

• Previously mined Brunswick Open Pit; and

• Core farm and core processing facility.


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Underground Infrastructure Underground infrastructure within the Augusta Mine is extensive and includes a 4.5 m wide by 4.8 m high access decline extending from surface to a current depth of 870 mRL. Access for the Cuffley Lode commences at the 1,030 mRL and extends to approximately the 895 mRL. At the 935 mRL, the Cuffley Incline extends off the Cuffley Decline and accesses mineral resources from the 945 mRL to the 1,005 mRL at year end. A second decline within Cuffley known as the 4800 Decline was commenced during 2014 to access the southern part of the Cuffley Main Lode which is positioned south of the East Fault. This Decline commences at the 960 mRL and extended to 900 mRL by year end. The gradient of all the declines and inclines development within Augusta and Cuffley has been developed at a grade of 1:7.

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 2,353 kVA shared between the Augusta Mine, Cuffley Mine and a 1,000 kVA supply at the Brunswick Processing Plant.

Powercor owns and manages the central Victorian electrical distribution infrastructure. Energy Australia (formerly Tru Energy) is the current electrical retailer.

Mandalay had planned to upgrade the Powercor distribution infrastructure to both the Augusta and Cuffley sites, but after significant evaluation has opted to maximise the mains electrical supply and supplement this feed with diesel generators on an as needed basis.

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.

Environmental Studies, Permitting and Social Impacts The environmental and community impacts associated with the proposed expansion of the current Costerfield Operation have been 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 State Development, Business and Innovation (DSDBI) facilitates this approval process and will engage with relevant referral authorities, as required. The DSDBI may prescribe


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certain conditions on the approval, which may include amendments to the environmental monitoring programme. 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.

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

• Historical production from the Costerfield Operation, predominantly the past twelve months completed by Mandalay;

• Manufacturers and suppliers;

• First principle calculations (based on Historical production values); and

• Costs include allowances for power, consumables and maintenance.

All cost estimates are provided in 2015 Australian dollars (AUD) and are to a level of accuracy of ± 10 per cent. Escalation, taxes, import duties and custom fees have been excluded from the cost estimates.

For reporting purposes summary tables provide estimates in AUD and US Dollars (USD). The USD have been estimated using the AUD:USD exchange rate of 0.85. The total capital required for the Costerfield Operation is summarised in Table ES- 3.

Table ES- 3: Costerfield Operation – Capital Cost Estimate

Item Total (AUD M)

CY 15 (AUD M)

CY 16 (AUD M)

CY 17 (AUD M)

CY 18 (AUD M)

Plant 4.4 2.5 0.5 1.4 0.0

Admin 0.2 0.2 0.0 0.0 0.0

Environmental 3.2 3.2 0.0 0.0 0.0

OH & S 0.0 0.0 0.0 0.0 0.0

Exploration 0.0 0.0 0.0 0.0 0.0

Mining 3.0 2.7 0.1 0.2 0.0

Ops Geology 0.0 0.0 0.0 0.0 0.0

Total-Plant and Equipment 10.7 8.6 0.6 1.5 0.0

Capital Development 7.7 7.8 0.0 0.0 0.0

Total Capital Cost 18.5 16.3 0.6 1.6 0.0

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 total operating costs for the Costerfield Operation is summarised in Table ES- 4.


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Table ES-4: Costerfield Operation – Operating Cost Estimate

Description Operating Cost

AUD M AUD/t USD M USD / t

Mining 73 169 62 144

Processing 27 63 23 53

Site Services , General and Administration 18 42 15 26

Total 118 274 100 223

Million dollars rounded to nearest million.

“/ t” rounded to the nearest dollar.

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. 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.

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 and is dependent on metal prices and exchange rates at the time of sale.

Royalties and Compensation The following royalty rates have been applied for the Costerfield Operation:

• Victorian Government Antimony Royalty: 2.75% sales less any selling expenses; and

• Victorian Government Gold Royalty: 0.0% of sales.

Economic Analysis The key project criteria and assumptions used in preparation of the cash flow analysis have been listed in Table ES- 5. A summary of the project economics are presented in Table ES- 6 and the Project sensitivity is presented in Figure ES- 5.

Table ES- 5 Project Criteria

Description Units Quantity

Proposed Mill Feed

Tonnes (kt) 431

Gold grade (g/t) 8.1

Antimony grade (%) 3.6

Project Life months 35

Average Production Rate t/mth 12,300

Maximum Mining Rate t/mth 13,800

Metallurgical Recovery Gold (%) 85 - 90

Antimony (%) 94

Gravity Gold % 42

Concentrate Grade Gold (g/t) 71


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Antimony (%) 53

Concentrate Selling Expenses AUD/dmt 213

Exchange Rate AUD: USD 0.85

Commodity Prices Gold USD/oz 1,200

Antimony USD/t 10,000

Table ES- 6: Project Economics

Description Units Quantity Units Quantity

Tonnes Milled Tonnes 430,725 Tonnes 430,725

Recovered Gold Ounces 98,039 Ounces 98,039

Recovered Antimony Tonnes 14,730 Tonnes 14,730

Operating cost AUD M 118.0 USD M 100

Operating Cost per Payable ounce AUD / Oz Eq1 728 USD/oz eq 727.5

Capital cost AUD M 19.4 USD M 11.3

Payable Gold Ounces 84,990 Ounces 84,990

Payable Antimony Tonnes 9,280 Tonnes 9,280

Payable (Saleable) Metal – Au Eq Oz Eq 162,322 Oz Eq 162,322

Net Revenue (less selling expenses) AUD M 220 USD M 187

After Tax Cash Flow AUD M 81 USD M 69

1 Oz Eq – Gold Ounces + (Antimony Price / Gold price) * Antimony Tonnes.

Tonnes and ounces rounded to nearest thousand.

Million dollars rounded to nearest million.

Figure ES- 5: Project Sensitivity


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Adjacent Properties The Costerfield Operation Mining Lease (MIN4644) 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 (EL3316, EL5490 and EL5406) are at an early stage and not relevant to discuss further in relation to this Technical Report.

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

Remnant Mining Remnant mineralisation exists above the 1,049 mRL, primarily in E Lode and to a lesser extent 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, the remnant mineralisation has not been included in the Technical Report and does not exist in the previously released Mineral Reserve statement for Augusta.

Interpretation and Conclusions 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.

Drilling pattern and density are also considered to be appropriate as sectional 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.

The reproducibility between laboratories of gold grades for 2014 is similar to that which was found in earlier check assay programmes from 2013 and 2012. In contrast, the reproducibility of antimony assays across laboratories decreased markedly from 92% RPD to 68% RPD. This reason for this difference is currently under investigation with a larger sample population being prepared for a three way check assay programme in early 2015. Preliminary investigations suggest that the apparent overcall of antimony grades by Onsite may be more reliable as their acid digest method has been optimised for high grade antimony through the addition of tartaric acid to prevent antimony precipitation out of the solution.

Reconciliation results show good precision and reasonable accuracy between the resource block model data and the processing plant data. 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 increased by 202,000 tonnes of ore with a contained metal increase of 4,000 oz gold and 3,500 t antimony. These increases equate to an overall drop in gold grade from 9.3 g/t to 7.5 g/t and antimony grade drop of 4.1% to 3.6%.

The resource cut-off has dropped slightly from 3.9 AuEq (g/t) to 3.8 AuEq (g/t). The gold equivalent factor used has risen slightly from 1.99 to 2.03. These two changes have the net effect of increasing the tonnes whilst reducing grade for an overall increase in contained metal.


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SRK makes the following observations regarding the mining operations:

• 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 form 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 continuing to increase and improve the geological information and thus mine designs and planning.

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 programmes 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 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 i.e. for each deposit (Cuffley and Augusta) should continue to be explored and applied by site operational personnel.


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Table of Contents

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

List of Abbreviations ..................................................................................................................................... v

Executive Summary ................................................................................................................................... viii

2 Introduction .................................................................................................................. 1

2.1 Scope of Work ..................................................................................................................................... 1 2.2 Work Programme ................................................................................................................................ 2

2.3 Basis of Technical Report ................................................................................................................... 2

2.4 Qualifications of SRK and SRK Team ................................................................................................ 2

2.5 Acknowledgement ............................................................................................................................... 3

2.6 Declaration .......................................................................................................................................... 3

3 Reliance on Other Experts ........................................................................................... 5

3.1 Marketing ............................................................................................................................................ 5

4 Property Description and Location ............................................................................. 6

4.1 Property Location ................................................................................................................................ 6

4.2 Land Tenure ........................................................................................................................................ 7

4.3 Underlying Agreements .................................................................................................................... 10

4.4 Environmental Liability ...................................................................................................................... 10

4.5 Royalties ........................................................................................................................................... 10 4.6 Taxes ................................................................................................................................................ 10

4.7 Legislation and Permitting ................................................................................................................. 11

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ......... 12

5.1 Accessibility ....................................................................................................................................... 12

5.2 Land Use ........................................................................................................................................... 12

5.3 Topography ....................................................................................................................................... 12

5.4 Climate .............................................................................................................................................. 12 5.5 Infrastructure and Local Resources .................................................................................................. 13

6 History ......................................................................................................................... 15

6.1 Introduction ....................................................................................................................................... 15

6.2 Ownership and Exploration Work ..................................................................................................... 15

6.2.1 Mid-East Minerals (1968–1971) ............................................................................................ 16

6.2.2 Metals Investment Holdings (1971) ....................................................................................... 16

6.2.3 Victorian Mines Department (1975–1981) ............................................................................ 16 6.2.4 Federation Resources NL (1983–2000) ................................................................................ 16

6.2.5 Australian Gold Development NL/Planet resource JV (AGD) (1987–1988) ......................... 16

6.2.6 Australian Gold Development NL (AGD) (1987–1997) ......................................................... 16

6.2.7 AGD Operations Pty Ltd (2001–2009) .................................................................................. 17

6.3 Historical Resource and Reserve Estimates .................................................................................... 21


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6.4 Historical Production ......................................................................................................................... 21

7 Geological Setting and Mineralisation ...................................................................... 23

7.1 Regional Geology.............................................................................................................................. 23 7.2 Property Geology .............................................................................................................................. 25

7.3 Stratigraphy of the Costerfield Formation ......................................................................................... 27

7.4 Structural Geology of the South Costerfield area ............................................................................. 29

8 Deposit Types ............................................................................................................. 30

8.1 Property Mineralisation ..................................................................................................................... 30

8.2 Deposit Mineralisation ....................................................................................................................... 33

9 Exploration .................................................................................................................. 38

9.1 Costean/Trenching ............................................................................................................................ 38 9.2 Petrophysical Analysis ...................................................................................................................... 38

9.3 Geophysics ....................................................................................................................................... 38

9.3.1 Ground Geophysics ............................................................................................................... 38

9.3.2 Airborne Geophysics ............................................................................................................. 39

9.4 Geochemistry .................................................................................................................................... 39

9.4.1 Mobile Metal Ion (MMI) .......................................................................................................... 39 9.4.2 Bedrock Geochemistry .......................................................................................................... 39

9.5 Aerial Photogrammetry Survey ......................................................................................................... 46

10 Drilling ......................................................................................................................... 47

10.1 Mandalay Resources (2009–Present) .............................................................................................. 47

10.2 2009/2010 ......................................................................................................................................... 47

10.3 2010/2011 ......................................................................................................................................... 47 10.4 2011/2012 ......................................................................................................................................... 47

10.5 2012/2013 Cuffley Lode Drilling ........................................................................................................ 48

10.6 2013/2014 Cuffley /N Lode Drilling ................................................................................................... 50

10.7 Drilling Methods ................................................................................................................................ 50

10.8 Collar Surveys ................................................................................................................................... 53

10.9 Downhole Surveys ............................................................................................................................ 53 10.10 Logging Procedures .......................................................................................................................... 53

10.11 Drilling Pattern and Quality ............................................................................................................... 54

10.11.1 Augusta ............................................................................................................................. 54

10.11.2 Cuffley ............................................................................................................................... 54

10.12 Interpretation of Drilling Results ........................................................................................................ 54

10.13 Factors that could Materially Impact Accuracy of Results ................................................................ 55

11 Sample Preparation, Analyses, and Security ........................................................... 57

11.1 Sampling Techniques........................................................................................................................ 57

11.1.1 Diamond Core Sampling ....................................................................................................... 57

11.1.2 Underground Face Sampling ................................................................................................ 57

11.2 Data Spacing and Distribution .......................................................................................................... 58


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11.3 Testing Laboratories ......................................................................................................................... 58

11.4 Sample Preparation .......................................................................................................................... 58 11.5 Sample Analysis ................................................................................................................................ 59

11.6 Laboratory Reviews .......................................................................................................................... 59

11.7 Assay Quality Assurance and Quality Control .................................................................................. 60

11.7.1 Standard Reference Material ................................................................................................ 60

11.7.2 Blank Material ........................................................................................................................ 66

11.7.3 Duplicate Assay Statistics ..................................................................................................... 67 11.7.4 Check Assay Programme – Sample Pulps ........................................................................... 68

11.8 Sample Transport and Security ........................................................................................................ 72

11.9 Conclusions ....................................................................................................................................... 72

12 Data Verification ......................................................................................................... 73

13 Mineral Processing and Metallurgical Testing ......................................................... 75

13.1 Metallurgical Testing ......................................................................................................................... 75

13.1.1 Historical Test work – Augusta Ore test work vs actual plant performance .......................... 75

13.1.2 Historical test work – Cuffley ore vs Augusta ore test work results ...................................... 78 13.1.3 Cuffley vs Augusta Lode Mineralisation ................................................................................ 78

13.1.4 Cuffley vs Augusta Hardness Comparison ........................................................................... 79

13.1.5 Cuffley vs Augusta Test work Recovery Conclusions ........................................................... 80

13.2 Recent test work – Cuffley South ...................................................................................................... 81

13.3 Processing Recovery Data on Cuffley vs Augusta orebody ............................................................. 84

14 Mineral Resource Estimates ...................................................................................... 87

14.1 Introduction ....................................................................................................................................... 87

14.2 Diamond Drillhole and Underground Face Sample Statistics ........................................................... 88

14.3 Data Interpretation and Domaining ................................................................................................... 91

14.4 Grade Capping .................................................................................................................................. 91

14.5 Estimation Domain Boundaries ......................................................................................................... 92

14.6 Vein Orientation Domains ................................................................................................................. 94 14.7 Bulk Density Determinations ............................................................................................................. 97

14.8 Variography ....................................................................................................................................... 98

14.9 Estimation Parameters ...................................................................................................................... 98

14.10 Block Model Estimation ................................................................................................................... 106

14.11 Block Model Validation .................................................................................................................... 106

14.12 Mineral Resource Classification ...................................................................................................... 110 14.13 Mineral Resources .......................................................................................................................... 111

14.14 Cut-off Grade Calculations .............................................................................................................. 117

14.15 Reconciliation .................................................................................................................................. 117

14.16 Comparison with Previous Mineral Resource Estimate .................................................................. 118

14.17 Other Material Factors .................................................................................................................... 119

15 Mineral Reserve Estimate ........................................................................................ 120


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15.1 Modifying Factors ............................................................................................................................ 120

15.1.1 Mining Dilution and Recovery ............................................................................................. 120 15.1.2 Mine Design and Planning Process .................................................................................... 121

15.1.3 Cut-off Grade ....................................................................................................................... 123

16 Mining Methods ........................................................................................................ 124

16.1 Introduction ..................................................................................................................................... 124

16.2 Geotechnical ................................................................................................................................... 125

16.2.1 Augusta Overview ............................................................................................................... 125

16.2.2 Cuffley Overview ................................................................................................................. 126 16.2.3 Rock stress .......................................................................................................................... 127

16.2.4 Stope Design Criteria .......................................................................................................... 127

16.2.5 Ground Support Requirements ........................................................................................... 128

16.2.6 Incline/Decline Location ...................................................................................................... 129

16.2.7 References .......................................................................................................................... 129

16.3 Augusta Mine Design ...................................................................................................................... 129 16.3.1 Method Selection ................................................................................................................. 129

16.3.2 Method Description ............................................................................................................. 129

16.3.3 Materials Handling ............................................................................................................... 131

16.4 Augusta Mine Design Guidelines .................................................................................................... 132

16.4.1 Design Parameters .............................................................................................................. 132

16.4.2 Mining Sequence ................................................................................................................. 132 16.4.3 Level Development .............................................................................................................. 133

16.4.4 Vertical Development .......................................................................................................... 133

16.4.5 Augusta Design Inventory ................................................................................................... 134

16.5 Cuffley Mine Design ........................................................................................................................ 135

16.6 Cuffley Mine Design Guidelines ...................................................................................................... 136 16.6.1 Design Parameters .............................................................................................................. 136

16.6.2 Mining Sequence ................................................................................................................. 136

16.6.3 Level Development .............................................................................................................. 136


16.6.5 Stoping ................................................................................................................................ 136

16.6.6 Materials Handling ............................................................................................................... 137 16.6.7 Cuffley Design Inventory ..................................................................................................... 137

16.7 Ventilation ....................................................................................................................................... 138

16.7.1 Augusta Ventilation Circuit .................................................................................................. 138

16.7.2 Cuffley Ventilation Circuit .................................................................................................... 139

16.7.3 Airflow Requirements .......................................................................................................... 139

16.7.4 Rise Sizes ........................................................................................................................... 140 16.7.5 Fan Duties ........................................................................................................................... 141

16.7.6 References .......................................................................................................................... 141

16.8 Backfill 141


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16.9 Mine Services and Infrastructure .................................................................................................... 142

16.9.1 Electrical and Communications ........................................................................................... 142 16.9.2 Compressed Air ................................................................................................................... 142

16.9.3 Process and Potable Water................................................................................................. 143

16.9.4 Explosives and Magazine.................................................................................................... 143

16.9.5 Emergency Egress .............................................................................................................. 143

16.9.6 Refuge Chambers ............................................................................................................... 144

16.10 Augusta Hydrogeology/Dewatering ................................................................................................ 144 16.10.1 Hydrogeology .................................................................................................................. 144

16.10.2 Dewatering ...................................................................................................................... 145

16.11 Cuffley Hydrogeology /Dewatering ................................................................................................. 146

16.11.1 Hydrogeology .................................................................................................................. 146

16.11.2 Dewatering ...................................................................................................................... 147

16.12 Mineral Reserve Schedule .............................................................................................................. 147 16.12.1 Development Schedule ................................................................................................... 147

16.12.2 Equipment Requirements ............................................................................................... 148

16.12.3 Personnel ........................................................................................................................ 148

16.13 Schedule Summary ......................................................................................................................... 149

17 Recovery Methods .................................................................................................... 151

17.1 Brunswick Processing Plant ............................................................................................................ 151

17.1.1 Crushing and Screening Circuit .......................................................................................... 151 17.1.2 Milling Circuit ....................................................................................................................... 151

17.1.3 Flotation Circuit ................................................................................................................... 152

17.1.4 Concentrate Thickening and Filtration ................................................................................ 152

17.1.5 Tailings Circuit ..................................................................................................................... 152

17.1.6 Throughput .......................................................................................................................... 152 17.1.7 Recovery ............................................................................................................................. 153

17.1.8 Concentrate Grade .............................................................................................................. 153

17.2 Plant Upgrades ............................................................................................................................... 158

17.2.1 Crushing and Screening Circuit .......................................................................................... 158

17.2.2 Milling Circuit ....................................................................................................................... 158

17.2.3 Flotation Circuit ................................................................................................................... 158 17.2.4 Concentrate Thickening and Filtration ................................................................................ 158

17.2.5 Tailings Circuit ..................................................................................................................... 158

17.2.6 Reagent Mixing and Storage ............................................................................................... 159

17.3 Recovery ......................................................................................................................................... 159

17.4 Services .......................................................................................................................................... 159

17.4.1 Water ................................................................................................................................... 159 17.4.2 Air ........................................................................................................................................ 159

17.4.3 Power .................................................................................................................................. 160


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18 Project Infrastructure ............................................................................................... 161

18.1 Surface Infrastructure...................................................................................................................... 161

18.2 Augusta Mine Underground Infrastructure ...................................................................................... 163 18.2.1 Ventilation System ............................................................................................................... 163

18.2.2 Dewatering System ............................................................................................................. 163

18.2.3 Infrastructure ....................................................................................................................... 163

18.3 Tailings Storage .............................................................................................................................. 163

18.4 Power Supply .................................................................................................................................. 164

18.5 Power Reticulation .......................................................................................................................... 164 18.6 Water Supply ................................................................................................................................... 165

18.7 Waste Rock Storage ....................................................................................................................... 166

18.8 Water Management......................................................................................................................... 166

18.9 Augusta to Brunswick ROM Pad Transport .................................................................................... 167

18.10 Diesel Storage ................................................................................................................................. 167

18.11 Explosive Storage ........................................................................................................................... 167 18.12 Maintenance Facilities .................................................................................................................... 167

18.13 Housing and Land ........................................................................................................................... 168

19 Market Studies and Contracts ................................................................................. 169

19.1 Concentrate Transport .................................................................................................................... 169

19.2 Marketing ........................................................................................................................................ 169

19.3 Contracts ......................................................................................................................................... 172

20 Environmental Studies, Permitting, and Social or Community Impact ............... 173

20.1 Environment and Social Aspects .................................................................................................... 173

20.1.1 Mine Ventilation ................................................................................................................... 173

20.1.2 Water Disposal .................................................................................................................... 173

20.1.3 Waste Rock ......................................................................................................................... 174

20.1.4 Tailings Disposal ................................................................................................................. 175

20.1.5 Air Quality ............................................................................................................................ 175 20.1.6 Groundwater ........................................................................................................................ 175

20.1.7 Noise ................................................................................................................................... 177

20.1.8 Blasting and Vibration ......................................................................................................... 177

20.1.9 Native Vegetation ................................................................................................................ 177

20.1.10 Visual Amenity ................................................................................................................ 177

20.1.11 Heritage ........................................................................................................................... 178 20.1.12 Community ...................................................................................................................... 178

20.1.13 Mine Closure and Revegetation ..................................................................................... 179

20.2 Regulatory Approvals ...................................................................................................................... 179

20.2.1 Work Plan Variation ............................................................................................................. 179

20.2.2 Other Permitting .................................................................................................................. 180

21 Capital and Operating Costs ................................................................................... 181


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21.1 Capital Costs ................................................................................................................................... 181

Total may not add-up due to rounding ............................................................................................ 181 21.1.1 Capital Lateral Development ............................................................................................... 182


21.1.3 Infrastructure ....................................................................................................................... 182

21.1.4 Mobile Plant ......................................................................................................................... 182

21.1.5 Processing Plant ................................................................................................................. 183

21.1.6 Drilling.................................................................................................................................. 183 21.1.7 Closure ................................................................................................................................ 183

21.2 Operating Costs .............................................................................................................................. 184

21.2.1 Lateral Development ........................................................................................................... 184

21.2.2 Labour Costs ....................................................................................................................... 185

21.2.3 Production Stoping .............................................................................................................. 185

21.2.4 Augusta to Brunswick ROM Pad trucking ........................................................................... 185 21.2.5 Processing ........................................................................................................................... 185

21.3 Concentrate Selling Expenses ........................................................................................................ 185

21.4 Royalties and Compensation .......................................................................................................... 186

22 Economic Analysis ................................................................................................... 187

22.1 Introduction ..................................................................................................................................... 187

22.2 Principal Assumptions ..................................................................................................................... 187

22.2.1 Metal Sale Prices ................................................................................................................ 187 22.2.2 Concentrate & Gold Sales ................................................................................................... 188

22.2.3 Exchange Rate .................................................................................................................... 188

22.2.4 Taxes ................................................................................................................................... 188

22.2.5 Royalties/Agreements ......................................................................................................... 188

22.2.6 Reclamation ........................................................................................................................ 188 22.2.7 Project Financing ................................................................................................................. 189

22.3 Economic Summary ........................................................................................................................ 189

22.3.1 Cash Flow Forecast ............................................................................................................ 189

22.3.2 NPV ..................................................................................................................................... 192

22.3.3 Sensitivity ............................................................................................................................ 192

23 Adjacent Properties .................................................................................................. 194

23.1 General Statement about Adjacent Properties ............................................................................... 194

24 Other Relevant Data and Information ..................................................................... 196

24.1 Remnant Mining .............................................................................................................................. 196

25 Interpretation and Conclusions............................................................................... 197

25.1 Geology ........................................................................................................................................... 197

25.2 Mining 198

26 Recommendations ................................................................................................... 199

26.1 Geology ........................................................................................................................................... 199


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26.2 Mining 199

27 References ................................................................................................................ 200

List of Tables Table ES-2-1: Summary of the Augusta, Cuffley and Brunswick Mineral Resource, inclusive of Mineral

Reserve ...................................................................................................................................... xii

Table 2-1: List of Qualified Persons .............................................................................................................. 4

Table 4-1: Granted Tenement Details ........................................................................................................... 7 Table 6-1: Historical Drilling Statistics for the Costerfield Property ............................................................. 15

Table 6-2: Historical Mine Production. Mandalay Resources – Costerfield Project .................................... 21

Table 6-3: Historical Mineral Resources. Mandalay Resources – Costerfield Project ................................ 22

Table 6-4: Historical Mineral Reserves. Mandalay Resources – Costerfield Project .................................. 22

Table 9-1: Auger drillhole totals by prospect ............................................................................................... 44

Table 10-1: Drillhole Summary ...................................................................................................................... 47 Table 10-2: Significant Cuffley Lode drillhole intercepts: 2012/2013 ............................................................ 49

Table 10-3: Significant Cuffley Lode drillhole intercepts: 2014 ..................................................................... 50

Table 11-1: Summary of Onsite duplicate statistics ...................................................................................... 67

Table 11-2: Summary of Onsite versus ALS Gold Duplicate Statistics ......................................................... 69

Table 11-3: Summary of Onsite versus ALS Antimony Duplicate Statistics ................................................. 69

Table 13-1: Test work vs Plant Reconciled Recovery Results for Antimony ................................................ 77 Table 13-2: Test work vs Plant Reconciled Recovery Results for Gold ........................................................ 77

Table 13-3: Test work Schedule and Grades ................................................................................................ 78

Table 13-4: Antimony Recovery Comparison Cuffley Lode vs Augusta Orebody ........................................ 81

Table 13-5: Gold Recovery Comparison Cuffley Lode vs Augusta Orebody ................................................ 81

Table 14-1: Changes made to models at year-end 2014 .............................................................................. 87 Table 14-2: Face and diamond drilling sample statistics .............................................................................. 89

Table 14-3: Sample statistics of the data selected for analysis .................................................................... 90

Table 14-4: Accumulated Gold statistics before and after top cuts ............................................................... 91

Table 14-5: Variogram Model Parameters .................................................................................................. 104

Table 14-6: Search Parameters for first estimation run .............................................................................. 105

Table 14-7: Block Model Dimensions .......................................................................................................... 106 Table 14-8: Mine Planning Regularised Block Model Dimensions ............................................................. 106

Table 14-9: Declustered (50 x 50 m) thickness weighted, composited sample grades and width compared to model grades ....................................................................................................................... 107

Table 14-10: Summary of the Augusta, Cuffley and Brunswick Mineral Resource, inclusive of Mineral Reserve .................................................................................................................................... 112

Table 14-11: Current Estimates Compared with the previous (December 2013) estimates for E Lode, W Lode, N Lodes, P Lodes, Cuffley lodes and Brunswick ........................................................... 119

Table 15-1: Mineral Reserves at Costerfield, as of 31 December 2014 ..................................................... 120

Table 15-2: Augusta Mine Recovery and Dilution Assumptions ................................................................. 121

Table 16-1: Mine Design Parameters.......................................................................................................... 132


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Table 16-2: Augusta Stope and Development Inventory ............................................................................ 134

Table 16-3: Mine Design Parameters.......................................................................................................... 136 Table 16-4: Cuffley Stope and Development Inventory .............................................................................. 137

Table 16-5: Augusta Minimum Airflow Estimate ......................................................................................... 140

Table 16-6: Cuffley LOM Minimum Airflow Estimate (per circuit) ............................................................... 140

Table 16-7: Cuffley Rise sizes ..................................................................................................................... 140

Table 16-8: Current Augusta Licence Maximum Quantities and Types of Explosives ............................... 143

Table 16-9: Jumbo Production Rates .......................................................................................................... 147 Table 16-10: Handheld Production Rates ..................................................................................................... 147

Table 16-11: Underground Mobile Equipment Fleet ..................................................................................... 148

Table 16-12: Personnel on Payroll by Department ....................................................................................... 149

Table 16-13: Underground Shift Mining Personnel ....................................................................................... 149

Table 16-14: Summary of design physicals .................................................................................................. 150

Table 17-1: Brunswick Mill Key Processing Results before Crushing Circuit Upgrade - February 2012 to August 2012 ............................................................................................................................. 156

Table 17-2: Brunswick Mill Key Processing Results after Crushing Circuit Upgrade - September 2012 to December 2013 ........................................................................................................................ 156

Table 17-3: Brunswick Mill Key Processing Results - January 2014 to December 2014 ........................... 157

Table 17-4: Brunswick Mill Concentrate Production before Crushing Circuit Upgrade - February 2012 to August 2012 ............................................................................................................................. 157

Table 17-5: Brunswick Mill Concentrate Production after Crushing Circuit Upgrade - September 2012 to December 2013 ........................................................................................................................ 157

Table 17-6: Brunswick Mill Concentrate Production – January 2014 to December 2014 ........................... 157

Table 18-1: Current Augusta Licence Maximum Quantities and Types of Explosives ............................... 167

Table 20-1: Permit Requirements ............................................................................................................... 180 Table 21-1: Costerfield Operation – Capital Cost Estimate ........................................................................ 181

Table 21-2: Operating Cost Inputs .............................................................................................................. 184

Table 21-3: Costerfield Operation - Operating Cost Summary ................................................................... 184

Table 21-4: Summary of Development requirements ................................................................................. 184

Table 22-1: Project Criteria ......................................................................................................................... 187

Table 22-2: Project Economics ................................................................................................................... 189 Table 22-3: Estimated Pre-Tax Cash Flow Summary ................................................................................. 190

Table 22-4: Sensitivity at 5% NPV After-Tax .............................................................................................. 192

Table 23-1: Augusta Mine Adjacent Properties (Source DSDBI Geovic, 2015) ......................................... 194

Table 23-2: Distance from the Augusta Mine Site to Significant Central Victoria Mining Landmarks ........ 194

List of Figures Figure 4-1: Costerfield Operation Location .................................................................................................... 6

Figure 4-2: Plan of Area – Mining Licence No: 4644 ...................................................................................... 8

Figure 4-3: Current AGD Mining and Exploration Lease Boundaries ............................................................ 9 Figure 5-1: Monthly Average Temperature and Rainfall .............................................................................. 13


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Figure 5-2: Augusta Box-cut, Portal and Workshop ..................................................................................... 13

Figure 5-3: Aerial View of Brunswick Processing Plant and Brunswick Open Pit ........................................ 14 Figure 7-1: Geological Map of the Heathcote – Colbinabbin – Nagambie Region ...................................... 24

Figure 7-2: Costerfield Property Geology and Old Workings ....................................................................... 26

Figure 7-3: Regional stratigraphic chart of the Costerfield region illustrating age relationships of defined sedimentary succession of the Darraweit Guim Province ......................................................... 27

Figure 7-4: Stratigraphic column of the Costerfield Formation illustrating the relative positions of newly defined (informal) stratigraphic units .......................................................................................... 28

Figure 8-1: Paragenetic history and vein genesis of the Costerfield region ................................................. 31

Figure 8-2: Two views of E Lode 1,070 Level South, looking south with annotation on right-hand side ..... 32

Figure 8-3: Close-up of mineralisation in west dipping E Lode 1070 Level looking south ........................... 32

Figure 8-4: Cuffley deposit looking North – 895L C2 North lode .................................................................. 34

Figure 8-5: Deposit wireframes and underground development .................................................................. 35

Figure 8-6: Structural relationship between the Costerfield Thrust and the Augusta and Cuffley Lode systems ...................................................................................................................................... 36

Figure 8-7: Costerfield geological systems showing Brunswick, Bombay/Minerva, Cuffley and Augusta deposits ...................................................................................................................................... 37

Figure 9-1: Augusta South Aircore Programme Results with Lode Locations ............................................. 40

Figure 9-2: Augusta South (Margaret Zone) – Auger drill 2011 bedrock geochemistry: antimony contours42

Figure 9-3: Auger geochemistry results displayed as antimony contours .................................................... 43 Figure 10-1: Known drillhole collar locations for Augusta and Cuffley ........................................................... 52

Figure 10-2: Example of cross section at 4,300 mN post drilling geological interpretation of the Augusta deposit ........................................................................................................................................ 55

Figure 11-1: G901-8 Gold Standard Reference Material – Assay Results for 2014 ...................................... 61

Figure 11-2: G906-8 Gold Standard Reference Material – Assay Results for 2014 ...................................... 62 Figure 11-3: MR11-01 Gold Standard Reference Material – Assay Results for 2014 ................................... 62

Figure 11-4: MR04/06 Gold Standard Reference Material – Assay Results for 2014 ................................... 63

Figure 11-5: MR30/20 Gold Standard Reference Material – Assay Results for 2014 ................................... 63

Figure 11-6: AGD08-01 Antimony Standard Reference Material – Assay Results for 2014 .......................... 64

Figure 11-7: MR11-01 Antimony Standard Reference Material – Assay Results for 2014 ............................ 64

Figure 11-8: MR04/06 Antimony Standard Reference Material – Assay Results for 2014 ............................ 65 Figure 11-9: MR30/20 Antimony Standard Reference Material – Assay Results for 2014 ............................ 65

Figure 11-10: Blank assay results for 2014 ...................................................................................................... 66

Figure 11-11: Scatter Plot for Onsite Gold Duplicates (g/t) .............................................................................. 67

Figure 11-12: Relative Paired Difference Plot for Onsite Gold Duplicates (g/t) ............................................... 68

Figure 11-13: Scatter Plot for Onsite versus ALS Gold Duplicates (g/t) .......................................................... 70

Figure 11-14: Scatter Plot for Onsite versus ALS Antimony Duplicates (%) .................................................... 70 Figure 11-15: Relative Paired Difference Plot for Onsite versus ALS Gold Duplicates (g/t) ............................ 71

Figure 11-16: Relative Paired Difference Plot for Onsite versus ALS Antimony Duplicates (%) ..................... 71

Figure 12-1: Cuffley Main lower drive south end showing mineralised structures (gold bearing quartz – pink, stibnite – yellow) and bedding .................................................................................................... 74

Figure 13-1: Amdel QEMSCAN Mineral Abundance ...................................................................................... 79 Figure 13-2: Grind Establishment Test work Comparisons at 75 µm............................................................. 80


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Figure 13-3: Grind Establishment Test work Comparisons at 53 µm............................................................. 80

Figure 13-4: Cuffley long section with metallurgical sample locations ........................................................... 83 Figure 13-5: Mill performance on Augusta Ore and Cuffley/Augusta Mix ...................................................... 84

Figure 13-6: Percentage of Cuffley and Augusta Ore Mined in 2014 ............................................................ 84

Figure 13-7: Mill recoveries and throughput for Cuffley/Augusta ore blend 2014 .......................................... 85

Figure 13-8: 2014 Cuffley/Augusta Blend Feed Grade and Recoveries ........................................................ 86

Figure 13-9: Gold Grade vs Recovery Relationship on Cuffley/Augusta blend (2014) .................................. 86

Figure 14-1: Long section of data collected for the sample comparison analysis showing Face sample (Red) and Drillhole (Blue) Data ............................................................................................................ 90

Figure 14-2: Long section of NM Lode showing fault intercepts identified as controls on mineralisation ...... 91

Figure 14-3: CM Lode Estimation domain boundaries and composite samples ............................................ 93

Figure 14-4: NM Lode Estimation domain boundaries and composite samples ............................................ 93

Figure 14-5: CM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor ......................................................................................................................................... 94

Figure 14-6: NM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor ......................................................................................................................................... 95

Figure 14-7: NE Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor ......................................................................................................................................... 95

Figure 14-8: Cuffley Deeps Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor ....................................................................................................................... 96

Figure 14-9: P1 and P2 Lodes Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor ....................................................................................................................... 96

Figure 14-10: Bulk Density Determinations ...................................................................................................... 97

Figure 14-11: Nugget variogram for gold accumulation in NM Lode ................................................................ 99

Figure 14-12: Nugget variogram for antimony accumulation in NM Lode ...................................................... 100 Figure 14-13: Nugget variogram for True Thickness in NM Lode .................................................................. 100

Figure 14-14: Primary direction variogram for gold accumulation in NM Lode .............................................. 101

Figure 14-15: Secondary direction variogram for gold accumulation in NM Lode ......................................... 101

Figure 14-16: Primary direction variogram for antimony accumulation in NM Lode ...................................... 102

Figure 14-17: Secondary direction variogram for antimony accumulation in NM Lode ................................. 102

Figure 14-18: Primary direction variogram for true thickness in NM Lode ..................................................... 103 Figure 14-19: Secondary direction variogram for true thickness in NM Lode ................................................ 103

Figure 14-20: CE Lode Gold Accumulation Swath Plot by Northing .............................................................. 108

Figure 14-21: CE Lode Antimony Accumulation Swath Plot by Northing ....................................................... 108

Figure 14-22: CM Lode Gold Accumulation Swath Plot by Northing ............................................................. 108

Figure 14-23: CM Lode Antimony Accumulation Swath Plot by Northing ...................................................... 109

Figure 14-24: NE Lode Gold Accumulation Swath Plot by Northing .............................................................. 109 Figure 14-25: NE Lode Antimony Accumulation Swath Plot by Northing ....................................................... 109

Figure 14-26: NM Lode Gold Accumulation Swath Plot by Northing ............................................................. 110

Figure 14-27: NM Lode Antimony Accumulation Swath Plot by Northing ...................................................... 110

Figure 14-28: NM Lode Resource Block Model with Resource Category Boundaries .................................. 113

Figure 14-29: NE Lode Resource Block Model with Resource Category Boundaries ................................... 114 Figure 14-30: CM Main Lode Block Model with Resource Category Boundaries .......................................... 114


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Figure 14-31: CE Lode Block Model with Resource Category Boundaries.................................................... 115

Figure 14-32: CD Lode Block Model with Resource Category Boundaries ................................................... 115 Figure 14-33: AS Lode Block Model with Resource Category Boundaries .................................................... 116

Figure 14-34: P1 and P2 Lode Block Models with Resource Category Boundaries ...................................... 116

Figure 14-35: Mill Reconciliation – Dry Tonnes .............................................................................................. 117

Figure 14-36: Mill Reconciliation – Gold (oz) ................................................................................................. 118

Figure 14-37: Mill Reconciliation – Antimony (t) ............................................................................................. 118

Figure 15-1: Overlaps between as-built and stope shapes which has been considered as depleted ......... 122 Figure 15-2: Overlaps between stope and development shapes which have been considered as depleted122

Figure 16-1: Section of Proposed Augusta and Cuffley Underground Design ............................................. 124

Figure 16-2: Plan View of Proposed Augusta and Cuffley Underground Design ......................................... 125

Figure 16-3: Geotechnical Domains – Section 5,010 mN ............................................................................ 127

Figure 16-4: Augusta Long-hole CRF Stoping Method ................................................................................ 130

Figure 16-5: Long section of Augusta stoping between 1,015 mRL and 989 mRL ...................................... 132 Figure 16-6: Typical Section Looking North through Augusta Level Development ...................................... 133

Figure 16-7: Long Section of Cuffley stoping areas ..................................................................................... 137

Figure 16-8: Proposed Cuffley Ventilation Circuit ........................................................................................ 139

Figure 16-9: Conceptual cross section of the groundwater flow paths ........................................................ 145

Figure 16-10: Groundwater elevation contour map of the areas surrounding the Augusta Mine .................. 145

Figure 17-1: Brunswick Processing Plant Summary Flowsheet ................................................................... 154 Figure 17-2: Brunswick Processing Feed Tonnages - January 2012 to December 2014 ............................ 155

Figure 18-1: Augusta Mine Site .................................................................................................................... 161

Figure 18-2: Brunswick Site Area ................................................................................................................. 162

Figure 19-1: Estimate of Global Antimony Demand by End-Use Segment.................................................. 170

Figure 19-2: Antimony metal prices 2010–2015 ........................................................................................... 170 Figure 19-3: World Antimony production and consumption, 2000–2020 (t Antimony) ................................. 171

Figure 20-1: Groundwater elevation contour map of the areas surrounding the Augusta Mine as at February 2014 ......................................................................................................................................... 176

Figure 22-1: Sensitivity Analysis ................................................................................................................... 193

Figure 23-1: Augusta Mine Adjacent Properties (Source DSDBI Geovic, 2015) ......................................... 194


SRK Consulting Page 1

2 Introduction SRK Consulting (Australasia) Pty Ltd (SRK) has worked in conjunction with 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 has been the sole ore source for the Brunswick Processing Plant until December 2013. At this time ore production started from the Cuffley Deposit located approximately 500 m to the north of the Augusta mine workings. The Cuffley deposit will be mined in conjunction with remanent 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;

• A conventional flotation mill (Brunswick) with a current name plate capacity of approximately 130,000 t/year of feed; and

• 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 the 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 by the List of Abbreviations.

2.1 Scope of Work The Scope of Work, as defined in a letter of engagement executed during November 2014 and subsequent discussions between Mandalay and SRK, includes the review of documents provided by Mandalay and prepared 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 of the Brunswick Mineral Resource Estimate;

• Review the mine design and mining schedules for Augusta prepared by Mandalay;

• Review the mine design and mining schedules for Cuffley Lode prepared by Mandalay;

• Review and comment on the project infrastructure aspects;

• Review and comment on the metallurgical and processing aspects;

• Review of the environmental considerations;

• Review of the capital and operating cost estimates;

• Review of the financial modelling; and

• Recommendations for additional work.



2.2 Work Programme 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 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 2015.

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 2014 and additional information provided by Mandalay throughout the course of SRKs 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 2014.

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

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

Internal SRK Peer Review of this Technical Report was completed by Anne-Marie Ebbels.



Anne-Marie Ebbels, Principal Consultant (Mining), BEng (Mining), MAusIMM (CP) (No 111006), conducted a peer review of non-geological aspects of this Technical report. Anne-Marie Ebbels is by virtue of her education, membership to 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 QP as defined in NI 43-101 for this report. The table defines the areas of responsibility for the QP 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 were completed by Mandalay personnel with supervision and review by SRK. The Mandalay personnel with significant roles were; Chris Davies, who undertook the geological modelling and resource estimation work; Mohammad Khani who undertook the reserves, scheduling and mine design work; Damon Buchanan who undertook the mineral processing and metallurgical test work; and Shannon Green and April Westcott.

2.6 Declaration SRKs opinion contained herein and effective 31 December 2014 is based on information collected by SRK throughout the course of SRKs investigations, which in turn reflect various technical and economic conditions at the time of writing. Given the nature of the mining business, these conditions can change significantly over relatively short periods of time. Consequently, actual results may be significantly more or less favourable.

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

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

SRK is not an insider, associate or an 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 18 November 2014

BEng (Mining), FAusIMM, CP (Mining)

Sections 1–6, Section 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 18 November 2014 MSc Mathematics & Planning (Geostatistics),

FAusIMM Sections 7–12, Section 14

Simon Walsh Principal Metallurgist & Director

Simulus Engineers (SRK Associate) *

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

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

*Simon Walsh was supported by Brett Muller who was the QP for SRK (2014) and completed a site visit on 10 March 2014.



3 Reliance on Other Experts 3.1 Marketing

Marketing information for this report, specifically Section 19, relies entirely on information by Roskill Information Services Ltd. The two reports referenced are as follows:

• Roskill, 2012. Antimony: Global Industry Markets and Outlook, 11th edition, 2012. Available from http://www.roskill.com/reports/minor-and-light-metals/antimony; and

• Roskill, 2011. ANOCA Ltd: Study of the antimony market, 17 October 2011. Available from http://www.ancoa.com.au/RoskillCRT.pdf.

A specific marketing study was not completed for this report.


http://www.roskill.com/reports/minor-and-light-metals/antimony


4 Property Description and Location 4.1 Property Location

The Costerfield Operation is located within the Costerfield mining district of Central Victoria. The operations are located 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, the Brunswick Processing Plant, Bombay Tailings Storage Facility (TSF), Cuffley Lode deposit and associated infrastructure.

The Augusta Mine (Augusta) is located at latitude of 36° 52’ 27” south and longitude 144° 47’ 38” east. The Brunswick Processing Plant is located approximately 2 km north west 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, four Exploration Licences and one application 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 30/06/2014

EL3310 Costerfield Granted AGD Operations P/L 59.0 GRATS 17/09/1993 17/09/2015

EL4848 Antimony Creek Surrendered AGD Operations P/L 18.0

GRATS 28/01/2005

Surrendered June 2014. Reapplied 25/07/2014

Refer to ELA5519

ELA5519 Antimony Creek Application

Mandalay Resources Costerfield

Operations Pty Ltd

11.0 GRATS

Under Application

See note below

EL5432 Peels Track Granted AGD Operations P/L 10.0 GRATS 23/08/2012 22/08/2017

MIN5567 Splitters Creek Granted AGD Operations P/L 30 Ha 20/02/2013 21/02/2023

EL5464 Antimony Creek Granted


Operations Pty Ltd

0.96 Ha 13/2/2014 12/2/2016

EL5452 Antimony Creek Granted


Operations Pty Ltd

2.07 Ha 26/6/2013 27/1/2016

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, EL 5464 and EL 5452 as shown in Figure 4-2 and Figure 4-3.

On 24 March 2014 Mandalay Resources ‘put their case’ to the DSDBI requesting that ‘exceptional circumstances’ be granted and that tenement ELA 5519 be renewed for a further 2 years. The minister did not grant ‘exceptional circumstances’ and as a result Mandalay decided to surrender the licence and re-apply for the tenement, in the hope that it was granted back to the company. Mandalay Resources re-applied for the same area of ground on 25 June 2014 and at the time of writing is currently under application (Figure 4-3).



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



Figure 4-3: Current AGD Mining and Exploration Lease Boundaries Source: Mandalay, 2014.



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

The MRSD Act, which is administered by the DSDBI, 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 owned by third party landowners that are situated within the boundaries of MIN 4644.

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 was reviewed at the end of 2014 and will be set at AUD 3.106M once the consultation period is complete at the end of January 2015. The rehabilitation bond is reviewed by the DSDBI every two years or when a substantive Variation to the Work Plan is approved.

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

Additional Bonds will be linked to the work plan that has been submitted for MIN 5567.

The total bond is AUD2.541M.

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

4.6 Taxes Mandalay reports that, as at December 2014, there is approximately AUD20.6M in tax loss carry forwards for AGD that will effectively eliminate any income tax liability in the next few years. 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; and

• 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; and

• 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 DSDBI and are registered against the tenement.

There are no Native Title issues relevant to MIN 4644.

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 the Heathcote–Nagambie Road via McNicols Lane which comprises 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 the Heathcote–Nagambie Road, approximately one kilometre further 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 the Department of Environment and Primary Industries (DEPI) 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 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 m above sea level (ASL) in the east, along Wappentake Creek, to 288 m ASL 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 to 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

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

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

The Cuffley operations use the infrastructure associated with the current Augusta operations.

Figure 5-2: Augusta Box-cut, Portal and Workshop



The Brunswick Processing Plant consists of an approximately 140,000 tonnes per annum (tpa) gravity-flotation gold-antimony processing plant, with workshop facilities, site offices, TSFs, core shed and core farm. It currently produces an antimony-gold concentrate which is trucked to the Port of Melbourne, 130 km to the south. There it is transferred onto ships for export to foreign smelters. 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.

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

Electrical power for both operations is sourced from the main national electricity grid. The Costerfield Operation has an existing arrangement for supply of 2,000 kW (2,353 kVA) 22kV at the Augusta Mine and 1,000 kVA 240/415 V at the Brunswick Processing Plant.

Process water for the Brunswick Processing Plant is drawn from standing water in the Bombay TSF, supplemented by water from the old Brunswick Open Pit throughout the summer months. 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 off site by a local contractor.

In October 2014 construction commenced on an Evaporation facility called the Splitters Creek Evaporation Facility. The purpose of the facility is to evaporate groundwater extracted from the Mandalay Resources Costerfield Operations thereby maintaining dewatering rates from the underground workings.



6 History 6.1 Introduction

From the 1860s, beginning with the initial discovery of the Costerfield Reef, until 1953, several companies have developed and mined antimony deposits within the Costerfield area. Some underground diamond drilling is known to have occurred during the period of 1934 to 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 and Table 6-1 presents a summary of the historical drilling statistics by 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



6.2.1 Mid-East Minerals (1968–1971) From 1968 to 1969 the price of antimony rapidly rose from USD0.45 per pound to USD1.70. 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 programmes. 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.

6.2.2 Metals Investment Holdings (1971) A series of diamond drillholes were 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 drillholes were situated to the north of the Tait’s Mine (north of Augusta), of which minimal detail remains.

6.2.3 Victorian Mines Department (1975–1981) A series of diamond drillholes were 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.

6.2.4 Federation Resources NL (1983–2000) Federation Resources 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 conducted 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 conducted desktop studies on the area above the Augusta Mine, noting the anomalous results of the soil geochemistry programmes conducted by The Victorian Mines Department and Mid-East Minerals, but did not conduct drilling at this location.

6.2.5 Australian Gold Development NL/Planet resource JV (AGD) (1987–1988) Australian Gold Development NL conducted a short reverse circulation drilling (RC) programme in conjunction with their 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 serious contamination.

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



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

2001 AGDs drilling of 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 a NQ-sized drill bit, the purpose of which was to maximize core recoveries. Triple-tube drilling was also employed in areas to maximize recoveries. Cobar Drilling Company Pty Ltd, based in Rushworth, was contracted for the drilling programme. Less competent rock adjacent to the mineralisation was successfully recovered during this programme 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.

It was identified that 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 at 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. Drillhole 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 drillhole MH028, which contained 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 drillholes for 1,514 m were drilled by AGD as part of an infill and extension programme to 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 being 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 programme when compared to previous drilling programmes, even though the majority of drillholes were drilled in the weathered zone.



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

2004/2005 Between October 2004 and April 2005, AGD completed a 26-hole diamond-drilling programme at the Augusta Deposit. Apart from 5 m percussion pre-collars and 4 RC geotechnical holes, the holes were drilled by HQ triple-tube diamond drilling.

The objectives of the diamond-drilling programme were:

• Improvement in mineralisation definition by increasing drillhole density;

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

• Increasing the Mineral Resource and Mineral Reserve.

2006/2007 AGDs 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 metres of diamond drilling. This comprised the following drillholes:

• 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; and

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

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 with following 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, being drilled adjacent to the Augusta box cut.

On completion of the Brunswick and W Lode drilling, both the LM75 and the 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 2006. By the end of the second quarter all the surface infrastructure had been completed together with open cut mining of E and C lodes. Decline development commenced during June 2006 and underground production by the end of the third quarter of 2006.

2007/2008 AGDs 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; and

• 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 drillholes confirmed W Lode continuity down-dip, with true thicknesses ranging from 0.254 to 0.814 m at 22.50 to 89.26 g/t gold and 16.19 to 47.20% antimony;

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

• 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 to 0.293 m at 6.82 to 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 drillholes, totalling 2,207.2 m that were drilled on the northern section of the Augusta Deposit, particularly from 4,411 mN to 4,602 mN.

The 11 surface drillholes 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 half way 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 AGDs drilling activities throughout 2008 and 2009 comprised drilling along strike and down-dip from the existing Augusta resource. A total of 2,585.95 metres of diamond drilling was completed.

Drilling during 2008 and 2009 was concentrated on the definition of the W Lode resource. Five drillholes tested the depth extent of 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 using 3 mining methods, floor benching, cut and fill and long-hole 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 comprised 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 targeting 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 at Hirds Reef prospect within EL 4848.



Drilling during this reporting period has been concentrated on the definition of the W Lode resource. Four (4) drill holes tested the depth extent of W lode. Another 6 holes were designed as infill holes to test ore shoots and gather geotechnical data. The significant results of Augusta Deposit Drilling through this period have been tabulated below.

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

2010/2011

Drilling from July 2010 to June 2011 was undertaken on two projects, 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 is 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. In total 9,890.7 m of diamond coring and 732 m of auger drilling was undertaken as part of the two projects from July 2010 to June 2011.

Capital development reached 976 (RL) Level (200 m below surface). Ore drive development was carried out on E and W Lodes.

2011/2012

Drilling was undertaken on 4 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 programme is now more regional.

The Bedrock Geochemistry auger drilling project is revealing 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 was undertaken as part of 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 E and W Lodes.



2012/2013 On 7 February 2013 AGD Operations Pty Ltd underwent a name change to ‘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:

• 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; and

• Increase confidence in areas of N Lode that are easily accessible from 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 Resources Costerfield Operations Pty Ltd tenements at Costerfield from July 2012 to June 2013.

From July 2012 to June 2013 Capital development reached 878 mRL (300 m below surface) and ore drive development was carried out on E, W and N Lodes.

6.3 Historical Resource and Reserve Estimates Mandalay Resources has reported Mineral Resources and Reserves from 2010 – 2014. Mineral Resources are presented in Table 6-3 and Mineral Reserves are presented in Table 6-4.

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

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 tonnes were extracted producing 25,000 ounces of gold and 4,200 tonnes of antimony. The production record from the start of 2010 to end of 2014 is shown in Table 6-2.

Table 6-2: 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



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

Measured Resource Indicated Resource Inferred Resource

Effective Date

USD/ Oz Au

USD/ Oz Sb

Cut-off

Grade (AuEq

g/t)

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

Table 6-4: Historical Mineral Reserves. Mandalay Resources – Costerfield Project

Proven Reserves Probable Reserves Total Reserves

Effective Date

USD/ Oz Au

USD/ Oz Sb

Cut-off

Grade (AuEq

g/t)

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

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



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; and

• Robinsons – Browns (R-B) 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 system(s) 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 (NW) and northeast (NE) faulting) has severely disrupted the mineralised system in the north.

Host rocks are Silurian Costerfield Formation siltstones (Costerfield siltstone). These siltstones are at least 600 m thick and are the oldest exposed rocks in the local area. Their most distinctive feature is intense bioturbation, with fucoidal textures appearing as irregular, elongate dark patches up to 5 mm wide and at least 25 mm long. Apart from these markings, there are few fossils or sedimentary structures evident.

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 past 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 to 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 (drillhole 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. 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 (see Figure 7-3). The Mt Ida Formation records the terminal faze 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 (that assigned by Talent, 1965) is chosen to be used within this report instead of the later re-assigned name of ‘Costerfield Siltstone’ (as re-defined by VandenBurg, 1988). It is recommended that the ‘Siltstone’ nomenclature be abandoned as it has become a misleading term inferring the unit is composed of siltstone dominant lithologies. Rather, the unit consists of a dominantly 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 mid-way 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, Quartzite Beds. Lithostratigraphic units defined from the Upper Costerfield Formation are named the Lower Siltstone Unit, Augusta beds and the Upper Siltstone Unit (see 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 enormous collection of geological data from the South Costerfield area. From this, construction of highly 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 in the range of 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 extremely 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 extremely 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 to represent 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 in the order of around 1 km.

An additional series of brittle faults are observed within this thrust system striking in a NNE direction (i.e. the East Fault). These faults have a sub-vertical dip and are generally observed as one to two metre 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 tentatively estimated to be on the scale of 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 occur 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 in uncertain as their bedding-parallel characteristics mean estimations of 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.

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 (Figure 8-6). Drilling in 2011 successfully intersected a displaced lode below the fault, now known as the Cuffley Main Lode.

8.1 Property Mineralisation The economic 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 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 August and Cuffley deposits is illustrated 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 Section 7-4 of report for detailed description of 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 on the 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 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 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 and Cuffley deposits of the Costerfield region. The Augusta Deposit currently comprises nine lodes: E Lode, W Lode, N Main Lode, NW Lode, NE Lode, NSW Lode, P1 Lode, P2 Lode, K Lode, and B Lode (Figure 8-5). The Cuffley Deposit currently being mined is comprised of four lode structures, Cuffley Main (CM Lode), Cuffley East (CE Lode), Cuffley Deeps (CD Lode) and Alison South (AS Lode).

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 less than 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 to 80˚ at around the 900 mRL. Its true thickness ranges from less than 0.1 m up to 2.7 m, with an average width of 0.36 m where it has been mined.

N Main Lode lies approximately 90 m to the west of E Lode. N Main Lode was previously referred to 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 less than 0.1 m up to 2.7 m, with an average width of 0.3 m where it has been mined. At present, the lode is open at depth.

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 less than 0.1 m up to 2.4 m, with an average width of 0.4 m where it has been mined.

The north-west 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 lode dips at about 85° to the east and occurs over a strike length of approximately 750 m, with a down-dip extent of approximately 250 m. It has an average true thickness of approximately 0.53 m. The Cuffley lode is situated under the flat (30° to 40°) west dipping fault, which has a reverse movement of 30–50 metres (Figure 8-6). At depth the lode turns from more vertical to a steep easterly dip and cross-cuts shallow west-dipping to flat bedding in the Costerfield formation, on an open parasitic, anticlinal fold ramping up to the crest of the Costerfield Dome to the east.

Within the northern extremity of the Cuffley lode, a large splay (Cuffley East/CE) exists diverging off the main lode (Cuffley Main/CM) northwards of the East Fault around the 5,000 mN northing (mine-grid northing). The plunge of this CE-CM intersection dips approximately 50–60˚, with an approximate north strike.

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 less than 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 50 m to the west of the northern section of E lode. The lode dips at approximately 55˚ to the west with a strike length of 80 m and a vertical extent of 70 m. It consists of bedded parallel, laminated quartz-stibnite veins of 0.1 m to 0.35 m width.

B Lode it interpreted to be a splay of the northern extent of E Lode and is located between 0–20 m to the west of E Lode. It is sub vertical to west dipping and strikes north-east. Its vertical extent is approximately 170 m and strike extent 100 m.

The Brunswick Lode lies approximately 600 m west and 1,300 m north of E Lode. The lode is sub-vertical and occurs over a strike length of approximately 350 m, with a down-dip extent of roughly 200 m. It has an average true thickness of about 1.28 m. At present, the lode is open at depth.

Figure 8-4: Cuffley deposit looking North – 895L C2 North lode Note: For scale, the end of rock bolt is 300 mm x 300 mm.



Figure 8-5: Deposit wireframes and underground development



Figure 8-6: Structural relationship between the Costerfield Thrust and the Augusta and Cuffley Lode systems



Figure 8-7: Costerfield geological systems showing Brunswick, Bombay/Minerva, Cuffley and Augusta deposits

North



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.

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; seven were waste.

The following petrophysical measurements were made:

• Mass properties:

− Dry bulk density;

− Apparent porosity;

− Grain density; and

− Wet bulk density.

• Inductive properties:

− Magnetic susceptibility;

− Diamagnetic susceptibility; and

− Electromagnetic conductivity.

• Galvanic properties:

− Galvanic resistivity; and

− 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

9.3.1 Ground Geophysics Based on the results of the petrophysical testing programme, a limited programme 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 at detecting the known mineralisation at Augusta.



9.3.2 Airborne Geophysics A low-level detailed airborne magnetic and radiometric survey was undertaken in 2008 by AGD 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 assist in interpreting the local buried structures.

9.4 Geochemistry

9.4.1 Mobile Metal Ion (MMI) 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, five-metre 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 appear to show a broad anomaly across the mineralisation, indicating that the technique could be useful for regional exploration.

9.4.2 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 to 1970, when a grid south of the 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 (22–57 m) diamond drillholes 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. Drillholes 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 latter part of the programme, 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 drillhole and laid out in sequence on the ground. Mandalay staff supervising the programme 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 has been 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, its likelihood is reduced because the sample material is 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 programme.

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, M032 drillhole intersected E Lode (55% antimony, 59 g/t gold) 35 m below surface and MH041 drillhole intersected E Lode up-dip (35.3% antimony, 36.8 g/t gold) at 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 a low-grade lode exists at shallow depth or a high-grade lode exists at depth.

Figure 9-1: Augusta South Aircore Programme Results with Lode Locations Source: Mandalay, 2012.

N



Augusta South Auger Bedrock Geochemistry Drilling (MIN 4644) Mandalay engaged Statewide Drilling to undertake the Augusta South auger geochemistry programme with an Edson 200 auger rig. The three traverse programme (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 programme 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 programme determined when bedrock was intersected by 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, with the coordinates labelled on it. 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 the assigned sample numbers and dispatched to Onsite Laboratories (Onsite) in Bendigo. The use of certified standards was routine practice throughout the sample assay programme. 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 is considered to be a satisfactory method because the results are only used internally within Mandalay as an indication of anomalism, and they are not used as input to resource estimation.

Two parallel NNW trending anomalous zones have been defined by this survey, as shown by 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 Programme From December 2011, Mandalay engaged Starwest Pty Ltd to undertake the Augusta East Auger drilling programme. This programme 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 7295.6 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 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 by 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.

N



A total of 1,375 auger holes were then drilled by Mandalay from 15 April to June 2014 for 3,906m. Holes were drilled over exploration licenses EL 3310 and EL 5432 and mining license MIN 4644 covering 6 prospect areas, Augusta, Cuffley, Brunswick, North and West Costerfield and Margaret’s Reefs. Results from duplicate samples taken during the auger programme show excellent correlation and therefore ensure good sampling practices were maintained throughout the programme. The tables below summarise all auger drilling totals by prospect location and tenement.

Table 9-1: Auger drillhole totals by prospect

Location Location Description Hole IDs No of Holes

Av Depth

Metres Drilled

Lease

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 2.8 690.4 MIN 4644, EL 3310

MAR Margaret’s Reef MAA0001-0542 536 3 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 2.3 772.3 MIN 4644, EL 3310

TOTAL 1,375 2.85 3,906.1

Prospect Location Summary is as follows:

Cuffley 76 holes were drilled on 2 lines over the underground Cuffley deposit to test the relationship between bedrock geochemistry and known gold-antimony ore bodies below surface. The Cuffley ore body 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 ore body at depth.

Augusta Mine Extension To the east, west and south of the existing Augusta mine site, 124 holes were drilled to explore for extensions of the known underground ore bodies. 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 2 lines drilled to the south displayed a narrow zone of high anomalism which correlates directly to extensions of known ore bodies. Diamond drilling between this area and the mine intersected no economic mineralisation and therefore this area represents a low priority drilling target.

Brunswick To the west and south of the Brunswick open cut 247 holes were drilled to test for extension of the known ore body. 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 ore body. Previous diamond drilling south of Brunswick was unable to locate an economic extension to the underground resource. These results can now be re-analysed in conjunction with the bedrock geochemistry “hit” to design further drilling.

Soil sampling is planned to cover the area further south to test the extent of the anomalism and if it connects with moderate results intersected on Newton’s Lane 500 m further south.



Margaret’s Reef The bulk of this year’s drilling programme was conducted over Margaret’s Reef 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 programme. Sample spacing of 10 m over the previous anomalous results gave a clearer indication of vein structures at depth.

Margaret’s appears to be made up of several reef vein systems as suggested from previous reverse-cycle (RC) and diamond drilling (DD). The veins strike approximately north-west which is a similar vein orientations to those 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 from current mining operations, was received from the northern most line at Margaret’s with 5.42 g/t gold and 3.25% antimony, suggesting economic mineralisation at surface.

Several high priority diamond drill targets have been planned, including a target beneath the above mentioned 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 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 further 336 holes were drilled at West Costerfield to test near historic workings to the east and continuity to the south of the previous auger programme. The previous geochemistry programme 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 and 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 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 is interpreted gold-antimony veins below surface similar in style to those intersected in a single diamond drill hole over True Blue.

Several diamond holes have been planned to test the high anomalism in this area which is considered a high grade target due to its proximity to the Brunswick mill site only 1.5 km to the east. There is ready access to drill the property as it covers a single owner where an exploration agreement is currently in place.

Recommendations and further work Soil sampling programmes will be continued in light of the close relationship to bedrock geochemistry where auger drilling access may be difficult over small landholdings and over forested rough terrain.

A comprehensive study of the auger geochemistry results has commenced with the primary objective to increase understanding of the mineralisation and alteration styles of the reef systems found over the exploration tenements.



9.5 Aerial Photogrammetry Survey 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 as well as the Augusta Mine Site in 2005.

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 extended 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: Drillhole Summary

Year Metres (Diamond)

Metres (Percussion/Auger)

2009–2010 458.9 547.0

2010–2011 10,622.7 732.0

2011–2012 18,581.4 7,295.6

2012–2013 24,329.08 3,838.0

2013–2014 20,817.0 3,906.0

Total 74,809.8 16,318.6

10.2 2009/2010 Drilling through 2009 and 2010 mainly comprised 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.

Augusta drilling during 2009 and 2010 concentrated on the definition of the W Lode Resource. Four drillholes 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 through 2010 and 2011 was undertaken on two projects – 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.

Augusta drilling during 2010 and 2011 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 and 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 Exploration from July 2011 to December 2012 was undertaken on four projects – 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 to endeavour 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 as part of 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 a 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 W and N Main Lode. 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 programme began in July 2011 (AD series of holes), following the MB007 discovery. As a follow-up programme, four holes were drilled (AD001-ADD004). 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 has involved drilling at least two holes on each mine grid cross-section on an approximate spacing of 80 m to 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 In 2012/2013 Mandalay Resources drilled 24,329.0 metres of diamond drilling, targeting the Cuffley Lode from surface (Table 10-2). These focussed 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 drillhole. MH335 and MH336 were drilled as wedges and therefore, they have significantly lower collar elevations and shallower dips than the other drillholes. Due to the narrow high-grade nature of the mineralisation, it is not meaningful to report significantly higher-grade intercepts within lower grade intercepts, and therefore they are not reported here.



Table 10-2: Significant Cuffley Lode drillhole intercepts: 2012/2013

Hole ID Collar

Easting (Mine Grid)

(m)

Collar Northing

(Mine Grid) (m)

Collar RL (Mine Grid)

(m) True Width

(m) Gold Grade

(g/t) Antimony

Grade %

AD030 15284.6 4970 1182 0.37 39.7 25.6

AD031 15284.6 4970 1182 0.3 39.1 34.6

AD032 15284.6 4970 1182 1.02 106.2 17.4

AD039 15348 4973 1182.5 0.11 32.8 33

AD040 15306 4916 1183 0.51 7 1.9

MH335 15445.8 4574.9 1020.2 0.18 2.4 0

MH336 15445.8 4574.9 1019.8 0.99 0.8 2.5

AD051 15170.7 4792.7 919.6 0.82 0.89 0.07

AD053 15165.5 4774.5 960.5 0.4 122.8 11.5

AD052 15162.1 4740.1 962.7 0.92 4.41 2.42

AD055 15162 4732.9 925.4 1.09 54.1 9.9

AD049 15193.8 5050.8 1022.6 0.14 178.9 30.4

AD062 15161.8 4762.6 981.8 0.1 18.7 0.5

AD063 15158.3 4748.8 947.9 0.09 4.9 2

AD064 15161 4797.8 945.7 0.33 14.5 13

AD046 15188.9 5099.2 907.9 0.15 16 22.6

AD057 15168.4 4745.4 884.4 3.16 1.63 1.29

AD058 15180.1 4779 884 1.05 3.2 0

AD056 15177.3 4839.9 919.7 0.31 29.2 4

AD048 15199.2 5089.3 1011.6 0.38 3.3 0

AD059 15183.3 4862.3 878.6 0.38 87.6 35.8

AD060 15180.3 4743.5 855 0.85 2 0

AD061 15188.4 4827.8 840.7 0.87 4.28 6.32

AD047 15192.9 5118.1 947.8 0.8 12.22 3.76

AD047W2 15190 5120 950 0.05 0.25 1.01

AD047W2 15190 5120 950 1.94 1.8 0.15

AD047W2 15190 5120 950 1.11 3.91 0.37

AD050 15194.8 5150.7 976 0.74 2.84 2.51

AD050 15192.6 5151.1 973 1.31 5.44 4

AD066 15195 4825.8 824.2 0.3 46.8 7.44

AD067 15187.4 4909.2 888.6 0 0 0

AD068 15183 4906.3 849 0.14 6.29 0.19

AD069 15189.5 4891.8 799.8 0.08 5.94 0

MH364 4890.7 15351.6 929.8 0.09 1 24.4

MH364 4892.7 15356.3 928 0.42 2.6 2.1

MH365 4894 15337.2 905 0.12 1.5 4.9

MH365 4904.6 15354.7 887.9 2.32 0.6 1.5

MH367 4862.3 15352.7 859.3 2.41 1.4 1.8

MH367 4862.6 15346.1 868.6 6.96 12.6 5.4



10.6 2013/2014 Cuffley /N Lode Drilling 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; and

• Infill and extension of 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 metres of diamond drilling was undertaken and 3,906 metres of auger drilling was undertaken on Mandalay Resources Costerfield Operations Pty Ltd tenements at Costerfield during 2013/2014 (Table 10-3). A total of 5,735 metres was drilled for the purposes of target testing, 9,390 metres for resource expansion and conversions and 5,692 metres for resource infill drilling. All drilling activity was conducted by Starwest Pty Ltd using 2 Boart Longyear LM90s, 1 Boart Longyear LM75, 1 pneumatic Kempe U2 and a modified Gemco 210B Track-mounted Auger.

Table 10-3: Significant Cuffley Lode drillhole intercepts: 2014

Hole ID Intercept Easting (Mine Grid)

Intercept Northing

(Mine Grid)

Intercept RL

True Width*

(m)

Gold Grade (g/t)

Antimony Grade (%)

AuEq (g/t) over

1.8m *

Total Hole

Depth (m)

AD070 15147.3 4696.4 916.2 0.12 8.4 7.92 1.7 183.5

AD071 15163.7 4696.8 850.3 0.15 43.10 0.01 3.6 260.8

AD072 15139.9 4628.0 930.5 0.8 4.0 0.9 2.7 198.4

AD073A 15145.0 4685.6 962.1 0.46 7.7 3.0 3.5 390.0

AD074 15153.5 4658.8 852.2 0.36 44.6 62.7 35.0 273.4

AD075 15142.6 4666.0 920.0 0.14 10.8 1.3 1.0 401.5

AD076 15137.2 4613.7 874.6 2.19 19.7 7.3 N/A 257./8

AD086A 15161.2 4630.2 990.9 1.02 6.8 15.9 22.5 194.9

AD088 15170.9 4928.2 776.1 0.10 14.1 1.80 1.0 300.6

AD091 15203.3 4995.5 838.8 0.29 25.1 9.9 7.3 242.3

AD092A 15197.4 4991.1 879.4 0.63 7.0 43.9 34.7 233.6

AD093 15179.8 5088.5 850.5 6.70 31.4 25.5 N/A 250.4

Note * Mineralised intercepts stated in this table may not be the same as the intercepts composited and used within the Resource estimation.

10.7 Drilling Methods Due to the extensive historic drilling conducted throughout the history of the Costerfield area, and 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 and below the 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 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 drillholes 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 drillholes is 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 to 100 m from the projected lode intercept.

Prior to 2011, various sized drillholes and methods were used during drilling. 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 is 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 drillhole collar locations for both Augusta and Cuffley.



Figure 10-1: Known drillhole collar locations for Augusta and Cuffley



10.8 Collar Surveys Since 2006, drillhole 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 of both underground and surface collar locations.

Presently, initial collar locations are sighted using a hand-held GPS, with drilling azimuths provided by compass. Holes are then surveyed on completion.

Between the late 1990s and 2001, most drillholes appear to have been located using a GPS. Drillhole collar locations prior to the 1990s were usually sighted via tape and compass. Where possible, historic drillholes were surveyed in 2005 by Adrian Cummins and Associates, but this was not always possible.

Collars surveyed after 2001 are recorded in the drillhole database as being surveyed. Unsurveyed/unknown drillholes are recorded as either GPS or unknown and are given an accuracy of within 1 m.

10.9 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 those cases, the survey is completed when suitable ground conditions are subsequently encountered.

Between 2001 and 2011, all drillholes appear to have been surveyed via 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.

10.10 Logging Procedures The following information only relates to drilling completed after 1 January 2010 and below the 1,000 mRL in the Augusta deposit.

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 through the use of 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 Microsoft Excel. RQD data is collected corresponding to drill runs and includes the “from” depth, “to” depth, run length in metres, the recovered length in metres, the recovery as a percentage, the length of recovered core greater than or equal to 10 cm, and the number of fractures. From this



data, an RQD value is calculated. This data is transferred daily from the tablet device to the company server and it is then loaded into the primary drilling data Microsoft Access database on a weekly basis.

Once depth marks are placed on the core, site geologists then log lithology, sample intervals, structural data, and geotechnical data (if applicable) directly to a Microsoft Access database using DrillKing logging software. All measurements of structural features, such as bedding, cleavage, faults, and shears are made using a kenometer 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 Microsoft Access database and are also scribed onto the core using wax pencil. Local Microsoft Access databases on the laptops are archived daily to the company server. They are then loaded into the primary drilling data Microsoft Access database weekly.

After geological logging has been completed, all trays are photographed before sampling. Once sampling is completed, the trays are put on pallets and moved to a permanent core storage area.

10.11 Drilling Pattern and Quality

10.11.1 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 the 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 W Lode is the primary target of the majority of this drilling, the pattern and density achieved on E and N Main Lodes can vary greatly. Where increased geological confidence is required, infill holes specifically targeting NE or E Lodes have been drilled at a nominal 40 metres. Surface drilling, targeting depth extensions of the Augusta Deposit, is generally conducted on 100 m sections along strike, with intersections spaced 80 to 100 m up- and down-dip.

10.11.2 Cuffley Initial drilling of the Cuffley Lode was intended to be done in a ‘W’ pattern on an approximate 50 m by 50 m offset grid. This pattern was started with AD001 through to and including AD004. To aid interpretation, this pattern was changed to a 100 m grid based on mine grid northings, with 50 to 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 maximize 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.

10.12 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 drillholes that are snapped to the drillhole 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.

10.13 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, potential for significant impact on results is minimised.

Sampling is also of a consistent and repeatable nature, with appropriate QA/QC methodologies employed. The assay method used is also considered to be appropriate for this style of mineralisation.

The nature of geological information collected during the logging process is considered to be 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.

1 Survey techniques are consistent with the Canadian Institute of Mining and Metallurgy and Petroleum (CIM) Exploration Best Practice Guidelines.




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.

11.1.1 Diamond Core Sampling Much of the drill core produced from the Costerfield area is composed of barren siltstone. Therefore, not all diamond drill core is required to be sampled. Sample intervals are determined and marked on the core by Mandalay geologists.

General rules that are applied in the selection of sample intervals are as follows:

• All stibnite-bearing veins are sampled;

• A waste sample is taken either side of the mineralised vein (ranging 50–100 cms);

• Areas of stockwork veining are sampled;

• Laminated quartz veins are sampled;

• Massive quartz veins are sampled;

• Siltstone is sampled where disseminated arsenopyrite is prevalent; and

• Puggy fault zones are sampled at the discretion of the geologist.

Mandalay staff sample the core. In order to obtain a consistent sample, the diamond drill core is always 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 2.7 m in length. The average sample length for drill core samples within the Augusta drill programme is 31 cm. Drillholes 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 requirement.

RC drilling-based samples 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 drillholes.

11.1.2 Underground Face Sampling Approximately 70% of all drive faces are sampled. Each development cut is approximately 1.8 m along strike. Samples are taken at a frequency of between 1.8 to 7 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;

• Sample locations are marked out so that the sample is taken in a direction that is perpendicular to the dip of the lode;

• The face and lengths of the sample are measured;

• The face is photographed and labelled with heading and northing information;



• 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 are also taken of the intervening siltstone;

• 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 used. A tab from the sample book is placed into the sample bag for use by the assay lab;

• The face is sketched on a face sample sheet and sample details recorded;

• The location of the face is derived from survey pickups of the floor and backs of the ore drive; and

• 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 drillhole intercepts is approximately 40 m by 30 m. This is reduced to 20 m by 20 m in areas of structural complexity. Face sampling along drives is done at a frequency of between 1.8 and 7 m along strike and 5 m to 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) and ALS (Brisbane) have also been used to verify the accuracy of Onsite.

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, which 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 previously marked with a sample number;

• The sample characteristics are marked on a sample ticket and placed in the bag;

• Calico bags are loaded in to plastic bags so that the plastic bags weigh less than 10 kg;

• An assay request sheet is completed and placed in the sample bag; and

• 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 as well as sample bags 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 > than 3 kg the sample is split and 50% of the sample utilised;

• The entire sample is then milled and pulverised to 90% passing 75 μm; and

• Samples are then 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. The following procedure is undertaken by Onsite for gold and antimony.

Gold grades are determined by fire assay/atomic absorption spectroscopy (AAS).

The following procedure is undertaken:

• 50 g of pulp is fused with 180 g of flux (silver);

• Slag is removed from the lead button and cupellation is used to produce a gold/silver prill;

• 0.6 mL of 50% nitric acid is added to a test tube containing prill, and the test tube is placed in a boiling water bath (100°C) until fumes cease and silver appears to be completely dissolved;

• 1.4 mL of hydrochloric acid (HCl) is added;

• On complete dissolution of gold, 8 mL of water is added once the solution is cooled; and

• Once the solids have settled, the gold content is determined by 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; and

• 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, the last visits being on September 2012, February 2013, October 2013, March 2014 and August 2014. Tours of the laboratory are normally completed in the presence of Onsite’s Laboratory Manager, Mr Rob Robinson.

A problem with illegible pulp labels was addressed by rectifying a problem with the labelling machine.



Notes and minutes from laboratory visits and meeting with laboratory staff are preserved on the Mandalay server.

Mandalay conducted a check assay programmes January 2013 and August 2014. This process involves obtaining a pulped sample from Onsite, splitting it, and submitting pulps to Onsite and to one or both of ALS or Genalysis. The results of the most recent programme are outlined in Section 11.7.4.

11.7 Assay Quality Assurance and Quality Control

11.7.1 Standard Reference Material Five standards that are currently in use have been made from material collected underground at Augusta (AGD08-01, MR11-01, MR 0.2, MR04-06 and MR30-20), and are routinely submitted to Onsite. Mandalay also routinely sends three commercially available, gold-only standards or Certified Reference Material (CRM), G901-8, G906-8 and G907-6 (sourced from Geostats Pty Ltd) to Onsite. A standard is sent with each batch of exploration samples (on average 1 standard per 15 samples) and with each batch of the underground face samples (on average 1 standard per 10 samples).

Results from January 2014 to December 2014 for Gold and Antimony are displayed in Figure 11-1 to Figure 11-9. 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 standard G901-8 (CRM Assay 47.24 g/t gold) displays good compliance for the period, with one assay outside the ±3SD limits. There is no significant bias or drift displayed in the assay results for the period (Figure 11-1);

• Gold standard G906-8 (CRM Assay 7.26 g/t gold) displays good compliance for the period, with no assays outside the ±3SD limits. There is a slight low bias for the period. There is no evidence of any significant drift in the accuracy of assay results for the period (Figure 11-2);

• Gold (and Antimony) standard MR11-01 (CRM Assay 115.7 g/t gold) displays exceptional compliance for the period given the very high grade of the standard with assays well within the ±3SD limits. There is a slight low bias showed for this period. There is no evidence of any significant drift in the accuracy in the assay results for the period (Figure 11-3);

• Gold (and Antimony) standard MR04/06 (CRM Assay 7.51 g/t gold) shows good compliance for the period, with no assays outside the ±3SD limits. There is a low bias displayed for the period, however this is a small dataset and this may not be significant. There is no evidence of any significant drift in the accuracy in the assay results for the period (Figure 11-4);

• Gold (and Antimony) standard MR30/20 (CRM Assay 23.64 g/t gold) displays good compliance for the period, with no assays outside the ±3SD limits. There is no evidence of bias or any significant drift in the accuracy in the assay results for the period (Figure 11-5);

• Antimony standard AGD08-01 (CRM Assay 7.08% antimony) shows good compliance for the period, with no assays outside the ±3SD limits. There is no evidence of bias or any significant drift in the accuracy in the assay results for the period (Figure 11-6);

• Antimony standard MR11-01 (CRM Assay 54.55% antimony) displays moderate compliance for the period, with 36 exceptions from 391 assays. This discrepancy has been investigated and it was found that this material was certified using much higher precision techniques (ICP-MS and



Fusion XRF) than those used by Onsite Laboratories for grade control assays, the large assay population now obtained from Onsite shows the same average result (54.55% antimony) with a larger standard deviation. When this is taken into account only one result falls outside of the ±3SD limits. There is no significant evidence for bias or drift in accuracy of the assay results, there was one cluster of high biased results in august but this is an isolated occurrence and is not considered material to the resource (Figure 11-7);

• Antimony standard MR04/06 (CRM Assay 3.50% antimony) shows good compliance for the period, with no assays outside the ±3SD limits. There is no evidence of bias or any significant drift in the accuracy in the assay results for the period (Figure 11-8); and

• Antimony standard MR30/20 (CRM Assay 31.08% antimony) shows good compliance for the period, with no assays outside the ±3SD limits. There is no evidence of bias or any significant drift in the accuracy in the assay results for the period (Figure 11-9).

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: G901-8 Gold Standard Reference Material – Assay Results for 2014


Figure 11-2: G906-8 Gold Standard Reference Material – Assay Results for 2014

Figure 11-3: MR11-01 Gold Standard Reference Material – Assay Results for 2014

0

2

4

6

8

10Au Standard G906-8

Assay Result -3 Std. Dev. +3 Std. Dev.

0

20

40

60

80

100

120

140Au Standard MR11-01




Figure 11-4: MR04/06 Gold Standard Reference Material – Assay Results for 2014

Figure 11-5: MR30/20 Gold Standard Reference Material – Assay Results for 2014

0

1

2

3

4

5

6

7

8

9

Au Standard MR04/06


0

5

10

15

20

25

30

Au Standard MR30/20




Figure 11-6: AGD08-01 Antimony Standard Reference Material – Assay Results for 2014

Figure 11-7: MR11-01 Antimony Standard Reference Material – Assay Results for 2014

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Figure 11-8: MR04/06 Antimony Standard Reference Material – Assay Results for 2014

Figure 11-9: MR30/20 Antimony Standard Reference Material – Assay Results for 2014

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11.7.2 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. In the case of gold at Onsite, this is >0.02 g/t.

For the period between Jan 2012 and Dec 2012, 73 out 149 samples (49%) of blank material exceeded the 0.02 g/t gold threshold. This result was deemed unacceptable and Mandalay worked with Onsite to resolve the issue.

The cause of the unacceptable assays was largely due to the use of Blank samples as “cleaners” whereby they were submitted immediately after a known high grade sample to remove any sources of contamination from the assay process for the next sample. It became apparent that the quartz blank wash used by the lab to remove contamination from between samples was not cleaning the equipment thoroughly enough.

Although the assay procedures should still provide an assay for blank material at acceptable levels approaching the detection limit, this was not the case. This approach has been changed so that the Blank standards are submitted randomly instead of after a known high grade sample. The 2013 Blank assay results were much improved with the maximum assay at 0.09 g/t gold, but there were still 9 of 37 samples (24%) that are above the acceptable limit of 0.02 g/t.

The 2014 Blank assay results show similar performance to 2013 with 92 of 345 results (26%) that are above the acceptable limit of 0.02 g/t. As above, the unacceptable assays were mostly caused by blanks following high grade assays to remove contamination. These results occurred during the first part of the year and combination of enforcing the procedure of random blank insertion and closer ongoing co-operation with Onsite appears to be remedying the situation as illustrated by Figure 11-10 below.

Given the very Gold high 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.

Figure 11-10: Blank assay results for 2014

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11.7.3 Duplicate Assay Statistics A summary of laboratory duplicate statistics assayed by Onsite for original gold assays versus duplicate gold assays for 2014s 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 11-11 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 11-12. 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 78% of pairs at less than 10% RPD, which demonstrates adequate but less than desirable precision in the gold assays by Onsite.

Table 11-1: Summary of Onsite duplicate statistics

Au Original Duplicate

Number of samples 213 213 Mean 21.50 20.73 Maximum 545.80 411.30 Minimum 0.00 0.00 Population SD 52.20 46.04 CV 2.43 2.22 Bias 3.57%

Correlation Coefficient 0.98

Percent of samples < 10% RPD 77.93

Figure 11-11: Scatter Plot for Onsite Gold Duplicates (g/t)

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Figure 11-12: Relative Paired Difference Plot for Onsite Gold Duplicates (g/t)

11.7.4 Check Assay Programme – Sample Pulps Duplicate statistics resulting from a gold and antimony check assay programme comparing Onsite and ALS Minerals (ALS) are presented in Table 11-2 and Table 11-3 respectively. The duplicates are assayed on separate aliquots of the same sample pulp. The check assay programme was undertaken by Mandalay in August 2014.

The gold scatter plot of this data (Figure 11-13) shows some bias (13%) between Onsite and ALS, with ALS returning higher grades than Onsite. The gold scatterplot shows consistent trend toward higher ALS grades at all levels. The antimony statistics show some bias (6%), with Onsite returning higher grades than ALS on average (Figure 11-14). The antimony scatterplot however does not show any consistent bias. The gold RPD plot shows that 66% of duplicate pairs show less than 20% RPD (Figure 11-15). This demonstrates poor reproducibility across laboratories. The antimony RPD plot shows that approximately 68% of duplicate pairs are less than 20% RPD (Figure 11-16), which demonstrates poor reproducibility across laboratories.

The reproducibility between laboratories of gold grades is similar to that which was found in earlier check assay programmes from 2013 and 2012. In contrast, the reproducibility of antimony assays across laboratories decreased markedly from 92% RPD to 68% RPD. This reason for this difference is currently under investigation with a larger sample population being prepared for a three way check assay programme in early 2015. Preliminary investigations suggest that the apparent overcall of antimony grades by Onsite may be more reliable as their acid digest method has been optimised for high grade antimony through the addition of tartaric acid to prevent antimony precipitation out of the solution.

The higher gold grades returned by ALS may not be representative, investigation of this issue has revealed that when ALS labs have been used in certification programmes for CRM’s they have a

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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 to be 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 versus ALS Gold Duplicate Statistics

Au Onsite ALS

Number of samples 42 42

Mean 50.74 57.60

Maximum 151.60 188.00

Minimum 2.11 2.03

Population Standard Deviation. 39.76 47.31

Coefficient of Variation 0.78 0.82

Bias -13.53%

Correction Coefficient 0.97


Table 11-3: Summary of Onsite versus ALS Antimony Duplicate Statistics

Sb Onsite ALS

Number of samples 35 35

Mean 31.68 29.78

Maximum 59.40 64.00

Minimum 1.28 0.78

Population Standard Deviation 17.52 19.04

Coefficient of Variation 0.55 0.64

Bias 5.99%

Correction Coefficient 0.96




Figure 11-13: Scatter Plot for Onsite versus ALS Gold Duplicates (g/t)

Figure 11-14: Scatter Plot for Onsite versus ALS Antimony Duplicates (%)

y = 1.1436x

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Figure 11-15: Relative Paired Difference Plot for Onsite versus ALS Gold Duplicates (g/t)

Figure 11-16: Relative Paired Difference Plot for Onsite versus ALS Antimony Duplicates (%)

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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, wrapped in plastic.

11.9 Conclusions SRK makes the following observations and recommendations:

• The sample preparation and analytical procedures undertaken by Mandalay are suited to the style of deposit being estimated;

• The level of compliance and bias displayed by the standard reference material samples is acceptable and demonstrates the reliability of the gold and antimony grades used to inform the Mineral Resource Estimate;

• The fact that around 26% of blank samples exceeded the acceptable threshold is of concern and requires further study by Mandalay. SRK does not consider that the high blank values are sufficient to have a material effect on the Mineral Resource Estimate;

• The marked drop in Sb reproducibility across laboratories is of concern to Mandalay. Differing analytical methods employed at each lab may be a contributing factor to this disparity. It is recommended that an additional check assay programme using three laboratories and a larger number of samples be undertaken in 2015;

• It is also recommended that Mandalay institutes a check assay programme on a quarterly basis rather than annually. This recommendation is carried forward from May 2014; and

• In 2014 two laboratory inspections were carried out by Mandalay personnel. It is recommended that this process be undertaken on a quarterly basis with formalised documentation.



12 Data Verification On 18 November 2014, SRK full-time employee Danny Kentwell (QP for Sections 6 to 12 and Section 14) visited the Augusta and the Brunswick Mine sites and was escorted by Chris Davis, Resource Manager for the Costerfield Operations. All drill-core for the Costerfield Property is processed at the Brunswick exploration core shed. The following data verification steps were undertaken:

• Discussions with site geologists regarding:

− 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; and

− 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 (gold bearing quartz – pink, stibnite – yellow) and bedding

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 and 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

Metallurgical testing is undertaken to provide information on the processing characteristics of the samples tested. This allows for plant design and the estimation of costs and recoveries expected through a full scale processing facility. In the case of the Costerfield Operation, processing of the Augusta ores has been ongoing for 7 years, and more recently a blend of Augusta and Cuffley ores since January 2014. This provides a much better understanding of the processing behaviour expected on these and similar ores. The use of historical operating data is a more accurate way to forecast future processing behaviour when processing similar ores rather than using test work data. In this case it has been used for the future LoM forecasting of throughput and recovery.

The historical Augusta test work and comparison of the results against plant operation is discussed below to demonstrate the confidence provided through the test work in its application to plant throughput and recovery forecasting. This confidence has now also been demonstrated with the Cuffley ore test work and operational data which is also discussed below. It provides confidence in the Cuffley data used for the preliminary economic assessment (PEA).

Mandalay began processing the new Cuffley ores from January 2014 in a blend with the existing Augusta ores at an average ratio of 44% Cuffley and 56% Augusta ore. This feed blend has continued into 2015 and this blend will continue to be processed for the LoM. The Cuffley ore had not been processed before and to allow for an understanding of how it was expected to behave in respect to throughput, recovery, concentrate quality and operating costs, additional test work was undertaken before it was processed. The Cuffley test work benefited from the historical operation because test work recovery behaviour versus plant recovery behaviour has been periodically assessed and has been found to compare well. This provides confidence in the Cuffley ore test work recovery results and the scale-up of recovery to plant scale. This is discussed further below.

13.1.1 Historical Test work – Augusta Ore test work vs actual plant performance Historical metallurgical test work was undertaken on samples from the Augusta deposit between 2004 to May 2012. The test work intent was to assess their physical attributes for crushing and grinding; and to determine metallurgical recovery of antimony (Sb) and Gold (Au). The following reputable and appropriately experienced laboratories were involved with various aspects of the metallurgical evaluation:

• ALS Ammtec - NSW (previously Metcon Laboratories) (Metcon); and

• AMDEL Ltd Mineral Services Laboratory - South Australia (Amdel).

The original test work samples were considered to be representative of the ores to be processed (i.e. they were in the pit design, were predominately drill core and representative of the early/ongoing LoM plan), and testing was extensive with both composite and variability testing done to confirm the likely range of metallurgical behaviours.

Amdel tested Augusta diamond drill core samples in 2004 prior to it being processed through the Brunswick processing plant. Testing included the ore’s amenability to crushing and grinding and the metallurgical response of the different ore types to gravity concentration, flotation and cyanide recovery.



Metcon has been responsible for ongoing ore characterisation test work since 2008. The aim of this test work has been to compare the recovery of the laboratory tested samples to the mill reconciled recovery. In total, ten separate ore characterisation test work campaigns have been completed on Augusta ores. Samples were selected via face sampling of ore drives or from mill feed belt cuts to coincide with what was being processed at the time through the processing plant.

The Metcon ore characterisation tests involve testing specifically targeted grind sizes and the ore amenability at these sizes to gravity concentration, flotation and cyanide recovery. Since July 2011, (i.e. test M2421), gold deportment (diagnostic leach) test work on the flotation tailings has been included in the ore characterisation tests to determine the percentage of cyanide soluble, sulphide locked and silicate locked gold. The three most recent test work campaigns have demonstrated a very close alignment with the laboratory test recoveries versus actual plant performance. It provides confidence that test work can be scaled directly to plant recoveries for the samples tested.

Table 13-1 and Table 13-2 outline the plant performance recovery for Sb and Au separately, compared to the historical test work for the Augusta processed ore.

For the first four ore characterisation tests (M1592, M1609, M1654 and M2091), there were no accurate plant flotation feed sizings to allow for an accurate recovery comparison due to the difference in the grind size of the test work and the plant operation as recovery is grind dependent. This is also why there is a range given for the test work recovery as the metallurgical testing was completed at different size fractions. These earlier comparisons were also during earlier operations were plant operating conditions had not been optimised and explains the different recoveries.

Tests M2115 and M2139 were completed only at a grind size P80 of 106 µm, which is not comparable to the actual plant flotation, feed size range that was between P80 76 µm and P80 60 µm at the time. This has been used to explain the difference between the tested versus actual recoveries although the variation is relatively large.

The most recent and representative tests, M2305, M2421 and M2610 highlighted in the tables below in green, were tested at similar flotation feed size ranges and conditions to the actual plant feeds size and conditions at the time and showed good correlation for antimony recovery at a marginally finer grind. The gold recovery comparison show two tests are consistent but M2421 is not. This was due to an unusually low plant gold grade for that month (July 2011) of just 4.51 g/t Au when normally it is above 8 g/t Au. This explains the difference between it and the test and demonstrates there is a grade recovery gold relationship as would be expected. The comparative testing over the last few years has shown a good arsenic and gold recovery correlation between test work and actual plant performance.



Table 13-1: Test work vs Plant Reconciled Recovery Results for Antimony

Ore Characterisation Sb Results

Laboratory Test work

Recovery%

Reconciled Plant

Recovery% Plant

flotation feed Plant vs Laboratory

Test work

Sb Sb P80 µm % Difference Sb Recovery

2008 January M1592 95–85 77

Data unavailable

2008 April M1609 76–97 83

2008 May M1654 98–99 79

2010 May M2091 97–99 90

2010 June M2115 @106 µm 98 90 76 Flotation feed size difference too great for

direct comparison 2010 July M2139 @106µm 73 88 60

2011 January M2266 @75µm 94 83 60 11

2011 February M2305 @53µm 94 92 47 2

2011 July M2421 @53um 97 96 47 1

2012 May M2610 @53um 97 96 47 1

Table 13-2: Test work vs Plant Reconciled Recovery Results for Gold

Ore Characterisation

Au Results

Laboratory Test work

Recovery%

Reconciled Plant

Recovery% Plant flotation

feed Plant vs Laboratory

Test work

Au Au P80 µm % Difference Au Recovery

2008 January M1592 71–73 78

Data unavailable

2008 April M1609 40–59 81

2008 May M1654 52–87 79

2010 May M2091 78–81 78

2010 June M2115 @106um 86 83 76 Flotation feed size difference too great for

direct comparison 2010 July M2139 @106um 78 83 60

2011 January M2266 @75um 94 82 60 12

2011 February M2305 @53um 89 89 47 0

2011 July M2421 @53um 91 85 47 6

2012 May M2610 @53um 89 90 47 -1

In summary, considering the improvements in plant recoveries from 2011 onwards on the existing Augusta ore, there is good correlation between the plant recovery and test work recovery results for both antimony and gold. This shows that the test work recovery results are a good indicator of the expected plant performance and can be applied to the new Cuffley feed ores.



13.1.2 Historical test work – Cuffley ore vs Augusta ore test work results A variety of metallurgical test work was completed on the Cuffley Lode to determine its metallurgical characteristics. The aim of this test work was to highlight any potential processing issues associated with the Cuffley mineralisation, as well as enable a processing comparison between the Augusta and Cuffley ores. The following reputable and appropriately experienced laboratories were involved with various aspects of the metallurgical evaluation:

• ALS Ammtec, Brookvale, New South Wales (Previously known as Metcon Laboratories) (Metcon); and

• AMDEL Limited Mineral Services Laboratory, South Australia (Amdel).

The Cuffley test work samples were considered to be representative of the ores to be processed. For Cuffley sample #1 approximately 3 metres of ¼ core were sampled from each of 7 drill holes in order to obtain a 35 kilogram representative sample of the Cuffley Lode. The length of sample was calculated using the angle between the lode structure and each individual drill hole with the intent to achieve a sample that would represent a 2.2 m wide development which would represent the head grade of the processing plant.

For Cuffley sample #2 approximately 3 metres of ¼ core was sampled, per hole from 6 drill holes in order to obtain a 30 kilogram sample of the lower grade boundary of Cuffley Lode. The length of sample was calculated using the angle between the lode structure and each individual drill hole with the intent to achieve a sample that would represent a 2.2 m wide development which would represent the head grade of the processing plant for the bounding area.

For Cuffley sample #3, approximately 2 metres of ½ core was sampled from three holes. One metre of core was sampled either side of the Cuffley Met Sample #2.

To determine the potential differences with regards to processing between the Augusta ore and Cuffley mineralisation, three different Cuffley samples and one Augusta orebody sample, labelled ‘2012 May Mill Feed’, were sent to Metcon for ore characterisation test work. A summary of the test work undertaken and the sample grades is shown below in Table 13-3.

Table 13-3: Test work Schedule and Grades

Cuffley Test work Schedule and

Grades Ore

Characterisation Ball Mill Work

Index QEMSCAN Mineralogy

Sb Grade (%)

Au Grade (g/t)

Cuffley #1

7.69 20.27

Cuffley #2

0.55 1.18

Cuffley #3

- -

2012 May Mill Feed 3.55 8.52

13.1.3 Cuffley vs Augusta Lode Mineralisation The Mill Feed May 2012 Sample and Cuffley #1 Sample were sent to Amdel Mineral Laboratories for QEMSCAN analysis, to determine the differences between the two samples and to potentially highlight any deleterious elements that could have a significant effect on the proposed processing methods.

The mineral abundance as shown in Figure 13-1 demonstrates very similar mineralisation characteristics when the difference in sample grades are taken into account (Cuffley #1 Sample grade of 20.27 g/t gold and 7.69% antimony and Mill Feed May 2012 Sample grade of 8.52 g/t Au and 3.55% Sb).



The similarities in the mineralisation and the recoveries achieved from the ore characterisation test work suggested that there is no significant difference in mineralisation that will have an adverse effect on the Cuffley ore’s processing recovery.

Figure 13-1: Amdel QEMSCAN Mineral Abundance

13.1.4 Cuffley vs Augusta Hardness Comparison The Bond Ball Mill Work Index for the Augusta ore was calculated at 15.5 kWh/tonne, compared to the Cuffley #3 Sample which was calculated as 16.0 kWh/tonne. This is a relatively small difference although is marginally higher for the Cuffley ores.

The Metcon ore characterisation test work included laboratory grind determinations which determine the appropriate grind time to get the correct feed size for the test work. This provides additional information on the sample’s physical behaviour and allows for further comparison. Figure 13-2 and Figure 13-3 show the difference in hardness between the two deposits at different grind sizes. From these results, it can be assumed that there is little difference in the hardness of the two deposits and it was not considered to be significantly different enough to warrant concern for plant throughput.



Figure 13-2: Grind Establishment Test work Comparisons at 75 µm

Figure 13-3: Grind Establishment Test work Comparisons at 53 µm

13.1.5 Cuffley vs Augusta Test work Recovery Conclusions The original conclusion was drawn that the Cuffley ores’ test work recoveries could be applied directly to plant performance. From previous analysis of test work versus actual plant performance it was estimated that the recovery on the Cuffley ore would decrease slightly for antimony to 95% (from 96% when processing only Augusta ore) and to increase slightly for gold to 92% (from 89% when processing only Augusta ore). The comparative results are provided in Table 13-4 and Table 13-5 below but since this work was completed, there has been a year of operating data for the Cuffley ores and this is now being used to forecast future throughput and recoveries.

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Table 13-4: Antimony Recovery Comparison Cuffley Lode vs Augusta Orebody


Gravity Recovery Antimony%

Flotation Recovery Antimony%

Total Recovery Antimony%

P80 75 µm

P80 53 µm

P80 38 µm

P80 75 µm

P80 53 µm

P80 38 µm

P80 75 µm

P80 53 µm

P80 38 µm

Cuffley #1 (M2569) 8.0 6.3 4.8 83.9 90.1 91.7 91.9 96.4 96.5

Mill Feed May 2012 (M2610)

7.9 5.8 5.4 87.9 91.3 91.8 95.8 97.1 97.2

Cuffley - Augusta difference 0.1 0.5 -0.6 -4.0 -1.2 -0.1 -3.9 -0.7 -0.7

Table 13-5: Gold Recovery Comparison Cuffley Lode vs Augusta Orebody


Gravity Recovery Gold %

Flotation Recovery Gold %

Total Recovery Gold %

P80 75 µm

P80 53 µm

P80 38 µm

P80 75 µm

P80 53 µm

P80 38 µm

P80 75 µm

P80 53 µm

P80 38 µm

Cuffley #1 (M2569) 55.7 52.0 53.4 40.2 39.8 42.2 95.9 91.8 95.6

Mill Feed May 2012 (M2610)

49.1 46.2 50.7 42.7 43.0 42.0 91.8 89.2 92.7

Cuffley - Augusta difference 6.6 5.8 2.7 -2.5 -3.2 0.2 4.1 2.6 2.9

13.2 Recent test work – Cuffley South In January 2014, additional test work was carried out as the original Cuffley orebody had been extended to the South and at depth. The new test work samples are referred to as Cuffley South. Seven of the original 11 intercepts used for this sample have already been mined and milled. Testing was completed at Australian Minmet Metallurgical Laboratories, in Gosford NSW (AMML). The samples taken for this test work are shown in the long section in Figure 13-4.

The head grade of the samples was deemed representative of the potential Cuffley South ore and assays reported at 9 g/t Au and 3% Sb. The sample provided for metallurgical test work in January 2014 was obtained from HQ size diamond drill core within the southern portion of Cuffley Main. The sample interval for each drill core was calculated using the relationship of the core to the orebody so that the samples were representative of normal ore development and mill feed. The sample interval was centralised on the orebody with both footwall and hanging wall of the lode structure being incorporated to mimic the proposed mining positioning of ore within development. A quarter of the core was taken from each designated hole and sample interval. Once all 11 samples were obtained they were combined to make one composite sample.

In the year since these samples were taken approximately 20% the Cuffley South area represented by the sample has been mined. The ore tonnes extracted from this area represents approximately 10% of all ore mined during 2014. This sample will be representative of approximately 25% of the ore that will be extracted within the proposed LoM schedule.



Test work on the Cuffley orebody showed that the recovery was 95% Sb and 92% Au at the flotation target size P80 53µm and at the average LoM feed grades. In comparison, the Cuffley South test work showed that at the target grind of P80 53µm, recoveries were 97% Sb and 90% Au, i.e. it is similar to the original Cuffley testing and there are no concerns over any significant change in metallurgical behaviour.

This test work programme was not as detailed as the initial Cuffley vs Augusta report but continues to support the high Sb and Au recoveries expected on the Cuffley ores at depth.

The best metric for forecasting future recoveries is the processing data, and as an Augusta/Cuffley mix has been processed since January 2014, this data will continue to be used in forecasting for Cuffley South ores.



Figure 13-4: Cuffley long section with metallurgical sample locations



13.3 Processing Recovery Data on Cuffley vs Augusta orebody Since January 2014 the Cuffley ore has been processed in conjunction with the existing Augusta ore. Approximately 60% of the area represented by Cuffley sample #1 has been mined in 2014 and represents approximately 40% of ore mined during the year. The rest of Cuffley sample #1 will equate to approximately 15% of the ore mined going forward. Cuffley sample #2 was on the periphery of the ore deposit and as such only represents about 2% of ore mined during 2014 but will represent about 10% of mining in the future. Figure 13-5 shows the mill performance before and after the transition to include Cuffley ore in the blend and Figure 13-6 shows the breakdown of the percentage of Cuffley and Augusta mined during 2014. The average for all of 2014 was 44% Augusta ore and 56% Cuffley ore. This does not correlate exactly to milled ratios due to minor stockpile variances. The current LOM plan has a blend of Cuffley/Augusta ore continuing to be mined in parallel at similar feed ratios. There is no further appreciable increase in milling rates forecast as the milling circuit is at capacity.

Figure 13-5: Mill performance on Augusta Ore and Cuffley/Augusta Mix

Figure 13-6: Percentage of Cuffley and Augusta Ore Mined in 2014



The mill performance since January 2014 purely on the Augusta/Cuffley blend is shown in Figure 13-7. Tonnages have remained relatively consistent, averaging 12,445 dry tonnes milled per month for 2014, with a maximum of 13,618 tonnes in August. Forecast throughput for 2015–2017 is 13,000 t/month. Based on 2014 throughput, this LoM milling capacity forecast is supported.

Figure 13-7: Mill recoveries and throughput for Cuffley/Augusta ore blend 2014

The Costerfield LoM and associated financial model is simplistic in that it does not assign a Sb grade recovery relationship. However, forecast grades are equivalent to those historically processed and therefore operating data can be used to establish future Sb recoveries. Forecast Sb recovery from 2015 to 2017 is 94% at an average Sb grade of 3.62%.

Gold grades vary significantly for the LoM and this will affect recoveries. This grade recovery relationship is discussed below. Forecast gold grades for 2015 average 9.43 g/t, so the recovery forecast remains at 90% for the year, based on historical recovery at similar grades. For 2016 the average gold grade is 7.57 g/t so the forecast gold recovery drops to 85.6%. For 2017 the average grade is 7.17 g/t, with a forecast recovery of 85%.

Figure 13-8 shows the historical Sb and Au feed grade and recovery data for 2014. It shows the forecast Sb recovery assumption to be well supported with an average recovery on the Cuffley/August blend in 2014 of 94.5%. There is a Sb grade versus recovery relationship but at 3.62% for LOM it still achieves a recovery of 94% and is considered to be relatively stable across the proposed feed grade range. The forecast Sb recovery is considered to be well supported by historical operating data and similar feed grades.

Operating data from 2013 to 2014 shows the Au recovery assumption to be well supported at similar gold feed grades with an average recovery on the Cuffley/August blend in 2014 of 89.6% however this is at a higher Au feed grade than that forecast in the later LoM. Figure 13-9 shows the gold grade versus recovery relationship based on recovery data for 2014. It demonstrates a very strong correlation and this has been used to forecast the recoveries at a lower LoM gold feed grade in 2016 and 2017. There is a high degree of confidence in the forecast gold recovery in the LoM at the forecast feed grades based on this correlation.



Figure 13-8: 2014 Cuffley/Augusta Blend Feed Grade and Recoveries

Figure 13-9: Gold Grade vs Recovery Relationship on Cuffley/Augusta blend (2014)



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 correctly assigns weights to composites of different lengths during estimation. The interpolation method used was ordinary kriging with the exception of Cuffley Deeps and Alison South with which inverse distance squared was used. The estimated grade is then back-calculated by dividing estimated gold accumulation and estimated antimony accumulation by estimated true thickness.

Not all models have been re-estimated due the absence of new data being captured during the year 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 2014

Lode Zone Code

New data captured

during 2014 New

Estimation New Resource Classification

Depleted during 2014

Reported above new

cut-off

E Lode 10 No No No Yes Yes

B Lode 15 No No No No Yes

W Lode 20 No No No Yes Yes

NM Lode 40 Yes Yes Yes Yes Yes

NE Lode 41 Yes Yes Yes Yes Yes

NSW Lode 45 No No No Yes Yes

P1 Lode 55 Yes Yes Yes Yes Yes

P2 Lode 56 Yes Yes Yes No Yes

K Lode 60 No No No No Yes

CM Lode 210 Yes Yes Yes Yes Yes

CE Lode 211 Yes Yes Yes Yes Yes

AS Lode 230 Yes Yes Yes No Yes

Brunswick 300 No No No No Yes

NW Lode NA Reinterpreted as part of the NM Lode Structure

For sample statistics, top cutting and estimation parameters of models not updated during this estimation last year’s NI 43-101 report can be referenced (Fairfield et al., 2014). From here on the models not re-estimated at 2014 year-end are called “2013 models” (E, B, W, NSW, K Lodes). The Brunswick model will be called Brunswick from here on.



14.2 Diamond Drillhole and Underground Face Sample Statistics Statistics for gold and antimony grades as well as true thickness for lodes newly re-estimated 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. There for 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 invariable 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 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)

164

0.0 319.3 31.7 1.45

Sb (%) 0.0 59.5 14.4 1.13

Vein Width (m) 0.027 2.100 0.246 1.02

Face Sample

Au (g/t)

1259

0.0 989.0 63.0 0.96

Sb (%) 0.0 67.1 31.0 0.59

Vein Width (m) 0.005 2.226 0.283 1.11

NE Lode

Drill Hole

Au (g/t)

52

0.0 135.2 18.4 1.47

Sb (%) 0.0 39.9 7.5 1.25

Vein Width (m) 0.024 1.006 0.210 1.06

Face Sample

Au (g/t)

67

0.0 146.2 45.3 0.67

Sb (%) 0.0 55.2 24.9 0.56

Vein Width (m) 0.010 0.795 0.147 1.12

P1 Lode

Drill Hole

Au (g/t)

12

0.3 162.8 39.8 1.44

Sb (%) 0.3 19.9 10.6 0.64

Vein Width (m) 0.048 0.286 0.142 0.53

Face Sample

Au (g/t)

127

0.2 1356.0 75.9 2.40

Sb (%) 0.1 54.7 19.3 0.70

Vein Width (m) 0.020 0.950 0.189 1.07

P2 Lode Drill Hole

Au (g/t)

17

0.2 145.3 31.0 1.54

Sb (%) 0.0 39.8 8.5 1.57

Vein Width (m) 0.054 0.670 0.208 0.84

CM Lode

Drill Hole

Au (g/t)

144

0.0 189.0 24.2 1.51

Sb (%) 0.0 62.7 10.8 1.33

Vein Width (m) 0.039 1.908 0.377 0.88

Face Sample

Au (g/t)

481

0.0 1136.0 67.9 1.37

Sb (%) 0.0 64.1 24.4 0.69

Vein Width (m) 0.005 3.010 0.417 0.98

CE Lode

Drill Hole

Au (g/t)

24

0.7 148.4 50.9 0.91

Sb (%) 0.7 55.4 22.6 0.64

Vein Width (m) 0.046 0.924 0.203 0.94

Face Sample

Au (g/t)

135

0.1 821.9 121.0 1.10

Sb (%) 0.0 65.2 32.5 0.58

Vein Width (m) 0.002 1.680 0.305 1.11

CD Lode Drill Hole

Au (g/t)

20

0.1 101.4 16.9 1.50

Sb (%) 0.0 50.0 5.4 2.26

Vein Width (m) 0.035 0.868 0.285 0.78

AS Lode Drill Hole

Au (g/t)

11

0.4 41.9 9.0 1.42

Sb (%) 0.0 24.4 6.3 1.25

Vein Width (m) 0.047 0.743 0.331 0.66



In order to show an example of the two factors described above a study has been undertaken on N Lode as it has the largest amount of data. To negate the influence of low grade intercepts on the drilling data all data was cut to the boundaries of the faces sampling data (Figure 14-1).

Figure 14-1: Long section of data collected for the sample comparison analysis showing Face sample (Red) and Drillhole (Blue) Data

This subset of N Lode data shows a higher average grade and a smaller vein width within the face sample data (Table 14-3). 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 compared to 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 with a numeric zone code applied to each (Table 14-1). The identified intervals within both the drill data and the 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 where considered during the process of wireframe modelling.

Sub-domaining of NM and CM Lodes was required to separate high grade and low grade populations to an acceptable degree, thus reducing the degree of heterogeneity. In the process of sub-domaining structural controls on mineralisation were considered and where the break in grade continuity could be established as being connected to a defined structure the intercept trace of that structure was used for the boundaries (Figure 14-2).

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

14.4 Grade Capping Statistical analysis was undertaken on the accumulated data to test for capping requirements for both gold accumulation and antimony accumulation. Each subdomain was analysed separately for capping requirements and capping was applied accordingly. This analysis was carried out with reference to the ongoing monthly model reconciliations to mine and plant production. These reconciliation tests have demonstrated strong concordance between modeled antimony grade and produced grade (Figure 14-37). No capping requirement was found to be needed for antimony accumulation. Statistics before and after grade capping are presented in Table 14-4.

For 2013 models statistical analysis of gold grades resulted in a gold grade cap (or top cut) of 150 g/t before accumulation. No grade cap was applied to the 2013 antimony data before accumulation. After accumulation limited top cutting was applied to some low grade subdomains (SRK 2014).

Grade caps were not applied in the 2009 resource estimate for Brunswick (Fredericksen, 2009).

Table 14-4: Accumulated Gold statistics before and after top cuts

Variable Lode Domain Top No. No. Cut Variable Min Max Mean CV



Cut Samples Samples

Au Acc NM 1 100 182 2 Uncut 0.00 197.8 6.8 2.7

Cut 0.00 100.0 6.1 2.0

Au Acc NM 2 100 139 1 Uncut 0.02 171.8 16.4 1.1

Cut 0.02 100.0 15.9 0.9

Au Acc NM 3 100 310 1 Uncut 0.00 206.6 17.8 1.1

Cut 0.00 100.0 17.4 0.9

Au Acc NM 4 100 527 4 Uncut 0.00 134.9 14.8 1.1

Cut 0.00 100.0 14.7 1.1

Au Acc NM 5 100 264 2 Uncut 0.02 141.3 9.4 1.6

Cut 0.02 100.0 9.2 1.5

Au Acc NE 1 20 119 1 Uncut 0.01 24.1 3.9 1.2

Cut 0.01 20.0 3.9 1.2

Au Acc CE 1 90 159 3 Uncut 0.01 231.7 22.1 1.3

Cut 0.01 90.0 20.3 1.0

Au Acc CM 1 100 166 13 Uncut 0.00 174.5 25.5 1.5

Cut 0.00 100.0 22.6 1.4

Au Acc CM 3 120 425 5 Uncut 0.01 356.9 20.1 1.5

Cut 0.01 120.0 19.2 1.3

Au Acc P1 1 50 139 7 Uncut 0.02 197.8 12.8 2.1

Cut 0.02 50.0 9.5 1.4

14.5 Estimation Domain Boundaries Structural controls on mineralisation have been identified through underground mapping and close structural interpretation of drill core (Figure 14-2). Within Cuffley Main and N Main lodes these relationships have been used to guide estimation domain boundaries.

All estimation domain boundaries are hard boundaries apart from the boundary between domain 2 and 3 within the Cuffley main model where the data from domain 2 is used to influence domain 3. This is to reduce overestimation within southern boundary of domain 3 (Figure 14-3, Figure 14-4).



Figure 14-3: CM Lode Estimation domain boundaries and composite samples

Figure 14-4: NM Lode Estimation domain boundaries and composite samples



14.6 Vein Orientation Domains To estimate true thickness from the drillhole 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 plain to best fit the dip and strike of the domain. The details of this plane was then recorded and added to the drill data within the particular domain.

These dip and strike domains are used to create volume correction factors within the z and y directions by 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, NM, NE, CD and P Lodes in Figure 14-5 to Figure 14-9.

Figure 14-5: CM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor

87 /099

79 /096 79 /271

82 /098

78 /096 74 /106

84 /086



Figure 14-6: NM Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor

Figure 14-7: NE Lode Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor

89 /110

83 /096

80 /280

88 / 071

72 /277

81 /277 85 /282

87 /269

87 /274

79 /281 88 /269



Figure 14-8: Cuffley Deeps Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor

Figure 14-9: P1 and P2 Lodes Dip and Dip Direction (dip/dip direction) Domains Coloured by Volume Correction Factor

81 /273

83 /077

87 /086

84 /251

88 /062

85 /286 85 /065

P1 P2



14.7 Bulk Density Determinations Estimation of bulk density was assessed using two methods as follows:

Bulk density (BD) 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))

Bulk density determinations of drill core samples using a water immersion method were measured. The samples used were whole pieces of diamond drill core, which were not coated in wax.

Figure 14-10 shows the measured bulk density values compared with the values calculated using the bulk density formula above. 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 14-10: 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 composited face sample and drillhole 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 2 dimensions after projecting the data to a constant easting.

Experimental variograms were calculated in the vertical 2D plane on 10 degree increments. The orientation of best continuity in grade accumulation and thickness was selected based on variographic analysis and verified by observations made during underground mapping and of plotted grades on longsections. If the variograms were considered to meaningfully reflect the true spatial continuity in the data, variogram models were fitted to the experimental variograms calculated in the direction of best continuity and 90 degrees to that direction in the 2D plane. Examples of experimental and modelled nugget, primary and secondary variograms, for CM and NM Lodes, are illustrated in Figure 14-11 to Figure 14-19.

In some instances, the direction of best continuity observed in the sample grades did not necessarily agree with the best continuity exhibited by the experimental variogram due to multiple grade trends related to structures associated with mineralisation with different orientations. The sample data for CM Lode, CE Lode, NM Lode, P1 Lode, had sufficient sample pairs to calculate meaningful variograms for grade accumulations and true thickness. The parameters for the fitted models are given in Table 14-5.

The accumulated grades and true thickness for NE Lode and P2 Lode were estimated using variogram parameters from comparable domains because there were insufficient sample pairs for meaningful variography in these lodes.

Although there were estimation subdomains within CM and NM Lodes the overall direction on grade and true thickness continuity was observed to remain constant therefor variographic analysis was conducted using data from combined subdomains within each of these lodes.

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 AS Lode and CD Lode where an inverse distance method was applied. Search ellipsoids were orthogonal to the 2D block model orientation. The following summarises the resource estimation process:

• Drillhole and face samples were projected into the plane of the estimate at 15,300 mE;

• The orientation of the long axis of the search ellipsoid for each lode was guided by the maximum continuity observed in the variography; and

• The anisotropy of the search ellipsoid for each lode was guided by the anisotropy observed in the grade and thickness distribution.

The search parameters for the models estimated at year end 2014 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 14-5.

Estimation was undertaken using a combined dataset of face sample and drillhole data. In the case of NM and CM Lodes grade domains were identified so estimation occurred individually on each domain then the domains were cut to their boundaries and combined to make the entire model. This process resulted in hard boundaries encapsulating these domains. One soft boundary exists within CM Lode where composites from domain 2 were used to inform the model within domain 3.



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 Mine. 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.

Modelled variograms for NM Lode are graphically represented in Figure 14-11 to Figure 14-19.

Figure 14-11: Nugget variogram for gold accumulation in NM Lode



Figure 14-12: Nugget variogram for antimony accumulation in NM Lode

Figure 14-13: Nugget variogram for True Thickness in NM Lode



Figure 14-14: Primary direction variogram for gold accumulation in NM Lode

Figure 14-15: Secondary direction variogram for gold accumulation in NM Lode



Figure 14-16: Primary direction variogram for antimony accumulation in NM Lode

Figure 14-17: Secondary direction variogram for antimony accumulation in NM Lode



Figure 14-18: Primary direction variogram for true thickness in NM Lode

Figure 14-19: Secondary direction variogram for true thickness in NM Lode



Table 14-5: Variogram Model Parameters

Variable Lode Rotations

Nugget 1st Spherical Structure 2nd Spherical Structure 3rd Spherical Structure

X Y Z Major Minor Spatial Variance Major Minor Spatial

Variance Major Minor Spatial Variance

Au Acc

NM

0 -90 0 0.10 11 6 116.5 26 20 41.2 77 28 59.3

Sb Acc 90 10 -90 0.19 8 6 25.5 25 20 82 92 25 7.1

True Thick 90 50 -90 0.21 23 22 0.06 95 28 0.02 - - -

Au Acc

NE

0 -90 0 0.10 11 6 116.5 26 20 41.2 77 28 59.3

Sb Acc 90 10 -90 0.19 8 6 25.5 25 20 82 92 25 7.1

True Thick 90 50 -90 0.21 23 22 0.06 95 28 0.02 - - -

Au Acc

P1

90 80 -90 0.38 47 12 210.36 - - - - - -

Sb Acc 90 80 -90 0.31 10 8 9.9 22 16 5.63 - - -

True Thick 90 80 -90 0.53 7 3 0.01 17 10 0.008 - - -

Au Acc

P2

90 80 -90 0.38 47 12 210.36 - - - - - -

Sb Acc 90 80 -90 0.31 10 8 9.9 22 16 5.63 - - -

True Thick 90 80 -90 0.53 7 3 0.01 17 10 0.008 - - -

Au Acc

CE

90 10 -90 0.41 10 3 14.8 25 7 295.3 - - -

Sb Acc 0 -90 0 0.14 6 8 10.2 40 17 20.5 - - -

True Thick 0 -90 0 0.50 9 3 0.005 17 11 0.049 - - -

Au Acc

CM

90 50 -90 0.27 23 8 319 84 23 166 - - -

Sb Acc 90 10 -90 0.07 6 6 56 20 8 82 - - -

True Thick 0 -90 0 0.19 10 3 0.066 26 5 0.038 50 11 0.032



Table 14-6: Search Parameters for first estimation run

Variable Lode

First Pass Second Pass Third Pass

Octant search

Min octant

Min sample

per octant

Max sample

per octant

Search Distance Min no.

sample Max no. sample


sample Max no. sample


sample Max no. sample Long

axis Short axis

Long axis

Short axis

Long axis

Short axis

Au Acc NM

100 40 4 12 250 100 3 16 500 200 1 3 On 1 1 8

Sb Acc 120 40 4 12 300 100 3 16 600 200 1 3 On 1 1 8

True Thick 120 40 4 12 300 100 3 16 600 200 1 3 On 1 1 8

Au Acc NE

100 40 4 12 250 100 3 16 500 200 1 3 On 1 1 8

Sb Acc 120 40 4 12 300 100 3 16 600 200 1 3 On 1 1 8

True Thick 120 40 4 12 300 100 3 16 600 200 1 3 On 1 1 8

Au Acc P1

70 25 3 6 175 62.5 3 16 350 125 3 16 On 1 1 8

Sb Acc 40 30 3 6 100 75 3 16 200 150 3 16 On 1 1 8

True Thick 40 25 3 6 100 62.5 3 16 200 125 3 16 On 1 1 8

Au Acc P2

70 25 3 6 175 62.5 3 16 350 125 3 16 On 1 1 8

Sb Acc 40 30 3 6 100 75 3 16 200 150 3 16 On 1 1 8

True Thick 40 25 3 6 100 62.5 3 16 200 125 3 16 On 1 1 8

Au Acc CE

30 10 4 10 75 25 1 3 150 50 1 2 On 1 1 8

Sb Acc 50 20 4 10 125 50 1 3 250 100 1 2 On 1 1 8

True Thick 20 15 4 10 50 37.5 1 3 100 75 1 2 On 1 1 8

Au Acc CM

125 45 4 12 312.5 112.5 3 16 625 225 3 16 On 1 1 8

Sb Acc 60 30 4 12 150 75 3 16 300 150 3 16 On 1 1 8

True Thick 90 40 4 12 225 100 3 16 450 200 3 16 On 1 1 8

Au Acc AS

100 35 2 4 250 87.5 2 4 500 175 1 2 On 1 1 8

Sb Acc 50 50 2 4 125 125 2 4 250 250 1 2 On 1 1 8

True Thick 70 20 2 4 175 50 2 4 350 100 1 2 On 1 1 8

Au Acc CD

50 50 2 4 125 125 2 4 300 300 1 2 On 1 1 8

Sb Acc 50 50 2 4 125 125 2 4 300 300 1 2 On 1 1 8

True Thick 50 50 2 4 125 125 2 4 300 300 1 2 On 1 1 8



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 drillhole samples, for 2 different cell sizes resulting in 2 models. When combined the high sample density blocks were cut to the boundary for the identified measured resource and added to the larger block model replacing the overlapping blocks. The cell sizes were selected based on the sample spacing of each area. Typically areas of high sample density are within developed areas in which face sampling has been conducted. Areas of low sample density include areas with drill intercepts with a data spread ranging from 80 m to 20 m. 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 14-7.

Table 14-7: 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 had been combined with depletion and resource categories applied the models were projected to the locations of the corresponding lode wireframe and expanded in the X direction to 10 m. The models from different lode structures are then combined into a global model that is used for mine planning and the creation of the reserve. The specifications for the global model are given in Figure 14-8.

Table 14-8: Mine Planning Regularised Block Model Dimensions

Block model origin Block dimension (m) Number of cells

X 15100 10 200

Y 4200 2.5 720

Z 720 5 90

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 with the estimated model grades in long section;

• Visual checks of the orientation of search ellipses and comparison to the sample and model grade distributions;

• Visual comparison of the estimated grades to previous estimates in long section;

• Comparison of the de-clustered sample grades with the model grades; and

• Plotting the sample and block estimated true thickness, gold and antimony grades on Swath plots.



The visual comparison of the sample thickness and grades with the estimated grades indicated that there was a favourable correlation. As expected a greater degree of smoothing of the grades was evident where there were fewer samples available.

Results from the comparison of the declustered, thickness weighted, composite grades with the model grades are shown in Table 14-9. An average of the tabulated results weighted by resource tones shows a 15% model undercall in gold grade and a 10% model undercall in antimony grade. The true thickness shows a model overcall of 4%. This is a reasonable result given the composited declustered data used is not top-cut. There are also localised variations in thickness, and grade due to the presence of large and small scale structural controls on mineralisation. Declustering over these areas leads to some mixing of grade populations. Some of the smaller and lower grade domains show greater miscorrelation due to small sample populations and small differences in the grade making a bigger difference to the magnitude of the discrepancy.

Table 14-9: Declustered (50 x 50 m) thickness weighted, composited sample grades and width compared to model grades

Au (g/t) Sb (%) True Thickness (m) Model/Composite

Lode Domain Declustered Composites Model Declustered

Composites Model Declustered Composites Model Au (g/t) Sb (%)

Vein Width

(m)

AS DOM 1 7.3 8.4 5.2 7.1 0.29 0.32 115% 137% 112%

CE DOM 1 64.5 65.1 25.1 23.1 0.22 0.23 101% 92% 105%

CM DOM 1 32.2 32.5 9.1 10.8 0.36 0.36 101% 118% 100%

CM DOM 2 4.9 3.9 2.6 1.7 0.32 0.33 80% 67% 103%

CM DOM 3 32.3 21.1 15.2 9.6 0.36 0.35 65% 63% 97%

NE DOM 1 22.6 15.5 10.4 6.5 0.21 0.23 69% 63% 109%

NM DOM 1 20.8 20.9 12.6 11.3 0.20 0.21 100% 90% 105%

NM DOM 2 64.3 59.1 36.0 37.1 0.22 0.26 92% 103% 118%

NM DOM 3 60.6 49.9 30.3 26.9 0.36 0.35 82% 89% 97%

NM DOM 4 62.5 57.6 29.3 31.3 0.24 0.24 92% 107% 100%

NM DOM 5 19.5 22.4 9.5 10.6 0.24 0.23 115% 111% 96%

P1 DOM 1 123.0 66.9 16.7 17.8 0.17 0.17 54% 106% 99%

P2 DOM 1 24.8 16.2 6.8 6.9 0.22 0.20 65% 102% 91%

The swath plots are illustrated for each of the major modelled lodes and in the case of NM and CM Lodes, represent a combination of all sub-domains (Figure 14-20 to Figure 14-27). They compare declustered, composited, accumulated grades with modelled accumulated grades. Composited data is declustered to 50 m x 50 m (Z and Y directions). Swath plot orientations illustrated here are by northing, along the strike of the lode which is the direction that has the greatest extent when compared to the east to west and the vertical directions.

Some of the discrepancies illustrated in the swath plots are attributed mixing of high and grade domains within the vertical swaths causing model grades to exhibit a miscorrelation with sample grades. This is most evident in the NM Lode Swath plots where model grade is consistently below composite grade. Analyses of the individual grade sub-domains display good concordance (Table 14-9). For further supporting swath analysis of NM Lode see Appendix A.



Figure 14-20: CE Lode Gold Accumulation Swath Plot by Northing

Figure 14-21: CE Lode Antimony Accumulation Swath Plot by Northing

Figure 14-22: CM Lode Gold Accumulation Swath Plot by Northing



Figure 14-23: CM Lode Antimony Accumulation Swath Plot by Northing

Figure 14-24: NE Lode Gold Accumulation Swath Plot by Northing

Figure 14-25: NE Lode Antimony Accumulation Swath Plot by Northing



Figure 14-26: NM Lode Gold Accumulation Swath Plot by Northing

Figure 14-27: NM Lode Antimony Accumulation Swath Plot by Northing

14.12 Mineral Resource Classification Mineral Resources have been classified with due regard to Mandalay’s experience in mining the deposit and the good reconciliation observed between previous block model resource estimates and the processing plant head grade during 2013 and 2014. Mandalay’s ongoing mining efforts continue to improve the geological confidence and the understanding of the controls on mineralisation which guides decisions made during the construction of the geological model. The classification criteria are the following:

• The Measured Resources are located within, and are defined by the developed areas of the mine. This criteria means that the estimate is supported by close spaced underground channel sampling and mapping;

• The Indicated Resource is located where drilling spacing on a nominal 40 mN x 40 mRL grid and there is good geological confidence in the geological model; and

• The Inferred Resource has 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 April 2014 (SRK, 2014).



14.13 Mineral Resources The Mineral Resources are stated here for the Augusta, Cuffley and Brunswick deposits with an effective date of 31 December 2014. The Mineral Resource is depleted for mining up to 31 December 2014. The figures also exclude remnant mineralisation that remains in pillars.

The Augusta, Cuffley and Brunswick deposits consist of a combined Measured and Indicated Mineral Resource of 999,000 tonnes at 7.5 g/t gold and 3.6% antimony, and an Inferred Mineral Resource of 519,000 tonnes at 5.3 g/t gold and 2.6% antimony.

This Mineral Resource includes a ROM stockpile comprised of 17,000 tonnes at 6.0 g/t gold and 2.6% antimony.

The Mineral Resources are reported at a cut-off grade of 3.8 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 USD12,000 per tonne as follows:

• AuEq =Au (g/t) + 2.03 x Sb (%)

The cut-off grade has decreased from, 3.9 g/t AuEq used in the previous resource report effective 31 December 2013, and the relative value attributed to antimony has increased slightly from 1.99 over the year. The changes in the cut-off grade and the AuEq formula are based on updated costs, recoveries and assumptions.

All relevant diamond drillhole and underground face samples in the Costerfield Property, available as of 30 November 2014 for the Augusta and Cuffley deposits were used to inform the estimate. The estimation methodology followed the approach used by SRK in December 2013 Resource Estimate reported in an SRK NI 43-101 Technical Report on the Property (SRK, 2014).

Mineral Resources for Brunswick have not been re-estimated since they were last reported in 2009 (Fredericksen, 2009). No drilling or mining has occurred on these structures since then. The Brunswick Mineral Resources are restated here for consistency with the other Costerfield lodes with an updated cut-off grade of 3.8 g/t AuEq and a 1.2 m minimum mining width criterion.

The effective date for the Brunswick Mineral Resources is 31 December 2014. The reporting of the Brunswick model in previous years has not been in line with the reporting methodology of Augusta and Cuffley Resources. This has been rectified in the 2015 estimation and has resulted in a drop in tonnes and grade of the Brunswick deposit.

Details of the Augusta, Cuffley and Brunswick Mineral Resources are stated in Table 14-10. Longitudinal projections of the models re-estimated at the end of 2014 are displayed in Figure 14-28 to Figure 14-34 with the same legend to enable comparison.



Table 14-10: Summary of the Augusta, Cuffley and Brunswick Mineral Resource, inclusive of Mineral Reserve

Lode Name Resource Category

Inventory (Tonnes)

Gold Grade (g/t)

Antimony Grade

(%)

Contained Gold (Oz)

Contained Antimony

(t)

Aug

usta

Dep

osit

E Lode Measured 55,000 5.4 2.9 10,000 1,600

Indicated 53,000 4.1 2.4 7,000 1,300

Inferred 86,000 3.1 2.7 9,000 2,400

B Lode Indicated 33,000 4.4 2.6 5,000 800

W Lode Measured 20,000 6.0 2.7 4,000 500

Indicated 25,000 5.3 2.8 4,000 700

Inferred 29,000 5.1 2.3 5,000 700

NM Lode Measured 54,000 11.5 6.0 20,000 3,300

Indicated 242,000 7.3 3.8 57,000 9,100

Inferred 121,000 4.6 2.3 18,000 2,800

NE Lode Measured 1,000 3.6 2.1 100 30

Indicated 52,000 3.4 1.6 6,000 800

Inferred 68,000 3.9 1.1 8,000 700

NSW Lode Indicated 9,000 3.7 2.4 1,000 200

P1 Lode Measured 14,000 9.8 2.5 4,000 300

Indicated 10,000 10.0 2.7 3,000 300

P2 Lode Indicated 12,000 6.9 3.2 3,000 400

Inferred 6,000 3.1 1.6 1,000 100

K Lode Indicated 29,000 2.9 2.0 3,000 600

Cuf

fley

Dep

osit

CM Lode Measured 40,000 17.7 6.9 23,000 2,800

Indicated 274,000 8.7 3.8 76,000 10,300

CE Lode Measured 11,000 16.9 5.2 6,000 600

Indicated 20,000 10.2 3.6 7,000 700

Inferred 6,000 13.7 6.5 2,000 400

CD Lode Inferred 56,000 8.9 3.4 16,000 1,900

AS Lode Indicated 28,000 4.8 3.7 4,000 1,000

Inferred 9,000 0.8 2.3 200 200

Brunswick Deposit Inferred 139,000 6.7 3.3 30,000 4,600

Stockpiles Measured 17,000 6.0 2.6 3,000 400

Total Measured + Indicated 999,000 7.5 3.6 242,000 35,900

Total Inferred 519,000 5.3 2.6 89,000 13,700

Notes:

1). Mineral Resources estimated as of December 31, 2014, and depleted for production through 31 December, 2014.

2). Mineral Resources stated according to CIM guidelines.

3). Mineral Resources are inclusive of Mineral Reserves.

4). Tonnes and contained gold (oz) are rounded to the nearest thousand; contained antimony (t) rounded to nearest hundred.




6). A 3.8 g/t Au Equivalent (AuEq) cut-off grade over a minimum mining width of 1.2 m is applied where AuEq is calculated at a gold price of $1,400/oz, antimony price of $12,000/t and exchange rate AUD:USD of 0.85.

7). The gold Equivalent value (AuEq) is calculated using the formula: AuEq = Au g/t + 2.03 * Sb%.

8). The Brunswick Mineral Resource has not been re-estimated since it was reported in Frederickson, 2009, Costerfield Gold and Antimony Project, Augusta and Brunswick Deposits. Frederickson Geological Solutions Pty Ltd.

9). The Brunswick resource reporting methodology has been reviewed and is now consistent with that of the Augusta and Cuffley Deposits.

10). The Mineral Resource estimation for Augusta and Cuffley deposits was performed by Chris Davis BSc, MAusIMM, who is a full-time employee of Mandalay Resources and was independently verified by Danny Kentwell MSc, BAppSc, FAusIMM, a full-time employee of SRK Consulting who is a qualified person under NI 43-101 and is the Competent Person for the Augusta and Cuffley Mineral Resource Estimates.

11). Danny Kentwell MSc, BAppSc, FAusIMM, full-time employee of SRK Consulting is a qualified person under NI 43-101 and is the Competent Person for the Brunswick Mineral Resource Estimate.

Figure 14-28: NM Lode Resource Block Model with Resource Category Boundaries Note: Black dots represent face and diamond drillhole samples used in the resource estimate.

Indicated

Inferred

Measured



Figure 14-29: NE Lode Resource Block Model with Resource Category Boundaries Note: Black dots represent face and diamond drillhole samples used in the resource estimate.

Figure 14-30: CM Main Lode Block Model with Resource Category Boundaries Note: Black dots represent face and diamond drillhole samples used in the resource estimate.

Indicated

Inferred

Measured

Indicated Inferred

Measured



Figure 14-31: CE Lode Block Model with Resource Category Boundaries Note: Black dots represent face and diamond drillhole samples used in the resource estimate.

Figure 14-32: CD Lode Block Model with Resource Category Boundaries Note: Black dots represent diamond drillhole samples used in the resource estimate.

Indicated

Inferred

Measured

Inferred



Figure 14-33: AS Lode Block Model with Resource Category Boundaries Note: Black dots represent diamond drillhole samples used in the resource estimate.

Figure 14-34: P1 and P2 Lode Block Models with Resource Category Boundaries Note: Black dots represent face and diamond drillhole samples used in the resource estimate.

Indicated

Inferred

Indicated

Inferred Measured

P1

P2



14.14 Cut-off Grade Calculations The 3.8 g/t AuEq cut-off grade was supplied by Mandalay and is based on recent cost, revenue, mining, and recovery data. AuEq is calculated at a gold price of USD1,400 per troy ounce and an antimony price of USD12,000 per tonne.

14.15 Reconciliation Reconciliation results for dry tonnes delivered to the Brunswick Processing Plant, gold (oz) and antimony (t) processed by the mill during 2014 are plotted in Figure 14-35, Figure 14-36 and Figure 14-37 respectively.

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 5% lower than produced by the plant. The tonnes of Antimony predicted by the model were 1% lower than produced by the plant. The 2013 models show poorer correlation, with Gold 16% lower and the Antimony 6% lower.

The tonnes predicted by the model were 8% lower than the tonnes processed by the mill which is largely attributed to additional tonnes from unplanned dilution. The most likely reason for the metal undercall by the model is that some of the dilution is mineralised. It is generally agreed that this is due to thin (<5 cm), discontinuous high grade veins in the hanging wall and footwall that are not included in the block model due to their erratic nature which makes them very difficult to model.

The reconciliation results give confidence to the sample collection procedures, the quality of the assays and the resource estimation methodology.

Figure 14-35: Mill Reconciliation – Dry Tonnes Note: Red mine recon. Line largely obscured by block model line.



Figure 14-36: Mill Reconciliation – Gold (oz)

Figure 14-37: Mill Reconciliation – Antimony (t)

14.16 Comparison with Previous Mineral Resource Estimate Table 14-11 compares the current model with the previous model, the latter of which has an effective date of 31 December 2013 (reported April 2014). Both sets of figures are diluted to 1.2 m. The following observations are made:

• Overall the measured and indicated resource tonnage has increased by 202,000 tonnes of ore with a contained metal increase of 4,000 oz gold and 3,500 t antimony. These increases equate to an overall drop in gold grade from 9.3 g/t to 7.5 g/t and antimony grade drop of 4.1% to 3.6%;

• The resource cut-off has dropped slightly from 3.9 AuEq (g/t) to 3.8 AuEq (g/t). The gold equivalent factor used has risen slightly from 1.99 to 2.03. These two changes have the net effect of increasing the tonnes whilst reducing grade for an overall increase in contained metal;

• Higher grade areas within E and W Lodes were depleted reducing resource tonnes and grade of remaining ore. Some high-grade remnant ore in to upper levels of E Lode also became serialised due to geotechnical instability during the year. This was taken out of the resource;

• Drilling and development along the N Lodes brought about an increase of inferred tonnes and a drop in measured and indicated tonnes; and

• Infill drilling and development on the Cuffley deposit doubled the measured and indicated tonnage whilst halving the inferred tonnes. The grade of the Cuffley deposit dropped due to mining in high grade areas of the deposit and converting lower grade material from inferred to indicated during the year. Through mining, high grade gold intercepts were also found to be part



of a parallel structure CE Lode rather than CM Lode as previously modelled. The result of this reinterpretation was a drop in gold grade within the CM Lode.

Table 14-11: Current Estimates Compared with the previous (December 2013) estimates for E Lode, W Lode, N Lodes, P Lodes, Cuffley lodes and Brunswick

Lode Resource Category

Tonnes Au (g/t) Sb (%)

Dec 2013

Dec 2014

14/13 (%)

Dec 2013

Dec 2014

14/13 (%)

Dec 2013

Dec 2014

14/13 (%)

E

Measured +

Indicated 116,000 108,000 93% 5.9 4.9 83% 3.3 2.7 83%

Inferred 83,000 86,000 104% 3.4 3.3 96% 3.0 2.8 93%

W

Measured +

Indicated 54,000 45,000 83% 6.3 5.5 88% 3.0 2.7 88%

Inferred 30,000 29,000 97% 5.2 5.4 103% 2.4 2.4 101%

NM, NE, NW, NSW

Measured +

Indicated 395,800 358,000 90% 7.7 7.3 95% 3.9 3.8 96%

Inferred 157,000 189,000 120% 5.0 4.3 85% 2.6 1.9 70%

P1, P2

Measured +

Indicated 1,640 36,000 2195% 21.1 8.6 41% 10.7 2.8 26%

Inferred 6,000 5.2 1.7

CM, CE, CD, AS

Measured +

Indicated 171,000 373,000 218% 17.9 9.7 54% 5.8 4.1 71%

Inferred 143,000 71,000 50% 12.6 8.0 63% 5.5 3.5 64%

Brunswick Inferred 157,000 139,000 89% 7.5 6.7 90% 3.9 3.3 85%

14.17 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 report.



15 Mineral Reserve Estimate From the Mineral Resource, a mine plan was designed based only on Measured and Indicated Resource blocks using predominantly the cemented rock fill blast hole stoping method. A cut-off grade of 5.0 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 the mine plan was demonstrated at metal prices of USD1,200/oz Au and USD10,000/t Sb prices. The Mineral Reserve estimate is presented in Table 14-1. The Measured and Indicated Resources within the mine plan were converted to Proven and Probable Mineral Reserves.

Table 15-1: Mineral Reserves at Costerfield, as of 31 December 2014

Category Inventory (Tonnes

Gold Grade (g/t)

Antimony Grade

(%)

Contained Gold (oz)

Contained Antimony

(t)

Proven 98,000 10.4 4.5 32,000 4,400

Probable 336,000 7.4 3.3 80,000 11,200

Proven + Probable 434,000 8.1 3.6 112,000 15,600

Notes:

1). CIM definitions were followed for classification of Proven and Probable Reserves.

2). Mineral Reserve estimated as of 31 December 2014.

3). Tonnes and Ounces are rounded to the nearest thousand; contained antimony rounded to nearest hundred.


5). Lodes have been diluted to a minimum mining width of 1.2 m for stoping and 1.8 m for ore development.

6). A 5.0 g/t Au Equivalent (AuEq) cut-off grade has been applied.

7). Commodity prices applied are gold price of USD1,200/oz, antimony price of USD10,000/t and exchange rate AUS:USD o 0.85.

8). The Au Equivalent value (AuEq) is calculated using the formula: AuEq = Au g/t + 1.97 * Sb%.

9). 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. The Mineral Reserve estimate was prepared by Shannon Green P.Eng., BEng, MAusIMM, a full-time employee of Mandalay Resources and was independently verified by Peter Fairfield, FAusIMM, a full-time employee of SRK Consulting and a qualified person under NI 43-101.

The Mineral Reserve does not include any portion of the Inferred Resources.

15.1 Modifying Factors

15.1.1 Mining Dilution and Recovery Mining Dilution 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; and

• 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 Augusta and Cuffley Lode mineralisation, Mineral Resources include an element of planned mining dilution, as Mineral Resources are reported to conform to 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 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 to date are consistent with the figures presented in Table 15-2. The long hole over break and dilution factors are considered to be consistent with operational experience as these measurements are based on stope inspection sheet measurement. Operating practices attempt to mine the stope as close to the lode width as possible.

The percentages of planned/unplanned dilution for the various Costerfield Operations mining methods are shown in Table 15-2.

Table 15-2: Augusta Mine Recovery and Dilution Assumptions

Mining Method Unit Airleg

Development

Airleg Half Uppers Airleg CRF

Long-hole Half

Uppers

Long-hole CRF

Planned Height Augusta m 2.8 3 4.2 4.2 4.2

Planned Height Cuffley m 2.8 4.2 4.2 7.2 7.2

Planned Width m 1.8 1.2 1.2 1.2 1.2

Overall Dilution % 15 0 15 25 25

Recovery Factor % 100 85 95 75 95

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. Airleg development dilution is based on end of month survey reports which compare actual drive area against the designed area. Dilution of airleg half uppers mining method is considered to be zero because the actual width of the panel is equal to or less than the planned width (i.e. 1.2 m). The remaining of stoping dilution figures are based on actual historical measurements.

Mining Recovery The recovery factor as shown in Table 15-2 represents the recovered proportion of planned mining areas, for the different mining methods. For stoping areas, visual inspections are carried out and recorded on stope inspection sheets if any ore is left in stopes. This data is used to estimate the recovery factor.

15.1.2 Mine Design and Planning Process All mine design work has been completed using Studio 5D Planner (S5DP) with all mine scheduling occurring in Enhanced Production Scheduler (EPS).

The mining shapes designed in Studio 5D Planner were assessed against the block model in order to calculate tonnes and grade. A 3D extended block model was used for the interrogation and mining depletion process of the designed mining shapes.

Mining depletion was accounted for as some of the designed mining shapes overlap the designed development or occurred within an area that had already been mined. Two types of mining depletion were considered;



• Mining depletion between the as-built (already mined area) and designed mining shapes; and

• Depletion between the designed development and designed stope shapes.

The first mining depletion type is accounted for in the block model by adding an attribute field to the mineral resource blocks that stated whether the block had already been depleted by mining and thus could not be reported in the Mineral Reserve. Using this depletion methodology, all overlaps between the mine design and the as-built was excluded from the Mineral Reserve. Figure 15-1 illustrates an example of overlap occurring between designed development and the as-built, while Figure 15-2 illustrates an example overlap occurring between designed development and planned stope shapes.

Figure 15-1: Overlaps between as-built and stope shapes which has been considered as depleted

Figure 15-2: Overlaps between stope and development shapes which have been considered as depleted

Development

Development

Stope

As-Built

As-Built

Stopes



The mining schedule evaluated mining areas within the Augusta and Cuffley Mineral Resources above the gold equivalent cut-off grade.

After completing the interrogation and depletion process, dependency rules including schedule constraints were applied to the designed shapes to link all the mining activities in a logical manner within in 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. Several scenarios were evaluated using EPS to ensure that the Mineral Reserve generated has the highest possible Net Present Value (NPV).

These design and scheduling programmes are also in use for quarterly and monthly planning purposes.

15.1.3 Cut-off Grade Estimation of the design cut-off grade was based on past mining experience gained within the Augusta deposit 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 current Augusta Mine as well as the planned future cost structure for the 10 m level interval spacing CRF mining method, a cut-off grade of 5.0 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. After completion of the economic modelling, the break-even cut-off grade was reviewed and estimated to be between 4.5 and 5.0 g/t AuEq.

For reporting purposes 5.0 g/t AuEq was applied to determine the Mineral Reserve estimate.

A further iteration was completed to review stopes that were included based on cut-off grade that required significant additional development. As a result of this some stoping areas were excluded from the Mineral Reserve in order to improve the NPV.



16 Mining Methods 16.1 Introduction

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 is serviced by a decline haulage system. 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.

The Augusta Mine employs predominantly airleg stoping methods as well as longitudinal uphole retreat working a bottom up sequence. These mining methods have been utilised throughout 2010 to 2014. 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 1,030 mRL. The Augusta Mine is scheduled to produce mill feed from three different mining methods: full face development, long-hole CRF stoping and half upper 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 plans to mine the extension of the Augusta Mineral Resources to the 829 mRL and the Cuffley Lode to the 814 mRL. The proposed stope outlines are presented in Figure 17-1.

Figure 16-1: Section of Proposed Augusta and Cuffley Underground Design



Figure 16-2: Plan View of Proposed Augusta and Cuffley Underground Design

16.2 Geotechnical

16.2.1 Augusta Overview Since 2006, the Augusta Mine has utilised seven different mining methods, a summary of these methods are as follows:

• Capital development undertaken with two boom jumbo;

• Level development undertaken with airleg (hand held drill);

• Flat back overhand ‘cut-and-fill’ development;

• Floor benching of level development;

• Crown ‘half upper’ stopes;

• Long-hole stopes with loose rock fill and varying level interval spacing (i.e. 4 m to 7 m); and

• Long-hole stopes with CRF and varying level interval spacing (i.e. 4 m to 10 m).

The chosen mining method for future stoping is long-hole stoping with CRF and varying level interval spacing. CRF and some cable bolting of the hanging wall on all levels prior to stoping provides the required support of the hanging wall and therefore rib pillars are unnecessary. To ensure medium term stability of the hanging wall, crown/sill pillars are left in place and extracted on the completion of a stoping panel as half upper stopes utilising a combination of hand held and long-hole mining techniques.

Numerical modelling using the program ‘FLAC3D’ was undertaken in March 2012 to examine the geo-mechanical performance of the planned long-hole extraction strategy, together with the performance of the CRF and crown/sill pillars at the Augusta Mine. The material properties of the weak, bedded/foliated siltstone, Orebody Domain, which makes up the hanging wall of the open stopes, was calibrated to the in situ conditions whereby excessive hanging wall failure and fall‐off occurs when the stope strike length is increased beyond 8 m. The primary failure mechanism was recognised to be tensile along the siltstone bedding/foliation/cleavage. The CRF pillars are predicted to remain stable throughout the extraction process.



16.2.2 Cuffley Overview The Cuffley Deposit is situated within the Costerfield area which is hosted by the Silurian Costerfield Siltstone. Lode mineralisation is associated with an 85° steeply dipping north-northwest trending shear zone. Underground operations within the Cuffley Deposit will be confined to a sequence of Turbidite Siltstones at a depth of between 200 m to 500 m below surface. This geotechnical assessment focuses on preliminary development and stoping conditions within the main turbidite sequences.

Proceeding from the surface to depth, the sedimentary sequence of host rocks is dominated by siltstones with minor sandstone interbeds. Due to this host rock lithological origin, five geotechnical domains have been defined. Major known fault intercepts were also investigated to produce a further five geotechnical domains; for a total of ten identified domains. Of these ten domains, six have been identified as important to determine major design principles for the Cuffley Deposit. The six geotechnical domains of interest are outlined below and shown in Figure 16-1.

• The Adder Fault ranges from 6 m to 20 m true thickness and dips at 20° to the south;

• The East Fault has been defined as having a true thickness of 600 mm and dips between 55 to 70° to the west. This structure is crucial to dominating placement of access infrastructure because it exists to the east of the Cuffley Lode where decline development and mine infrastructure has been conceptually designed;

• The Flat Shear tends to strike parallel with bedding and ranges in thickness from 500 mm to 2 m true width;

• The Capital Turbidite Host domain has a ‘very good’ Q rating rock mass quality. The true thickness of this unit is 100 m. Poor rock mass conditions are only encountered in the vicinity of fault structures. There is predominately only one joint set parallel to bedding and a high percentile RQD of 90%;

• Continuing beneath the Capital Turbidite Host Unit is a turbidite sequence referred to as the Eagle Formation due to its visually distinct bands of siltstone 1 cm to 2 cm thick. This domain has a ‘very poor to poor’ Q rating rock mass quality with an RQD of less than 50%. It has been interpreted with a true thickness of 90 m. Three or more joint sets exist in conjunction with infill material whose properties result in low friction values; and

• The Beach Unit also has a Q rating rock mass quality of ‘very poor to poor’ and an RQD of less than 50%. It has not been defined at depth due to drilling constraints. It includes three or more joint sets, including quartz veins, with low friction infill materials.



Figure 16-3: Geotechnical Domains – Section 5,010 mN

The Adder Fault, East Fault and Flat Shear geotechnical domains have a Q description of ‘exceptionally poor to extremely poor’ and an RQD of less than 50%.

16.2.3 Rock stress In the history of the Augusta Mine, twenty unconfined compressive strength (UCS) tests have been completed; fifteen of which have focused on the footwall and hanging wall rock mass of the lode material. This testing has concluded that UCS average values within the mineralised zones are similar and are approximately 100 MPa.

In situ rock stress measurements have not been undertaken at the Augusta Mine. For mining stress studies, principal stresses used are based upon those reported within a paper written by Lee, Mollison and Litterbach (2010) which collates over 1,000 stress measurements throughout Australia and New Zealand. The Costerfield Region (Augusta Mine) is within the Lachlan province where the principal stress orientation is west-northwest to east-southeast and with magnitudes of δ1= 55: δ2 = 35: δ3 = 30 for measurements greater than 500 m below surface.

16.2.4 Stope Design Criteria Review of the performance and placement of crown/sill pillars within the Augusta Mine has proven that the empirical rule of thumb that a crown pillar height should be at least equal to the maximum width of stopes to be mined, in the Augusta case this is approximately 3–4 m and has been validated.

The final stage of crown/sill pillar extraction consists of ‘half upper’ open stope retreat utilising hand held mining or long-hole retreat techniques. The chosen method is dependent on the level interval spacing. As the level interval spacing has been increased to 10 m within Cuffley, the majority of the crown/sill pillar recovery is completed using long-hole methods.



The majority of future CRF stoping is based on a level interval spacing of 10 m floor to floor (7 m stope height). A level spacing of 10 m results in a development extraction sequence ratio of approximately 30%. This is the preferred option because it has been observed that ground conditions, through the development process, deteriorate the closer the level interval spacing. One main advantage of the previous closer level spacing was the opportunity to support the mid span of stoping blocks.

A central mine access pillar has also been developed for the majority of stopes remaining below the 1,049 mRL in Augusta and all access development within Cuffley has been based on a central mine access pillar. The primary reason for the use of a central access pillar is to provide a smooth production profile due to increased development drive ore tonnes as the level can be mined in both north and south directions compared to a sole direction only (it doubles production development drive heading availability).

The early years of mining history at Augusta (above 1,075 mRL) has left behind what are termed ‘remnant’ mining areas. These areas are challenging for safe and effective extraction because the stress effects are assumed to be quite complex in the highly structured ground. Removal of a crown pillar in this region could increase stresses in an un-mined area that does not have a robust ground support system to adjust to the changes in deformation loading. This becomes critical if there are multiple structures involved and water present.

16.2.5 Ground Support Requirements The main ground control design method used at the Augusta Mine for permanent infrastructure areas, intersections and stopes is the ‘Rock Tunnelling Quality Index, Q’, Barton et al., (1974, 1993 Chart update). The use of the ‘UNWEDGE’ simulation software also assists with additional wedge stabilisation design and ground control material selection.

Stope support systems at the Augusta Mine have been based on the Mathews et al., (1981) Stability Graph Method for cable bolt design. Records of stope performance are kept and site empirical rules, with respect to stability, are based on back analysis with the Stability Graph Method.

The following ground support reinforcement is used underground at Augusta Mine:

• Black and galvanised friction bolts;

• Chemical resin bolts;

• Single strand bulbed-cable bolts;

• Black and galvanised mesh of all development areas;

• Black and galvanised steel mesh ‘W’ straps; and

• In-cycle fibrecrete, 75 mm in thickness, for major structures encountered by capital development.

Rock support and reinforcement design methods continue to evolve and develop as the mine matures. All permanent development is supported with galvanised friction bolts, chemical resin bolts and galvanised mesh. Major intersections are reinforced with single strand bulbed-cable bolts that are installed prior to the establishment of cross cuts.

All production drive development is supported with black steel friction bolts, chemical resin bolts and a combination of black steel mesh and black steel mesh W Straps. The turnout area formed by the central mine access pillar method is supported with galvanised ground support that extends 15 m in either direction and aims to provide stiffer ground support for final close out stopes (i.e. completion of stoping on a level).



Ground support standards for capital and production drive development exist for Augusta and as it has been validated that ground conditions within Cuffley are very similar to Augusta, the same ground support standards have been adopted. Pattern bolting with mesh to grade line or below, with cable bolts for spans greater than 5.5 m is a general description of the required ground support for all development. Fibrecrete will be used as required, particularly when faulted ground is intersected by capital development.

16.2.6 Incline/Decline Location The current incline/decline design stand-off from the Cuffley Lode ranges from 30 m to 70 m. The stand-off distance is dependent on the proximity of the decline to major structures that has led to instability issues in the past.

The proximity of the decline to the stoping blocks in the historical Augusta underground workings has been approximately 30 m. The decline has remained stable and unaffected by stoping through these areas.

16.2.7 References The geotechnical content contained within Section 16, relating to Cuffley, is based on work undertaken by Beck Engineering Pty Ltd. The work primarily focused on the geotechnical logging and interpretation of the Cuffley drill core and was undertaken in two stages:

• A desktop review of Mandalay’s proposed scope of work for the geotechnical logging to confirm sufficiency for this Technical Report; and

• A site visit to review geotechnical logging and interpretation of rock mass conditions and structures.

Two independent reports were completed as part of the geotechnical logging and assessment programme, these were as follows:

• Beck Engineering – Review of Cuffley geotechnical logging scope of work, 30 March 2012; and

• Beck Engineering – Review of Cuffley geotechnical logging and interpretation, 16 July 2012.

16.3 Augusta Mine Design

16.3.1 Method Selection Long-hole CRF stoping has been selected as the preferred mining method for the remaining Augusta Mineral Resource. This is based on the orebody geometry as well as the favourable experience gained through the use of this method as opposed to other trialled mining methods.

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 via the use of a half-upper handheld mining methods.

16.3.2 Method Description Augusta mines five individual lodes (W, NM, NE, P and E Lodes) have widths varying from 0.1 m to 1.5 m and dips of 45° to 85° which are favourable conditions for long-hole CRF stoping using handheld and mechanised mining techniques.

The sub-level spacing of 7 m floor to floor (4.2 m backs to floor), has been established to ensure stable spans, acceptable drilling accuracies and blasthole lengths. New areas of Augusta north of the older workings have been developed using 10 m floor to floor level spacing.



The production cycle for airleg long-hole CRF stoping, as illustrated by Figure 16-4, comprises the following:

• Develop access to the orebody;

• Establish bottom sill drive and upper fill drive;

• Drill airleg blastholes in a staggered ‘Dice 5’ pattern depending on ore width. Nominal stope design width is 1 m;

• Blast 2 m strike length of holes and extract ore;

• Place rock bund at brow of stope and place rock tube in stope. Rock tube is tightly rolled steel mesh placed in leading edge of stope prior to filling and eliminates the need for boring reamer holes in next stoping panel;

• Place CRF into the stope;

• Remove rock bund at brow of stope; and

• Commence extraction of adjacent stope once CRF has cured.

Figure 16-4: Augusta Long-hole CRF Stoping Method

Source: after Potvin, Thomas, Fourie, 2005.

The half upper stope geometry is engineered to negate the need to use remote control loaders for ore mucking and the mining cycle comprises the following:

• Drill 2.5 m–3 m blind airleg long-holes for a strike length of 3 m;

• Place a large waste bund toward the rear of the stope;

• Blast holes and extract ore;

• Leave a 2 m rib pillar where required by ground conditions; and

• Commence next stope.



16.3.3 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. A portion of suitable material is screened and utilised underground as roadbase.



16.4 Augusta Mine Design Guidelines

16.4.1 Design Parameters The mine design parameters that have been used in the design of the Augusta Mine are summarised in Table 16-1.

Table 16-1: Mine Design Parameters

Description Dimensions Gradient

Decline 4.8 mH x 4.5 mW 1:7 down

Decline sumps 4.8 mH x 4.5 mW 1:6 down

Stockpiles 4.8 mH x 5.0 mW 1:50 up

Level cross-cut 4.8 mH x 5.0 mW 1:50 up

Truck tip 6.5 mH x 5.0 mW 1:50 up

CRF mixing bay 4.8 mH x 4.5 mW 1:10 down

Operating level access 2.8 mH x 2.4 mW 1:7 up; 1:7 down

Production development 1.8 mH x 2.8 mW 1:50 up

Level spacing 10.0 m N/A

16.4.2 Mining Sequence The Augusta mining sequence will follow a bottom-up sequence, mining from the northern and southern extents retreating toward a central access. This sequence enables a consistent production profile to be maintained as it allows for dual development headings on each level. The current Augusta mining methods and extraction sequence is shown in Figure 16-5.

Figure 16-5: Long section of Augusta stoping between 1,015 mRL and 989 mRL



16.4.3 Level Development Ore drive development is designed to ensure the lode is positioned in the face to allow production drill positioning during subsequent stoping. Ore drives are excavated and supported by handheld mining methods that have been proven to be generally stable and productive. An arched profile is excavated to minimise dilution during development and for stability. All ore drives are developed under geology control. A typical cross section through the Augusta level development is shown by Figure 16-6.

Figure 16-6: Typical Section Looking North through Augusta Level Development

16.4.4 Vertical Development Ventilation rises of 2.4 m diameter have been excavated and supported between levels to extend the primary ventilation circuit. Mining of these ventilation rises was undertaken using ladder rising at an angle of 65° and on completion a 0.9 m diameter modular escape ladder way was installed as a means of secondary egress.

In addition, a 3 m diameter raise bore from surface to approximately the 1,020 mRL was completed in 2012. This shaft acts the primary exhaust shaft for Augusta, allowing the former Augusta return air rise (RAR) to become a fresh air intake.



16.4.5 Augusta Design Inventory The proposed underground stoping and development for the Augusta Mine is summarised in Table 16-2.

Table 16-2: Augusta Stope and Development Inventory

Level Development Stoping Total

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

3 0 0.00 0.00 1,121 5.02 3.29 1,121 5.02 3.29

829 125 2.61 1.88 3,023 5.78 3.15 3,148 5.65 3.10

839 3,497 4.30 2.58 4,735 5.88 3.94 8,231 5.21 3.36

849 5,581 4.48 2.97 8,648 6.45 3.99 14,229 5.68 3.59

859 3,763 5.29 2.76 6,719 7.79 4.43 10,482 6.89 3.83

869 3,826 5.48 3.39 2,189 9.42 5.84 6,014 6.91 4.28

879 3,922 5.39 3.34 7,152 7.68 4.60 11,074 6.87 4.15

884 0 0.00 0.00 1,158 5.07 2.58 1,158 5.07 2.58

889 2,781 4.99 2.63 5,177 7.05 3.70 7,958 6.33 3.33

891 0 0.00 0.00 1,370 4.97 2.39 1,370 4.97 2.39

898 211 2.35 0.82 1,263 5.17 2.44 1,474 4.77 2.20

899 2,806 4.96 2.61 4,808 6.68 3.53 7,614 6.04 3.19

905 0 0.00 0.00 936 10.24 3.85 936 10.24 3.85

909 2,908 5.22 2.76 2,283 7.81 4.16 5,191 6.36 3.38

919 1,482 6.93 3.44 6,015 7.34 3.82 7,496 7.26 3.75

926 0 0.00 0.00 136 12.61 5.85 136 12.61 5.85

929 1,789 5.36 2.69 4,367 8.94 4.46 6,157 7.90 3.95

939 2,041 7.69 3.87 6,114 9.55 4.81 8,155 9.08 4.57

949 1,936 9.33 5.82 5,661 8.83 4.86 7,598 8.96 5.10

959 0 0.00 0.00 5,666 8.91 4.67 5,666 8.91 4.67

969 0 0.00 0.00 4,087 9.30 4.64 4,087 9.30 4.64

976 1,132 2.30 1.58 1,523 2.90 3.13 2,655 2.64 2.47

979 0 0.00 0.00 2,329 10.53 6.16 2,329 10.53 6.16

980 1,675 8.00 2.48 3,548 9.94 3.26 5,224 9.32 3.01

986 1,082 1.88 1.80 1,470 2.60 2.16 2,552 2.29 2.01

989 0 0.00 0.00 1,742 14.06 6.98 1,742 14.06 6.98

990 1,709 6.10 2.30 3,166 8.39 3.17 4,875 7.59 2.87

996 1,138 2.48 1.44 1,599 2.95 1.53 2,738 2.76 1.49

999 0 0.00 0.00 385 12.47 6.54 385 12.47 6.54

1000 1,713 6.10 2.30 3,626 7.45 3.02 5,340 7.01 2.79

1005 0 0.00 0.00 644 12.74 6.24 644 12.74 6.24

1006 408 1.76 1.15 0 0.00 0.00 408 1.76 1.15

1010 1,693 4.94 2.22 3,237 6.75 3.04 4,930 6.13 2.76

1019 0 0.00 0.00 1,621 12.97 5.91 1,621 12.97 5.91

1020 1,745 4.81 2.16 4,737 5.19 2.63 6,482 5.09 2.50




Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

1029 317 5.40 1.99 1,517 12.68 6.35 1,834 11.42 5.60

1030 1,751 3.18 1.24 1,570 4.76 1.86 3,321 3.93 1.53

1034 0 0.00 0.00 1,126 3.10 3.38 1,126 3.10 3.38

1040 0 0.00 0.00 1,095 3.71 3.10 1,095 3.71 3.10

1045 0 0.00 0.00 1,517 4.80 3.60 1,517 4.80 3.60

1049 0 0.00 0.00 547 3.56 2.47 547 3.56 2.47

1060 0 0.00 0.00 2,462 3.71 2.76 2,462 3.71 2.76

1070 0 0.00 0.00 708 4.19 3.09 708 4.19 3.09

905subN 0 0.00 0.00 486 10.72 6.50 486 10.72 6.50

905subS 0 0.00 0.00 2,730 12.02 6.92 2,730 12.02 6.92

Total 51,032 5.21 2.79 126,013 7.59 4.06 177,045 6.90 3.69

Source: Resrve_150130_v10_Levels.xlsx.

16.5 Cuffley Mine Design The Cuffley Lode is accessed via a single decline that has been driven from the existing Augusta development at the 1,030 mRL. The decline has been excavated at 1:7 down and has been driven at 4.5 m wide by 4.8 m high. The Cuffley Decline currently extends to approximately the 895 mRL. At the 935 mRL, the Cuffley Incline extends off the Cuffley Decline and accesses mineral resources from the 945 mRL to the 1,005 mRL at time of reporting. 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 extended to 900 mRL by year end.

Mining is undertaken in Cuffley using a central access pillar recovery sequence using ‘bottom-up’ stoping panels. This extraction sequence necessitates the formation of crown pillars of which the majority are recoverable.

The decline stand-off distance (60 m) adopted for Cuffley was based upon infrastructure requirements, stockpile locations and the distance required to develop at a maximum grade of 1:7 to obtain the required sub-level spacing. The decline is in the hangingwall of the resource.



16.6 Cuffley Mine Design Guidelines

16.6.1 Design Parameters The mine design parameters used in the Cuffley design are summarised in Table 16-3.

Table 16-3: Mine Design Parameters

Description Dimensions Gradient

Decline 4.8 mH x 4.5 mW 1:7 down

Decline sumps 4.8 mH x 4.5 mW 1:6 down

Stockpiles 4.8 mH x 5.0 mW 1:50 up

Level cross-cut 4.8 mH x 5.0 mW 1:50 up

Truck tip 6.5 mH x 5.0 mW 1:50 up

CRF mixing bay 4.8 mH x 4.5 mW 1:10 down

Operating level access 2.8 mH x 2.4 mW 1:7 up; 1:7 down

Production development 1.8 mH x 2.8 mW 1:50 up

Level spacing 10.0 m N/A

16.6.2 Mining Sequence The mining sequence will follow 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 for the bottom-up mining sequence to commence from the 935 Level with no crown pillars planned in between this level and the 1025 Level. Elsewhere, crown pillars are usually formed at every fourth level allowing the stoping panel to comprise three CRF stoping blocks.

16.6.3 Level Development Production drive development is designed to be mined to ensure the ore is positioned in the face in order to minimise the hanging wall exposure. All production development is developed under geology control. Production drives are excavated and supported by handheld mining methods that have been proven to be generally stable and productive at the Augusta Mine in similar host rock ground conditions.

16.6.4 Vertical Development Ventilation rises of 3.5 m x 3.5 m are designed to be excavated and supported between levels to extend the existing primary exhaust system both above and below the Cuffley exhaust shaft bottom. Ladder rising will continue to be utilised for installation of escape ways providing a second means of egress.

16.6.5 Stoping Stoping of the remaining Cuffley Lode is proposed to utilise the same stoping methods employed at Augusta, namely longitudinal long-hole stoping in conjunction with CRF and crown half-uppers. Figure 16-8 is a long section of the Cuffley production drive development and stoping panels.

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 within the resource model.



Figure 16-7: Long Section of Cuffley stoping areas

16.6.6 Materials Handling All mineralisation and waste is proposed to be trucked from the mine via the internal Cuffley incline/decline as well as the 4800 decline. The bottom ore development level is currently situated at the 814 mRL, some 350 m below surface. Material is trucked to Augusta via the Cuffley decline and hauled to surface via the existing Augusta Mine decline. A private contractor rehandles the mill feed and transfers it to the Brunswick ROM pad to be stockpiled, blended and crushed prior to be being fed into the Brunswick Processing Plant.

All other waste hauled to surface is to be stockpiled at the Augusta Waste Rock Storage Facility. A portion of suitable material will is screened and utilised underground for roadbase.

16.6.7 Cuffley Design Inventory The planned stope and development inventory for Cuffley is summarised in Table 16-4.

Table 16-4: Cuffley Stope and Development Inventory


Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

814 4,874 2.81 1.43 8,841 4.78 2.22 13,715 4.08 1.94

824 5,363 4.32 1.69 5,174 6.32 2.28 10,537 5.30 1.98

834 4,891 4.78 1.75 8,969 6.67 2.67 13,860 6.00 2.35

844 5,378 5.26 2.02 9,411 7.33 2.78 14,789 6.58 2.50

854 6,724 4.77 2.07 5,851 7.22 3.55 12,574 5.91 2.76

864 6,423 6.40 2.81 11,751 8.54 3.79 18,174 7.78 3.44

874 6,999 6.37 2.70 13,109 9.40 4.10 20,107 8.35 3.61

884 7,429 7.31 3.29 7,393 10.30 4.58 14,822 8.80 3.93

894 4,512 7.70 3.03 7,828 11.63 4.17 12,340 10.19 3.75

895 1,081 5.90 3.15 3,977 16.57 7.08 5,058 14.29 6.24




Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

Inventory (Tonnes)

Grade (g/t Au)

Grade (% Sb)

904 4,212 10.05 3.24 5,595 12.77 4.07 9,807 11.60 3.71

905 0 0.00 0.00 4,344 19.24 7.30 4,344 19.24 7.30

914 2,508 6.23 2.15 6,400 15.37 4.24 8,908 12.80 3.65

915 663 8.78 2.93 4,031 17.83 5.86 4,693 16.55 5.45

924 3,713 10.83 2.93 4,736 16.38 4.54 8,450 13.94 3.83

925 0 0.00 0.00 2,135 13.50 4.84 2,135 13.50 4.84

934 0 0.00 0.00 2,269 21.96 6.49 2,269 21.96 6.49

935 0 0.00 0.00 1,896 12.88 5.15 1,896 12.88 5.15

944 0 0.00 0.00 1,188 18.53 3.75 1,188 18.53 3.75

945 344 6.58 3.41 3,606 9.63 4.36 3,950 9.37 4.28

953 1,432 4.45 1.87 1,576 6.96 3.80 3,007 5.76 2.88

954 0 0.00 0.00 2,214 11.00 2.75 2,214 11.00 2.75

955 0 0.00 0.00 5,110 10.71 5.05 5,110 10.71 5.05

963 2,087 4.97 2.36 1,784 6.10 4.07 3,872 5.49 3.15

964 0 0.00 0.00 1,711 8.79 3.01 1,711 8.79 3.01

965 0 0.00 0.00 4,245 10.24 4.87 4,245 10.24 4.87

973 1,496 4.72 3.10 1,073 5.20 4.26 2,569 4.92 3.58

974 0 0.00 0.00 825 7.41 3.32 825 7.41 3.32

975 0 0.00 0.00 2,958 11.60 6.20 2,958 11.60 6.20

985 471 6.74 4.82 3,874 13.31 6.07 4,346 12.59 5.94

995 1,092 6.67 4.35 4,186 13.68 5.24 5,278 12.23 5.05

1005 2,762 8.53 3.19 3,992 11.70 4.16 6,754 10.41 3.76

1015 2,077 8.26 3.00 3,885 8.09 4.95 5,962 8.15 4.27

1025 2,256 4.95 3.54 2,284 6.98 4.92 4,540 5.97 4.23

Total 78,787 6.27 2.58 158,221 10.58 4.21 237,009 9.15 3.67

Source: Resrve_150130_v10_Levels.xlsx.

16.7 Ventilation

16.7.1 Augusta Ventilation Circuit The current Augusta ventilation circuit comprises fresh air being drawn down the decline through the Augusta portal, with exhaust air exiting the mine through a series of short rises from 900 mRL to the 1,020 mRL and then a return air rise that extends to the surface from the 1,020 mRL. Fresh air from the decline is force ventilated into the active levels using secondary ventilation fans.

Two single stage 132 kW axial fans have been built into a bulkhead at the 1,020 mRL and act as the Augusta primary ventilation fans. These fans draw a total of 80 m3/s of air through the mine. Currently only one fan is in operation and draws a total of 45 m3/s which is sufficient enough volume to ventilate the limited stoping activities currently occurring within the Augusta 900 level.



16.7.2 Cuffley Ventilation Circuit The current Cuffley ventilation circuit comprises of fresh air being drawn from the connection to the Augusta workings at the 1,030 mRL. From the 1,030 mRL fresh air travels down the Cuffley decline to the 4800 decline where the primary ventilation flow splits. Primary ventilation flow travels down the 4800 decline and reports to the 949 mRL RAW from where it travels to the Cuffley RAR (950 mRL) exhausting to the surface. Primary ventilation flow continuing down the Cuffley decline splits at the incline turnout. From this point primary ventilation flow travelling up the incline reports to the 995 mRL RAW and primary ventilation flow travelling down the decline reports to the 915 mRL RAW. All exhaust ventilation flow travels back through a series of short rises and airways back to the Cuffley RAR where it exhausts to surface.

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 140m³/s of primary ventilation flow into the Cuffley workings.

Figure 16-8 shows the ventilation infrastructure for Cuffley and the connection to the Augusta ventilation circuit.

Figure 16-8: Proposed Cuffley Ventilation Circuit

16.7.3 Airflow Requirements The Victorian Mining Act 1958 doesn’t stipulate any minimum ventilation requirements, but the Occupational Health and Safety Act 2004 stipulates that “the importance of health and safety requires that employees, other persons at work and members of the public be given the highest level of protection against risks to their health and safety that is reasonable practicable in the circumstances”.

The minimum airflow required in Western Australia is stated as 0.05 cubic metres per second per kilowatt (m3/s/kW) of the maximum rated engine output (Mine Safety and Inspection Regulations 1995), this is considered to be “best practice” and therefore this standard has been applied to the ventilation calculations.

Table 16-5 summarises the current and future Augusta diesel fleet requirements and estimated primary airflow requirements.



Table 16-5: Augusta Minimum Airflow Estimate

Task Model Number Operating

Engine Power (kW)

Machine Airflow Requirement

(m3/s)

Task Airflow Requirement

(m3/s)

Truck MT436 1 298 14.9 14.9

Production LHD LH203 2 72 3.6 7.2

Tool Carrier JLG307 1 75 3.8 3.8

1020 Magazine 15.0

Total 445

40.9

Note: LHD – load-haul-dump.

Table 16-6 summarises the current and future Cuffley Lode diesel fleet requirements and estimated primary airflow requirements.

Table 16-6: Cuffley LOM Minimum Airflow Estimate (per circuit)

Task Model Number Operating

Engine Power (kW)

Machine Airflow Requirement

(m3/s)

Task Airflow Requirement

(m3/s)

Development LHD 1700G 1 258 12.9 12.9

Truck MT436B 1 298 14.9 14.9

Production LHD LH203 4 72 3.6 14.4

Tool Carrier JLG307 1 75 3.8 3.8

Grader 140H 1 138 6.9 6.9

Total 841 52.9

As Cuffley consists of essentially two circuits (i.e. 4800 Decline/Cuffley Decline & Cuffley Incline), the total primary ventilation requirement for Cuffley is estimated to be 105.8 m3/s.

16.7.4 Rise Sizes The parameters of the existing Augusta rises are as follows:

• Augusta Fresh Air Rise (surface to 1,020 mRL), 2.4 m diameter;

• Augusta RAR (below the 1,020 mRL), 2.4 m diameter;

• Augusta RAR (1,020 mRL to surface), 3.0 m diameter;

• Cuffley RAR (950mRL to surface), 3.0 m diameter;

• Cuffley RAR (above the 955mRL), 3.5m diameter; and

• Cuffley RAR (below the 955mRL), 3.5m diameter.

Based on the 105.8 m3/s requirement, the main Cuffley decline airflow will be at a maximum velocity of 4.5 m/s, which is well below the upper design limit of 6 m/s, which is the point at which dust would become airborne. The Cuffley exhaust rise will have a velocity of approximately 15 m/s, which is at the middle range of velocites that main exhaust shafts should be operated at.

The parameters of the designed Cuffley rises are in Table 16-7.

Table 16-7: Cuffley Rise sizes

Cuffley Rises Diameter

Cuffley RAR (Internal Raises) 3.5 m x 3.5 m

Cuffley RAR (950 mRL to Surface) 3.0 m diameter



16.7.5 Fan Duties The existing Augusta primary fans comprises two Zitron ZVN-1-14-132/4 single stage 132 kW axial fans installed in a bulkhead on the 1,020 mRL. The operating parameters of each fan are:

• Lower operating fan static pressure of 1,000 Pa at 55 m3/s; and

• Higher operating fan static pressure of 2,800 Pa at 35 m3/s.

At present, only a single fan is operating at a measured pressure of 1,130 Pa for a combined volume capacity of 48 m3/s.

The existing Cuffley primary fans comprises of 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; and

• Higher operating fan static pressure of 2,600 Pa at 30 m3/s.

At present, only two fans is operating at a measured pressure of 1,030 Pa for a combined volume capacity of 102 m3/s.

16.7.6 References The information presented in Section 16 is supported by a review undertaken by Brendan Clarke of Ozvent Consulting Pty Ltd as reported in Costerfield Operations Ventilation Review, 2014.

The scope of work for the ventilation review was:

• Review and confirm Augusta and Cuffley primary airflow requirements;

• Review and confirm the mine Ventsim model is suitable for long term planning and adjust model as necessary;

• Review ventilation requirements for the Cuffley long term plan. Confirm most viable long term solution;

• Review Cuffley internal shaft design. Recommend optimal shaft diameter;

• Review Augusta and Cuffley current primary and secondary ventilation systems including fans, vent walls and regulators, and identify areas for improvement; and

• Review main primary vent/regulator wall design for Cuffley 950 Level diamond driller access.

16.8 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 capital and operating development with the addition a cement slurry mix that results in a final product composing of 5% cement.

The use of paste fill was also considered as a possible alternative but it was found that the tailings from the Brunswick Processing Plant were unsuitable for backfill purposes due to the high moisture and clay content. The additional treatment required to make the tailings suitable for paste fill purposes resulted in paste fill being deemed too expensive.

CRF is mixed in batches of varying sizes with a Caterpillar 1700 (or equivalent) loader (1700 loader). The cement slurry is delivered underground to mixing bays via a cement agitator truck. The CRF has a cement factor of 5%. The hydrated cement mix is batched on surface using a cement silo on contract by Mawsons Ltd.



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.

The CRF is mixed mechanically in the mixing bays by means of 1700 loader prior to being trammed to the fill point of the stope to be filled using Toro 151 (or equivalent) loader. A bund is placed at an appropriate distance from the top of the stope to minimise potential for loader to overbalance into the stope void. Care is taken during placement of the CRF that the rock-tube is not displaced which is secured by chains during the filling process.

Nominal curing time is 12 hours and after approximately eight hours the rock bund placed at the brow of the stope is bogged away. There has been no observed failure of the CRF during this process.

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 are kept of batch quantities for all batches. The CRF has proved effective in minimising dilution during subsequent panel extraction as well as providing better ground stability and has eliminated the requirement for rib pillars.

16.9 Mine Services and Infrastructure

16.9.1 Electrical and Communications Electrical power for the Augusta site is provided by Powercor from the main electricity grid. The existing arrangement for electricity supply comprises:

• 2,000 kVA, 22 kV supply substation (Augusta Substation) at Augusta mine site; and

• 2,500 kVA, 22 kV supply substation (Cuffley Substation) at 8 Philips Lane.

A total available supply to these two substations is limited by the grid supply to 2,353 kVA. The present and projected consumption for the life of mine is below this threshold. The Augusta mine workings are fed from the Augusta Substation, while the Cuffley mine workings are fed from the Cuffley Substation.

High voltage (11 kV) power is fed underground from the Augusta Substation via a service hole to a 1,500 kVA 11 kV/1 kV substation at the 1,020 mRL. From this location, a high voltage (11 kV) cable is lowered further via an additional service hole to a 1,500 kVA 11 kV/1 kV substation at the 935 mRL.

High voltage (11 kV) power is fed underground from the Cuffley Substation via the Cuffley RAR to a 1,500 kVA 11 kV/1 kV substation at the 955 mRL.

From these underground substations, the power (1kV) is reticulated to the working areas via cable and distribution boxes. So far as is 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.

The Mine utilises a leaky feeder system for communications that will be extended as the mine develops into new production areas. Underground refuge chambers have direct communication to surface.

16.9.2 Compressed Air The existing Augusta Mine surface compressed air plant comprises four 600 cfm compressors. The overall plant capacity is 1,132 L/s (2,400 cfm).

Compressed air is delivered underground to the 1020 level in the Augusta Mine via an existing 10-inch poly line located in the Augusta RAR. The underground compressed air distribution system



from the 1020 level consists of 4-inch HDPE poly piping run in the decline shoulder. Each level is then supplied form the decline via 2-inch HDPE poly piping. Ring mains are installed where possible to increase the systems 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;

• Blasthole cleaning for development rounds;

• Pneumatic long-hole drills; and

• Long-hole cleaning.

16.9.3 Process and Potable Water Process water for the Augusta underground and surface operations is sourced from the Augusta Central Evaporation Pond located at the entrance to the site. Process water is distributed to the working levels via HDPE poly pipe in the decline. Process water is not potable (i.e. not for drinking).

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 20 Litre containers.

16.9.4 Explosives and Magazine 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 contract 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 bottom of the return air rise to surface at the 1,020 mRL. Table 16-8 states the current Augusta Magazine licence allowances.

Table 16-8: 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

16.9.5 Emergency Egress The ladder way design at Augusta comprises an enclosed ladder way with drop bars to allow personnel to rest during egress. At present the ladder way system extends from surface to the 880 mRL to ensure a secondary means of egress is provided throughout the Augusta Mine. The existing ladder way is entirely located within the exhaust airway system. The top portion (above 1,020 mRL) is located in fresh air, while below that is situated in the exhaust airway.

Emergency evacuation from Cuffley requires a combination of internal ladder ways and refuge chambers. An agreement has been established with a specialised crane contractor to provide a suitably sized crane to be mobilised to the collar of the Cuffley exhaust shaft for emergency egress



purposes. This crane connects to an Emergency Gig which is then lowered down the Cuffley exhaust shaft and is landed behind the bulk head at the 950 mRL. This Emergency Gig can be used to safely transport three personnel at a time.

The ladder way design from Cuffley is stand alone and comprises an enclosed ladder way within an 800 mm diameter rise which has drop bars to allow personnel to rest during egress. It is designed to link up the initial development of Cuffley on the 958 mRL up to the 995 mRL and down to the 820 mRL. Once personnel evacuate to the 958 mRL, emergency egress will be provided via the Emergency Gig.

16.9.6 Refuge Chambers Underground refuge chambers are installed underground in response to hazards posed by irrespirable atmospheres. Augusta has a fresh air base located at 3 Level and 1040 Level as well as a 4 man refuge chamber at the 905 Level. Cuffley has a 16 man refuge chamber at the 965 mRL and 950 mRL as well as a 20 man chamber at the 915 mRL.

The position of the refuge chamber facilities enables all personnel to be within 750 m of a refuge chamber, as recommended in the Western Australia ‘Refuge Chambers in Underground Metalliferous Mines’ Guideline (Department of Consumer and Employer Protection, 2008). Considering future mining activity, the expected number and distribution of personnel as well as the possible repositioning of existing refuge chambers, it has been determined that an additional 6-man portable Mine Refuge Chamber will suffice.

It is not intended for refuge chambers to substitute as a second means of egress, but to provide refuge during fire or containment when ladder ways may be inoperative or inaccessible.

16.10 Augusta Hydrogeology/Dewatering

16.10.1 Hydrogeology The hydrogeology of the Augusta Mine site consists of two main aquifers, the Recent Shallow Alluvial Aquifer (SAA) and the Silurian Regional Basement Aquifer (RBA), described below:

• The SAA comprises silts, sands and gravels, and is a perched groundwater system occurring sporadically across the site and within the confines of the creek and valley floors. There is clear evidence that this perched aquifer is laterally discontinuous and hydraulically isolated from the basement aquifer. There is also evidence to show that the alluvial aquifer is directly linked to the surface water systems; and

• The RBA comprises Silurian metasediments and forms the basement aquifer, where groundwater mainly occurs within and is transmitted through fracture systems. This aquifer appears to be semi-confined to confined, with groundwater occurring in fractures beneath the upper weathered profile, at depths of greater than 20 m below natural surface. The Augusta Mine dewatering activities are solely from this aquifer. Due to the nature of fractured rock systems, and the difference between many of the larger fractures in this system, it is difficult to specify aquifer properties for this unit.

In general, regional groundwater flow is from the south to the north. It follows the general trend of the topography also. The topography immediately surrounding the mine slopes downward from the southwest to the northeast. This trend broadly follows the trend of topography and groundwater flow across most of northern Victoria. Groundwater flows from the mountains immediately to the southwest of the mine site, and further to the southeast. The regional sinks are to the north, with the water eventually flowing into the deep lead system and into the Murray Darling system. Piezometer



measurements show the contours of the potentiometric surface, which clearly shows this flow direction, as illustrated by Figure 16-9.

Figure 16-9: Conceptual cross section of the groundwater flow paths

Depth to groundwater in the RBA is strongly dependent on the distance from the centre of pumping from the Augusta underground workings. There is also a dominant north south orientation to the drawdown pattern due to fault structures in the Silurian age mudstones which strike north, as shown by Figure 16-10. Close to the mine, groundwater levels are greater than 100 m below surface and rise steeply to be less than 15 m within 500 m in an east to west direction and 1,500 m in a north to south direction. Where the alluvial aquifer occurs, depth to groundwater is typically less than 4 m below surface.

Figure 16-10: Groundwater elevation contour map of the areas surrounding the Augusta Mine

16.10.2 Dewatering Dewatering of the Augusta Mine used to be undertaken from a combination of surrounding bores, shafts and directly from the decline. As the mine has extended to depth, dewatering Augusta has been via the existing rising main infrastructure. The process of dewatering in front of the mining levels is achieved by leaving the diamond drill holes drilled from underground open to drain. Due to



the fractured nature of the aquifer, the groundwater inflows are not always predictable. Flow rates have continued to increase as the decline advances and more drives are developed.

Removal of the process water and ground water is achieved by a staged pump station system comprising Weartuff WT084 mono pumps located at 890 mRL. They each have the capacity to pump 240 m of total dynamic head with a total system flow of 12.5 L/s.

The lower working levels utilise electric submersible pumps to direct water to the 890 mRL mono pumps. The 890 mRL mono pump lifts water to the 945 mRL permanent pump station which pumps water to surface via the Cuffley rising main. The mine owns three additional WT084 pumps which can be used as travelling mono pumps, allowing them to be positioned where required.

16.11 Cuffley Hydrogeology /Dewatering

16.11.1 Hydrogeology Much of the conceptual hydrogeology of the Cuffley Lode is based on information obtained at the Augusta Mine as well as some groundwater bores located in the vicinity of Cuffley. The regional hydrogeology comprises of two main aquifers, the SAA and the RBA, described below:

• The SAA comprises of silts, sands and gravels, and is a perched groundwater system occurring sporadically across the site and within the confines of the creek and valley floors. There is clear evidence that this perched aquifer is laterally discontinuous and is less common in the area around Cuffley due to its distance from surface water systems; and

• The RBA comprises of Silurian metasediments and forms the basement aquifer, where groundwater mainly occurs within and is transmitted through fracture systems. This aquifer appears to be semi-confined to confined, with groundwater occurring in fractures beneath the upper weathered profile, at depths of greater than 20 m below natural surface. Groundwater at Cuffley is likely to occur within the fractures of the RBA.

The Cuffley Lode is displaced from the historical workings of the overlying Alison Reefs via the Adder Fault. This fault is a sub-horizontal puggy fault and is considered to be responsible for partly dewatering the former Alison workings via the Augusta Mine to the east. Future dewatering of the RBA that exists around the Cuffley Lode will be undertaken by opening diamond drill holes that have not been grouted as well as expanding the underground workings to increase the footprint into which groundwater can be captured.

It is likely that the future dewatering of Cuffley Lode will be similar to groundwater flows and behaviour experienced at the Augusta Mine. Therefore, due to pre-dewatering of Cuffley, it is anticipated that groundwater inflow will be either less or equal to that experienced at the Augusta Mine. To date, this has proved to be the case with dewatering levels virtually stabilising as Augusta mining activities have decreased and mining activities within Cuffley have increased.

The risk to groundwater users and surface water through extraction at Cuffley is considered to be minimal. This is due to no groundwater users being present within the vicinity of the planned Cuffley workings and the RBA being too saline and yields too low to be a feasible target. In addition, the perched SAA, which is associated with the soaks and surface water system is disconnected from the RBA. This has been observed via long term monitoring of several bores within the RBA and SAA. It is expected that the cone of depression associated with dewatering Cuffley will mimic that seen at the Augusta Mine, being predominately north to south in direction, with minimal drawdown across strike. However, due to the proximity of Augusta to Cuffley and the access decline between the two working areas, the cones of depression will merge between the two mines to former a larger cone of depression.



16.11.2 Dewatering Removal of process water and ground water is achieved by a sole permanent pump station located at the 945 mRL. The pump station comprises three Weartuff WT088 mono pumps that have a capacity to each pump 240 m of total dynamic head with a total system flow of 37.5 L/s. The 945 mRL monos pumps direct mine water to surface via the Cuffley rising main.

The lower working levels utilise electric submersible pumps to direct water to the 945 mRL permanent pump station. This system will need to be extended as the mine progresses vertically. This will occur with the assistance of the three additional WT084 mono pumps currently owned by the mine.

16.12 Mineral Reserve Schedule

16.12.1 Development Schedule Development towards the Cuffley Lode commenced in the second half of calendar year CY2013 and is anticipated to be completed during the fourth quarter of CY2015.

The maximum capital development rates are dependent on the number of headings available for a single jumbo. The estimated capital development advance rates per month used in the development schedule are shown in Table 16-9. The handheld development rates are based upon the assumptions as shown in Table 16-10. The production development production rate is less than the waste development rate as it allows time for face sampling and poorer ground conditions.

Table 16-9: Jumbo Production Rates

Headings Available 1 2 3

1 Jumbo (m/mth) 140 160 180

Table 16-10: Handheld Production Rates

Description Value

Production crews 5

m Advance/cut 1.8

Cuts/face/week 3

m Advance/face/week 5.4

Available faces/crew/week 3

m/week/crew 16.2

Total m/week 81.0

Waste crews 1

m Advance/cut 1.8

Cuts/week 14

Total m/week 25.2

Weeks/mth 4.3

m/Mth (production) 348

m Waste/mth 108

Total m /mth 457



16.12.2 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 existing mobile equipment fleet is summarised in Table 16-11.

Table 16-11: Underground Mobile Equipment Fleet

Equipment Type Existing Fleet

2-Boom Jumbo 1

Production Drill 2

LHD - 1700 Loader 2

LHD - 151 Loader 7

Haulage Trucks 2

Charging Rig 1

Telehandler 1

Service Tractor 2

Light Vehicles 10

Total 28

16.12.3 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.

Shift Schedule Mining costs have been estimated using a continuous mining operation, 24 hours a day, 365 days/year. Augusta is a residential mine and therefore all employees will commute from their place of residence.

Operators and maintenance personnel work seven days on, then seven days off, 11-hour shifts alternating between dayshift and nightshift.

Augusta support staff work a standard Australian working week of five days on, then two days off, with an eight-hour dayshift only.

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 to relieve for sickness, annual leave and general absenteeism.

Current personnel numbers for total work force levels by department are shown in Table 16-12. It is expected the current staffing levels will be sufficient to successfully operate both the Augusta and Cuffley underground operations simultaneously.



Table 16-12: Personnel on Payroll by Department

Classification Number

Executive 2

Mining 96

Technical Services 13

Sustainability 6

Administration 14

Exploration 5

Plant 28

Total 193

Table 16-13 lists the mining personnel required for each Augusta underground mining crew on single shift basis.

Table 16-13: Underground Shift Mining Personnel

Classification Number

Shift Supervisor 1

Jumbo Operator 1

Loader Operator 5

Magazine Keeper 1

Truck Driver/Nipper 2

Long-hole Driller 2

Airleg Operator 6

Service Crew 2

Electrician 1

Charge Hand 1

Total 22

16.13 Schedule Summary A summary for the key physicals in the Mineral Reserve schedule are presented in Table 16-14.



Table 16-14: Summary of design physicals


Jumbo Capital Development m 2,115

Handheld Operating Development (Waste) m 2,353

Handheld Operating Development (Ore) m 9,456

Ventilation Raises m 90

Escape-way Raises m 126

Development Ore Tonnes tonnes 129,820

Development Ore Grade Au g/t 5.9

Development Ore Grade Sb % 2.7

Development Ore Grade AuEq g/t 11.1

Stoping Ore Tonnes tonnes 284,234

Stoping Ore Grade Au g/t 9.3

Stoping Ore Grade Sb % 4.1

Stoping Ore Grade AuEq g/t 17.4

ROM Ore Tonnes tonnes 11,640

ROM Ore Grade Au g/t 6.1

ROM Ore Grade Sb % 3.0

ROM Ore Grade AuEq g/t 12.0

Underground Stockpile Ore Tonnes tonnes 5,031

Underground Stockpile Ore Grade Au g/t 1.9

Underground Stockpile Ore Grade Sb % 0.6

Underground Stockpile Ore Grade AuEq g/t 3.1

Total Ore Tonnes tonnes 430,725

Total Ore Grade Au g/t 8.1

Contained Au ounces 112,206

Total Ore Grade Sb % 3.6

Contained Sb tonnes 15,670

Total Ore Grade AuEq g/t 15.3

Contained Metal AuEq ounces 211,456



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. The concentrator operates 24 hours per day, 7 days per week. Crushing is done on day shift and 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 second stage of milling to recovery a gold rich gravity concentrate which can be tabled and sold as a separate gold concentrate product which is sent to a refinery or blended with the final flotation concentrate. Secondary milled products are classified on 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. Concentrate is bagged for shipment to customers in China. The tailings are thickened before being pumped to a tailings storage facility (TSF) to the north of the Brunswick Processing Plant. The flowsheet is simple, conventional and is suited to processing the Costerfield ores in the LoM plan. With the inclusion of a mobile crushing unit, the capacity of the plant has been successfully upgraded to over 13,000 tonnes per month (t/mth). A summary processing flowsheet is provided in Figure 17-1. A more detailed processing description is provided below.

17.1.1 Crushing and Screening Circuit The crushing and screening plant consists of a new primary crushing circuit in closed circuit with a 16 mm vibrating screen. It uses a diesel powered mobile impact crusher. This crushed ore is conveyed to two x 120 tonne fine ore bins in parallel utilising the existing two stage jaw crushing circuit conveyors for transportation of ore to these bins. The old two stage jaw crushing circuit is still available for use providing additional redundancy. The primary jaw crusher has a closed side setting (CSS) of 45 mm, discharge is fed over a scalping screen with oversize being fed to the secondary jaw crusher. This delivers a product of approximately 50% passing 5 mm. A crushed ore reclaim system is also utilised so that crushed ore can be stockpiled on the ground in the event the ore bins are full. Crushed ore can then be reclaimed from the stockpile and fed directly into the primary ball mill if and as needed to improve overall milling circuit utilisation and enable crushing on day shift only.

17.1.2 Milling Circuit The fine ore is reclaimed from the crushed ore bins or stockpile via feeders which 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 concentrate that is sent either to the concentrate thickener or to the gold room for tabling. The gravity concentrate is typically sent directly to a refinery as a separate saleable product. The gravity tailing is pumped to classifying cyclones, the overflow of which becomes the flotation plant feed and 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.



A closed circuit mobile crushing plant (Powerscreen XH320SR) with a screen aperture of 16 mm was trialled during September 2012 to assess the impact it would have on the throughput of the existing processing plant. The trial results proved that this additional crushing capacity was able to improve throughput to over 10,000 t/mth. An even larger capacity mobile crushing unit (Finlay Terex) was employed on a hire basis and eventually purchased. It has been successful in increasing the throughput tonnage even further. The current mill capacity is capable of handling in excess of 13,000 t/mth. Plant throughput is shown below in Figure Figure 17-2.

17.1.3 Flotation Circuit 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 are used for cleaning duties. The cleaner capacity was increased from four to six cells in August 2014 to further improve product quality. 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.

17.1.4 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.

17.1.5 Tailings Circuit The flotation tails are settled in a thickener. Overflow is recycled through the plant as process water and the thickened solids are pumped to a tailings storage facility (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.

17.1.6 Throughput The Brunswick mill was expanded in 2012 with the implementation of a mobile crushing circuit upgrade. This has resulted in a step change in the plant capacity. The forecast plant throughput on the current ore feed blend is discussed in Section 13. The average forecast throughput for the remaining LoM is 12,306 t/mth, but is typically around 13,000 t/mth. Throughput for the 2014 calendar year was, 12,445 t/mth, with a maximum of 13,618 tonnes on a blend of 44% Cuffley/56% Augusta ore. Throughput is gradually trending higher as shown below in Figure 17-2. This historic throughput supports the forecast average and peak throughput in the LoM plan. The increased portion of Cuffley ore in the feed blend after 2015 is not expected to negatively impact throughput. The physical properties are similar and there is scope to increase the plant throughput further. The forecast throughput is therefore considered to be a reasonable estimate. Supporting historical data is provided below in Table 17-1 to Table 17-3.



17.1.7 Recovery The current plant recovery on the ore feed blend is discussed in Section 13. In summary, current 2014 YTD recoveries on a blend of 44% Cuffley:56% Augusta ore are 94.5% Sb and 89.6% Au. The recovery drop with the increased plant throughput in 2012 was negligible at around 1% for Sb and Au respectively. The minor drop in Sb is more likely to be associated with a decrease in Sb grade and has improved since the implementation of the additional cleaner cells. Recoveries have not changed significantly with the introduction of the Cuffley ore into the blend. The LoM plan will transition to mill a feed blend coming increasingly from Cuffley in 2015. The forecast recoveries are 94% Sb for the LoM and 90% Au in 2015 with a minor decrease in years 2016 and 2017 due to decreasing gold feed grade. Both are considered to be reasonable recovery estimates given both are supported by historic recoveries at a similar feed grades and grade/recovery relationships. It is also supported by test work. Further confidence is provided by the consistent recoveries of both Sb and Au across a range of feed grades. Supporting historical data is provided below in Table 17-1 to Table 17-3.

17.1.8 Concentrate Grade The current concentrate antimony grades have remained stable at 54% antimony. Two additional cleaner cells were added to the circuit after the crushing circuit upgrade to ensure product quality was maintained at the same recovery. Supporting historical data is provided below in Table 17-4 to Table 17-6.



Figure 17-1: Brunswick Processing Plant Summary Flowsheet



Figure 17-2: Brunswick Processing Feed Tonnages - January 2012 to December 2014



Table 17-1: Brunswick Mill Key Processing Results before Crushing Circuit Upgrade - February 2012 to August 2012

Feb-12 Mar-12 Apr-12 May-12 Jun-12 Jul-12 Aug-12

Average Feb-12 - Aug-12

DMT Milled 5,929 7,392 7,269 6,648 6,894 6,833 6,979 6849

Antimony recovery (%) 96 96 96 96 97 97 96 96

Gold recovery (%) 90 88 88 90 91 91 90 90

Antimony feed grade (%) 4.39 4.29 4.60 5.40 4.32 4.77 3.76 4.50

Gold feed grade(g/t) 9.10 8.25 8.22 8.78 8.03 8.81 7.09 8.32

Production prior to temporary crusher test period (Powerscreen unit) between 31 August and 11 October 2012. Terex crusher installed from 12 October 2012.

Table 17-2: Brunswick Mill Key Processing Results after Crushing Circuit Upgrade - September 2012 to December 2013

Sep 2012

Oct 2012

Nov 2012

Dec 2012

Jan 2013

Feb 2013

Mar 2013

Apr 2013

May 2013

Jun 2013

Jul 2013

Aug 2013

Sep 2013

Oct 2013

Nov 2013

Dec 2013

Avg Sep 2012

to Dec 2013

DMT Milled 9,299 10,151 10,242 11,523 9,984 8,243 8,351 9,906 11,546 11,166 12,389 12,592 11,113 11,572 10,994 11,786 10,678

Antimony recovery

(%) 96 96 96 94 96 96 96 96 95 96 95 96 95 95 95 94 95

Gold recovery

(%) 89 90 90 86 89 90 91 92 89 89 89 90 91 91 90 90 90

Antimony feed grade

(%) 4.16 4.58 4.37 3.33 4.38 5.37 4.65 3.72 3.80 4.05 3.86 4.66 4.86 3.71 3.91 3.52 4.18

Gold feed grade (g/t) 7.08 9.24 9.40 6.27 8.85 9.59 9.96 10.32 8.43 9.54 9.21 9.85 10.34 8.27 7.98 7.38 8.56



Table 17-3: Brunswick Mill Key Processing Results - January 2014 to December 2014

Jan 2014

Feb 2014

Mar 2014

Apr 2014

May 2014

Jun 2014

Jul 2014

Aug 2014

Sep 2014

Oct 2014

Nov 2014

Dec 2014

Avg Jan 2014

to Dec 2014

DMT Milled 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 94 95 94 96 94 94 94 95 95 94 95 95

Gold recovery (%) 91 90 90 89 89 88 88 87 91 91 90 92 90

Sb feed grade (%) 3.97 3.71 3.27 3.33 4.35 3.80 3.78 4.10 5.03 4.37 3.89 4.12 3.98

Au feed grade (g/t) 9.05 7.93 7.93 8.27 8.24 8.62 9.09 9.18 11.82 11.68 10.69 12.06 9.55

Table 17-4: Brunswick Mill Concentrate Production before Crushing Circuit Upgrade - February 2012 to August 2012

Feb 2012 Mar 2012 Apr 2012 May 2012 Jun 2012 Jul 2012 Aug 2012 Avg Feb 2012 to Aug 2012

Con DMT 467 578 597 629 529 566 455 546

Sb Con% 53 53 54 55 55 56 55 54

Table 17-5: Brunswick Mill Concentrate Production after Crushing Circuit Upgrade - September 2012 to December 2013

Sep 2012

Oct 2012

Nov 2012

Dec 2012

Jan 2013

Feb 2013

Mar 2013

Apr 2013

May 2013

Jun 2013

Jul 2013

Aug 2013

Sep 2013

Oct 2013

Nov 2013

Dec 2013

Avg Sep 2012

to Dec 2013

Con DMT 691 818 785 673 787 782 716 678 741 779 847 1,038 947 776 769 725 785

Sb Con% 54 55 55 54 53 54 52 52 54 54 54 54 54 53 53 54 54

Table 17-6: Brunswick Mill Concentrate Production – January 2014 to December 2014

Jan 2014

Feb 2014

Mar 2014

Apr 2014

May 2014

Jun 2014

Jul 2014

Aug 2014

Sep 2014

Oct 2014

Nov 2014

De 2014

Avg Jan 2014 to Dec 2014

Con DMT 877 750 770 705 985 784 878 1,002 1,089 975 831 854 875

Sb Con % 54 53 52 52 54 53 53 53 55 54 54 55 54



17.2 Plant Upgrades During 2014 the cleaner circuit was increased in capacity from four to six Denver DR15 cells to increase the cleaner residence time at the higher flows. This further debottlenecks the Brunswick Plant after the 2013 crushing and gravity circuit upgrades and allows for the antimony concentrate grade target of 54% to be maintained.

On the basis of current plant performance, no major capital projects are envisaged to achieve the forecast LoM throughput. Ongoing optimisation and minor, low capital cost debottlenecking projects should however continue to improve operations and increase capacity.

17.2.1 Crushing and Screening Circuit The initial trial in September 2012 of a portable crusher significantly improved the capacity of the mill. A larger portable crushing unit is now a permanent part of the process flowsheet configuration and enables a throughput of over 13,000 t/mth.

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. No further major work is planned at this time.

17.2.2 Milling Circuit The milling circuit remains unchanged. The finer crushed ore feed size allows the target throughput of over 13,000 t/mth to be achieved. Duty/stand-by pumps have been installed on the primary mill discharge hopper and Knelson (Gravity) Concentrator feed. No further major work is planned at this time.

17.2.3 Flotation Circuit During 2014 the cleaner circuit capacity was increased from four to six Denver square DR15 cells via refurbishment of offline equipment. Small improvements to the tank cell launders have also improved the efficiency of the circuit. This allows recoveries as well as final Sb concentrate grades to be maintained at the higher throughput. No further major work is planned at this time.

17.2.4 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.

17.2.5 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%).

Additional TSF storage capacity is to be provided during 2015 with the first stage lift on the Brunswick TSF. The first stage Brunswick TSF lift will provide an additional 197,000 m3 of storage. The Brunswick TSF first stage lift construction works are scheduled to commence in the fourth quarter of 2014. This raise will provide sufficient tailings storage capacity until the second quarter of 2017.



In addition to the planned first stage Brunswick lift, planning approval has been granted for the following:

• Bombay facility 327,000 m3 total increase (comprising of a 183,600 m3 lift and a separate 143,400 m3 lift); and

• Brunswick additional 110,000 m3.

The total Brunswick and Bombay approved increase in tailings storage allows for over seven years capacity at current milling rates.

After the 197,000 m3 lift on the Brunswick TSF has been completed and in use, the Bombay 183,600 m3 lift is the next to be built. This will provide approximately 19 months of capacity.

17.2.6 Reagent Mixing and Storage No upgrade work is required for the reagent mixing and storage area.

17.3 Recovery No major recovery projects are planned. Ongoing optimisation work will be undertaken as normal practice. All improvements planned to the existing circuit will be to maintain current levels of recovery or marginally improve it.

17.4 Services

17.4.1 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, tailing thickener overflow and TSF decant return water. The majority of the raw make-up water is provided by dewatering of the underground operations (approximately 1.6 ML/day). The plant operates with a positive water balance with excess water being disposed of. Mandalay Resources constructed a 2 ML/day permeate reverse osmosis (RO) plant at the Brunswick mill in 2014. This was run in conjunction with the 0.5 ML/day hire plant until Christmas 2014. The 0.5 ML/day hire plant was decommissioned and 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 will be completed in the first quarter of 2015. This will have 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.

17.4.2 Air The Brunswick Concentrator 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 filter press runs off high pressure air. The processing facility has adequate air to meet the LoM requirements.



17.4.3 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 current capacity of over 13,000 t/mth. The current calculated peak load is estimated at 1,000 kVA and occurs when the entire primary crushing circuit is operational (with the jaw crushers not running – however, if the mobile crusher unit is not operable the existing open circuit jaw crushing circuit can be reinstated).

The hired 0.5 ML/day reverse osmosis plant (decommissioned in December 2014) runs off mains power. The new RO plant is powered via diesel generator sets. This will continue to be its power source in the future in the future. The processing facility has adequate electrical power supplies to meet the LoM requirements. It can be simply expanded by adding additional generator sets in the future if required.



18 Project Infrastructure 18.1 Surface Infrastructure

The Costerfield Operations surface facilities are representative of a modern gold-antimony mining operation. The Augusta Mine site, refer to 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;

• Ventilation exhaust raises;

• Ventilation intake raise; and

• Augusta Mine dam to manage rainfall run-off.

Figure 18-1: Augusta Mine Site



The Brunswick site, refer to Figure 18-2, comprises the following:

• Gold-antimony processing plant and associated facilities;

• Central administration complex;

• Process plant workshop;

• Tailings storage facilities;

• Run of Mine stockpiles;

• Reverse Osmosis Plant;

• Previously mined Brunswick Open Pit; and

• Core farm and core processing facility.

Figure 18-2: Brunswick Site Area



18.2 Augusta Mine Underground Infrastructure

18.2.1 Ventilation System The primary ventilation infrastructure consists of two exhaust shafts known as the Augusta RAR and Cuffley RAR. These two shafts are both 3 metres in diameter and have been supported with shotcrete. The Augusta RAR extends from surface to the 1,020 mRL, with a series of short raises from the 1,020 mRL to depth to the 900 mRL or approximately 280 m below surface. This shaft system ventilates the Augusta Mine via two single stage 132 kW axial fans located in a bulkhead at the 1,020 mRL. Presently only one 132 kW fan is operation.

The Cuffley RAR extends from surface to the 945 mRL, with a series of short 3.5 m x 3.5 m raises extending down to the 915 mRL and up to the 995 m RL. This shaft system ventilates the Cuffley Mine via three single stage 110 kW axial fans located in a bulkhead at the 945 mRL. Presently only two 110 kW fans are in operation.

Fresh air is supplied via the decline portal and a small diameter FAR. The total primary ventilation flow is approximately 140 m3/s with 115 m3/s entering the portal and 25 m3/s entering the Augusta FAR. This primary ventilation flow is distributed through the mine using secondary fans positioned in the decline/incline that force ventilate the development levels. The primary ventilation flow exits the mine via the Augusta RAR at a rate of 45 m3/s and the Cuffley RAR at a rate of 95 m3/s.

18.2.2 Dewatering System The mine dewatering system consists of transportable pump stations and submersible pump systems that pump mine water (groundwater inflows and mine process water) directly to the 945 permanent Pump Station where it is pumped to surface via a rising main. This rising main extends to the Brunswick Pit which is used to settle out the solids before the clean water is directed to the evaporation/storage dams adjacent to the Augusta Surface facilities for evaporation and further distribution.

18.2.3 Infrastructure An underground crib room exists at the 1,085 mRL and the underground magazine exists at the 1,020 mRL.

In addition to the fixed plant, Mandalay owns, operates and maintains the majority of the mobile mining equipment including production drills, loaders, trucks, and ancillary equipment required to undertake ore development and production operations. The mobile mining equipment used to excavate the capital development is supplied, maintained and operated by an external contractor.

18.3 Tailings Storage Since operations began in the 1970s, two tailings dams have been constructed and operated, one of which have since been decommissioned and partially rehabilitated:

• Brunswick TSF; and

• Bombay TSF, that is currently operational.

Both TSFs were constructed based on a conventional paddock style/turkeys nest type design with earthen embankments.



All tailings are currently deposited in the Bombay TSF, on which a two metre lift was completed in the second half of 2013. This expansion brings the total storage capacity of the Bombay TSF to 361,000 m3 and will enable tailings to be deposited until the second quarter of 2015.

Additional TSF storage capacity is to be increased during 2015 with a 3.5 m embankment raise to be completed on the Brunswick TSF in the second quarter of 2015 which will provide an additional 197,000 m3 of tailings storage. The current strategy is to complete the Brunswick TSF embankment raise which will provide sufficient tailings storage capacity until the second quarter of calendar year 2017, after which it is planned to complete a 3.0 m raise on the Bombay TSF which will provide an additional 183,000 m3 of storage. Mandalay has received conditional approval to raise the height of the Bombay TSF.

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 2,353 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. Energy Australia (formerly Tru Energy) is the current electrical retailer.

Mandalay had planned to upgrade the Powercor distribution infrastructure to both the Augusta and Cuffley sites, but after significant evaluation has opted to maximise the mains electrical supply and supplement this feed with diesel generators on an as needed basis.

18.5 Power Reticulation Power for the Augusta underground operations is fed from a high voltage 22 kV Powercor feed that feeds:

• The Augusta surface switch room via a 1 MVA transformer that steps the voltage down to 415 V; and

• The surface switchyard via a 2 MVA transformer that steps the voltage down to 11 kV.

The 415 V supply is stepped up to 1,000 volt 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 V 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 1,020 mRL. From this location, the substation steps the voltage down to 1,000 volt and the power is reticulated to the working areas via cable and distribution boxes.

An additional 1.5 MVA 11 kV substation has been installed at the Augusta 935 mRL to service mine working load requirements below this level.

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 feed 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. 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. It is planned to relocate the 1.5 MVA substation from the Augusta 935 Level to the Cuffley 849 Level in mid-2015 to supply the at depth extension of the 4800 Decline.

A Power Factor Correction Unit, installed in 2013 on the 11 kV mine supply, increases the effective use and availability of power to the Augusta Mine. It is anticipated that this Power Factor Correction Unit will be relocated to the Cuffley Substation in 2015 as this substation will supply the majority of the power underground from 2015 onwards.

18.6 Water Supply Water for the Augusta underground and surface operations is sourced from the Augusta Evaporation Facility 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 permanent 945 Pump Station into which all mine water is fed from the underground pump system.

The Brunswick Processing Facility sources water from a number of sources including recycled process water from the Bombay TSF as well as standing water within the old Brunswick Pit.

The site is licenced to extract 700 ML per year from the Augusta and Cuffley underground workings.

Mandalay has a permit to discharge permeate water from the Reverse Osmosis Plant situated at the Brunswick site. The permit permits the discharge of up to 360 ML of permeate into the Mountain Creek South diversion, which runs into the Wappentake system, over a nine month period of the calendar year. A no flow period of 90 consecutive days per annum must occur as part of the permit which reflects the natural drying out of the creek in the dry part of the year. Permeate from the RO plant is utilised across site for a number of uses including dust suppression on roads and washing vehicles.



18.7 Waste Rock Storage An expansion of the waste rock storage facility located within the Augusta Mine footprint was recently approved by the regulatory authorities. Permission was granted to increase the height of the Augusta Waste Rock Storage Facility by 3.0 metres which has provided approximately 70,000 m3 of additional storage capacity. This volume reflects the temporary rock storage requirements of the current Life of Mine (LOM) plan. Currently no waste rock is being trucked off the mine site to the Brunswick TSF or Bombay TSF as all temporary waste rock storage facilities at the Brunswick site are at capacity.

The waste rock currently stored at the Augusta and Brunswick sites is to be used for future tailings storage facility expansions as well as rehabilitation requirements.

Waste rock is also being used to meet underground cemented rock fill (CRF) backfill requirements. Filling of mining voids using CRF steadily reduces the waste rock storage requirements and has been factored into the mine plan.

18.8 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 Open 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 of three dams with a total storage capacity of 150 ML. When some smaller catchment and operational dams such as the Mine Dam is 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, tailing thickener overflow and Bombay TSF decant return water. The majority of the raw make-up water is sourced from the Brunswick Pit. Whilst the process plant utilises water from a closed circuit, make-up process water is required to supplement water evaporated at the tailings dam.

Total evaporation/water disposal capacity including discharge of Reverse Osmosis treated water is currently estimated at 593 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. Hence, the current disposal of mine and process water has presented a significant challenge to the mine as a further 119 ML per year of evaporation capacity is required to dispose of the excess water.

To combat this water disposal issue, additional infrastructure is being constructed 2015 ensure sufficient capacity is available to dispose of all underground mine water (groundwater inflows and mine process water).



Mandalay purchased 30 Ha of land and a new Mining Licence (MIN 5567) was granted February 2013, with the intent of constructing an evaporation facility on the land which will provide an additional 120 ML per year (net) evaporation. Approval was granted in May 2014 and construction of the new Evaporation Facility known as the Splitters Creek Evaporation Facility commenced in the fourth quarter of 2014. The purpose of the facility is to evaporate groundwater extracted from the Costerfield Operations and thereby maintain dewatering rates from the underground workings.

In addition to the Splitters Creek evaporation facility, in early 2014, Mandalay submitted an application to the EPA to construct and operate a reverse osmosis water treatment plant at the Brunswick Plant to produce 2ML of treated water per day. Now operational, the water from his plant is discharged to a local waterway up to 9 months per year at a rate up to 2 ML/day.

With the new evaporation facility and reverse osmosis water treatment plant now operational, water disposal requirements will be met for at least the next three years.

18.9 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.

18.10 Diesel Storage A self bunded diesel storage tank of 60,000 L capacity exists at the Augusta Mine site. This diesel storage caters for all underground and surface diesel needs for both Augusta and Brunswick.

18.11 Explosive Storage The current Augusta Magazine is located at the 1020 Level 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

18.12 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 within the Augusta Mine.

A small maintenance/boilermaker workshop exists at Brunswick to assist with undertaking minor maintenance activities.



18.13 Housing and Land Mandalay owns ten 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 the Brunswick TSF embankment raise.



19 Market Studies and Contracts 19.1 Concentrate Transport

The concentrate is discharged directly into 1.5 t capacity bulk bags ready for transportation. The concentrate bags are loaded onto a road-train that transports the concentrate by road from the Brunswick Processing Plant to the Port of Melbourne. The average payload of each B-double3 road-train is approximately 42 t.

At the Port of Melbourne, the concentrate bags are loaded into shipping containers that are then stacked until required for ship loading. Shipments are normally scheduled at least once per month on a CIF basis to the destination port.

A third party trucking company collects the concentrate from the Brunswick site, transports, stores and loads the concentrate. The concentrate charges for transport from the Brunswick concentrate pad onto the ship in Melbourne has been estimated at AUD33 per dry metric tonne (dmt). 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 currency of sale 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 review of market reports by Roskill, the United States Geological Survey and public comments by major consumers such as Campine.

Globally, world antimony mine production in 2014 was estimated to have been 160,000 t contained antimony. China is the world’s largest producer of antimony, accounting for approximately 78% of world production in 2014. The Augusta mine’s production of 5,623 t of antimony therefore accounted for approximately 3.5% of global mine production, and 16% of production outside China.

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 (Figure 19-1). Antimony consumption is forecast to continue to grow moderately moving forward, but from a lower base following the drop in consumption after 2011 due to substitution of antimony in flame retardants in response to higher prices (Figure 19-2).4

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 2015.



Figure 19-1: Estimate of Global Antimony Demand by End-Use Segment Source: Roskill, USGS and industry reports.

Figure 19-2: Antimony metal prices 2010–2015

Moving forward the emergence of a substantial supply deficit is forecast in the 2016–2020 period, largely due to an expected reduction in mine output in China due to the lack of new discoveries, a determined effort by the Chinese Government to limit environmental damage from smaller producers and continued crackdown on illegal mining and smuggling (Figure 19-3). Many Chinese producers have already had to increase the level of concentrate imports in order to sustain domestic refined production, such that China imported an estimated 24,000 t of antimony contained in ores and concentrates in 2014.

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. These market dynamics are likely to result in the antimony price remaining in the range of USD7,000 to USD11,000 /t.

7,000

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Metal Bulletin Asian Metal



Figure 19-3: World Antimony production and consumption, 2000–2020 (t Antimony) Source: 2000–2012 Roskill, 2013–2020 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 Earths Elements (REE).

• USA: The “2013 Department of Defense Strategic and Critical Materials 2013 Report” reports foresees a shortfall of 20,500 tons of antimony in a four-year period, and recommends mitigation options to address this shortfall including strategic stockpiling of ~11,000t.5

• British Geological Survey: Ranked Antimony #3 on the “2012 Risk List”, down from its #1 ranking in the 2011 version of the same report.6

• European Union: An EU reported 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, the only metal among all critical raw materials to have a large deficit, defined as >10% of the market.7

5 “Strategic and Critical Materials 2013 Report on Stockpile Requirements”, U.S. Department of Defense, Office of the Under Secretary of Defense for Acquisition, Technology and Logistics, January 2013 (http://mineralsmakelife.org/assets/images/content/resources/Strategic_and_Critical_Materials_2013_Report_on_Stockpile_Requirements.pdf)

6 “Risk List 2012- Current Supply Risk Index for Chemical Elements or Elements Groups which

are of Economic Value," British Geological Survey (www.bgs.ac.uk/downloads/directDownload.cfm?id=2643&noexcl=true&t=Risk%20list%202012)

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19.3 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.

7 “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)



20 Environmental Studies, Permitting, and Social or Community Impact

20.1 Environment and Social Aspects

20.1.1 Mine Ventilation A ventilation shaft has been installed in Cuffley to maintain suitable air quality and volumes within the expanded underground mine. The Cuffley ventilation shaft is located on private land owned by Mandalay Resources and will act as the primary exhaust for the Cuffley development area. The facility includes a transformer and connection to an adjacent power supply line, as well as a rising main pipeline for removing groundwater from the Cuffley development area. The site is fenced to prevent public access.

20.1.2 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, except in the event of an ‘emergency’.

The climate in Central Victoria enables water to be removed through evaporation. Average pan evaporation is approximately 1,400 mm per year according to the nearest Bureau of Meteorology monitoring station at Tatura. 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 as 1,048 mm. Rainfall recorded in Heathcote in 2010 and 2011 was higher than average, with 979 mm and 759 mm respectively. Rainfall in 2013 and 2014 was below average at 554 mm and 510 mm respectively.

The Costerfield Operation currently operates a series of water storage and evaporation dams, including the following major storages facilities:

• Three evaporation dams;

• Brunswick Open Pit;

• Brunswick TSF; and

• Bombay TSF.

High efficiency evaporation sprays are also utilised at the Brunswick Open Pit to enhance evaporation rates.

The mining of Cuffley is expected to result in an increase in surface water disposal requirements, with up to 700 ML/year needing to be evaporated or disposed of in some way.

Current evaporation and re-use capacity is calculated to be approximately 370 ML/year, which is insufficient to sustainably manage current water disposal requirements. Therefore, a further 320 ML/year evaporation capacity is required to dispose of the 700 ML/year expected from future mine dewatering requirements.

Mandalay has purchased 30 Ha of land and a new Mining Licence has been granted with the intent of constructing an evaporation facility on the land which will provide an additional 120 ML per year (net) evaporation. Approval has been granted and in May 2014 the new Evaporation Facility known as the Splitters Creek Evaporation Facility commenced.



The purpose of the facility is to evaporate groundwater extracted from the Mandalay Resources Costerfield Operations and thereby maintain dewatering rates from the underground workings. The facility shall be constructed to include a series of shallow evaporation terraces that follow the natural topographic contours. Groundwater pumped from a dam or tank at the Augusta Evaporation Facility will be discharged at the highest point of the facility. Water will be allowed to cascade down the slope via the evaporation terraces to a deeper Storage Dam at the lowest point. A pump will be installed to reticulate 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.

In addition to the Splitters Creek evaporation facility, in early 2014, Mandalay submitted an application to the EPA to construct and operate a reverse osmosis water treatment plant to produce 2 ML of treated water per day. Now operational, the water from his plant is discharged to a local waterway 6 months per year. The 2 ML/day permeate reverse osmosis (RO) plant located at the Brunswick mill is being run in conjunction with the 0.5 ML/day hire plant. The 0.5 ML/day hire plant will then be decommissioned and the 2 ML/day plant will remain in operation as per regulatory approvals.

With the new evaporation facility and reverse osmosis water treatment plant operational, water disposal requirements will be met for at least the next three years.

20.1.3 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 a 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 the 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 the old Brunswick TSF, and has been used to permanently cap part of the TSF in accordance with the approved rehabilitation plan.

At the end of August 2012, the waste rock stockpile next to the Augusta Mine box-cut had reached its approved maximum capacity and some waste rock has been transported to the Brunswick TSF as capping material in preparation for the final closure of the facility. Waste rock has also been transported to the Bombay TSF to increase the height of the TSF.

Surplus waste rock will continue to be stored on the surface for future TSF lifts. A number of different locations are being considered for future waste rock storage, some of which will require a variation to the Work Plan. Waste rock will also be used in backfilling of the underground stopes, thereby reducing the quantity of waste rock brought to the surface. There is some opportunity to reduce the volume of waste rock to be stored through the re-use of the rock as engineering fill by third parties. The DSDBI has approved such off-site use, and markets are being investigated to evaluate whether significant volumes of rock can be managed in this manner, thereby reducing on site waste rock storage requirements. However, sufficient waste rock will need to be retained in order to fulfil rehabilitation and TSF expansion requirements.



20.1.4 Tailings Disposal In December 2014, Mandalay Resources has commenced construction of the raise the old Brunswick TSF facility. These facilities will provide sufficient tailings storage capacity for at least the next five years production.

20.1.5 Air Quality The approved Environmental Monitoring Plan for the Augusta Mine includes an air quality monitoring programme based on eight dust deposition gauges at various locations surrounding the Costerfield Operation plus 5 dust deposition gauges at the Splitters Creek evaporation facility and has been designed to confirm dust emissions remain below the prescribed limits. The monitoring data is provided to the regulatory authorities and Community Representatives through the quarterly Environmental Review Committee meetings. Given the similar nature of the activities associated with the deepening of the Augusta Mine, the current air quality monitoring programme is expected to be appropriate going forward.

Control measures in place to manage dust emissions from the operations currently and going forward include:

• Road watering programme with treated groundwater;

• Moisture control of mill feed during processing; and

• Maintaining moisture on TSFs and waste rock stockpiles.

Operation of exhaust ventilation shafts is not currently a source of dust emissions, and future ventilation arrangements will not contribute to dust emissions.

20.1.6 Groundwater Deepening of the Augusta Mine and the potential development of the Cuffley Lode is expected to result in an increase in groundwater ingress into the underground workings and therefore will result in an increase in mine dewatering rates. It has not been possible to accurately estimate future dewatering rates, however, a maximum dewatering rate of 700 ML/year has been assumed based on experience to date, compared to current dewatering rate of approximately 600 ML/year. A groundwater licence to extract 700 ML per year has been granted by Goulburn-Murray Water.

A conceptual hydrogeological model has been developed for the site based on current groundwater monitoring data. The model shows a cone of depression in the bedrock aquifer trending in a north to south orientation, parallel to the deposits, as shown in Figure 20-1.

The current cone of depression indicates some dewatering has already occurred along the line of the Cuffley Lode. The cone of depression is expected to extend north as development proceeds north from the Augusta Mine to the Cuffley Lode.

The conceptual hydrogeological model indicates that the Augusta and Cuffley deposits are located in the regional groundwater aquifer. The regional groundwater aquifer is confined to semi-confined and comprises 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 of 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, it appears that the current dewatering activities at Augusta do not affect the alluvial aquifer. Therefore, there is no impact to local landowners or the surface water system.

Figure 20-1: Groundwater elevation contour map of the areas surrounding the Augusta Mine as at February 2014



20.1.7 Noise The approved Environmental Monitoring Plan for the Costerfield Operation includes a noise monitoring programme 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 operations is a sensitive issue for near neighbours, and the company operates a 24-hour, 7 days a week 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, identify and implement immediate and longer term actions. All complaints are discussed at the quarterly Environmental Review Committee meetings.

The proposed expansion of the current Costerfield Operation is not expected to significantly change the nature of noise emissions from the site. Construction of new waste rock storage, TSF or evaporation facilities may require some additional noise monitoring which will be identified as part of the WPV approval process. Existing resources and procedures are expected to be adequate to accommodate any required modifications to the noise monitoring programme.

20.1.8 Blasting and Vibration DSDBI prescribes blast vibration limits for the protection of buildings and public amenity. The Company has procedures in place and monitoring equipment to measure blast vibration levels in order to assess compliance with the prescribed limits.

20.1.9 Native Vegetation The Costerfield Operation has been developed 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. Offset obligations for the Costerfield Operations activities to date has required the securing of 0.504 habitat hectares of various EVCs plus a number of old trees and consequently Mandalay retains a surplus of native vegetation offset for potential future needs.

The current expansion of the Costerfield Operation through construction of the Splitters Creek Evaporation Facility and Brunswick & Bombay Tailings Storage Facilities is expected to have minimal impact on native vegetation and the Peel Lane site is anticipated to contain sufficient offset credits to meet the site’s foreseeable future needs.

20.1.10 Visual Amenity The key aspect of the Costerfield Operation that may affect visual amenity is the construction of new groundwater evaporation facilities.

Community consultation will take place as part of the planning for any such facilities, and mitigation measures will be investigated and implemented where appropriate. Screening



vegetation shall be considered as part of the design of each of these facilities, in consultation with the relevant land manager and near neighbours.

20.1.11 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 these areas of interest that exist within the current Mining Licence (ML4644).

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; and

• New Alison Mine Shaft.

The expansion of the mining operations is not anticipated to result in any disturbance of historic mine workings or other heritage features.

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 footprint of the mine expansion does not fall into any designated Areas of Cultural Heritage Sensitivity, as per Map 7824 Heathcote, Department of Planning and Community Development, July 2012, therefore, it is not anticipated that the development plans will trigger the need for a Cultural Heritage Management Plan. However, Mandalay acknowledges the requirements of the Aboriginal Heritage Act and Regulations in relation to the protection of Aboriginal Cultural Heritage.

20.1.12 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 DSDBI 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.

Activities associated with community engagement in relation to the proposed expansion of the current Costerfield Operation, are described in a general sense within the Community Engagement Plan.

Key stakeholders for the mine expansion project are near neighbours, the regulatory authorities and the Company’s Environmental Review Committee. Most of these stakeholders have been provided with an initial briefing on the proposed expansion plans



through either the Environmental Review Committee quarterly meetings or during individual meetings with key regulators.

Community engagement is considered to be most important in relation to construction of new evaporation facilities. Whilst several of the land owners in close proximity to the proposed evaporation facility have objected to the planning application, the local Council has endorsed the project as an appropriate water management infrastructure to sustain ongoing mining activities at Costerfield.

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.

Mandalay has instituted a quarterly Environmental Review Committee which involves community and regulatory representatives, and this forum gives all parties the opportunity to find out about current and future issues at the mine and to provide their input.

20.1.13 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 up front, 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 will 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 removal of all plant and returning the disturbed area back 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 bond associated with the Company’s Mining Licence MIN4644 was reviewed and increased in mid-2012.

20.2 Regulatory Approvals

20.2.1 Work Plan Variation Future changes to mining activities such as potential changes to waste rock storage facilities will require a WPV to be approved. The DSDBI facilitates this approval process and will engage with relevant referral authorities, as required. The DSDBI may prescribe certain conditions on the approval, which may include amendments to the environmental monitoring



programme. 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.

20.2.2 Other Permitting In addition to the approval of a WPV, the proposed 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.

DPI Compliance with Native Vegetation Management Framework for removal of native vegetation associated with the power supply, evaporation facility and expansion of TSF footprints..

EPA EPA consent to discharge reverse osmosis 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, predominantly the past twelve months completed by Mandalay;

• Manufacturers and suppliers;

• First principle calculations (based on historic production values); and

• Costs include allowances for power, consumables and maintenance.

All cost estimates are provided in 2015 Australian dollars (AUD) and are to a level of accuracy of ± 10 per cent. Escalation, taxes, import duties and custom fees have been excluded from the cost estimates.

For reporting purposed summary tables provide estimates in AUD and US Dollars (USD). The USD have been estimated using the AUD:USD exchange rate of 0.85.

The capital and operating cost estimates have been consolidated into the file Costerfield Operations Financial Model_20150306.xlsx.

21.1 Capital Costs Table 21-1summarises the estimated total capital requirements for the Costerfield Operation. A detailed breakdown of the individual capital items included in the TEM was sourced from the 2015 budget document Version 3 and updated in January 2015.

Table 21-1: Costerfield Operation – Capital Cost Estimate

Item Total (AUD M)

CY 15 (AUD M)

CY 16 (AUD M)

CY 17 (AUD M)

CY 18 (AUD M)

Plant 4.4 2.5 0.5 1.4 0.0

Admin 0.2 0.2 0.0 0.0 0.0

Environmental 3.2 3.2 0.0 0.0 0.0

OH & S 0.0 0.0 0.0 0.0 0.0

Exploration 0.0 0.0 0.0 0.0 0.0

Mining 3.0 2.7 0.1 0.2 0.0

Ops Geology 0.0 0.0 0.0 0.0 0.0

Total-Plant and Equipment 10.7 8.6 0.6 1.5 0.0

Capital Development 7.8 7.8 0.0 0.0 0.0

Total Capital Cost 18.5 16.3 0.6 1.6 0.0

Total may not add-up due to rounding



21.1.1 Capital Lateral Development Decline development quantities have been based on the 3D mine designs prepared for the Cuffley Mine. 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, ventilation drives, sumps, truck tips and CRF mixing bays.

The unit cost for lateral development is based on the agreed development rates with the 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.

A Contractor has been engaged to complete a fixed package of work including the capital development required to complete the primary access development of the Cuffley Lode.

21.1.2 Vertical Development Vertical development quantities include all ventilation and escapeway raises to support the mine design. The estimated unit cost for ventilation raises has been based on the agreed development rates for the Contractor undertaking the capital development plus Cuffley historical costs for explosives and miscellaneous consumables.

The Contractor is excavating the ventilation raises using long-hole drilling and blasting techniques. It has been assumed that the short return airway extensions between levels will continue to be completed by a Contractor.

Point Break Mining, a specialist vertical development contractor, is currently contracted to undertake the vertical airleg rising and fit out of the escapeway raises within Cuffley. The estimated unit cost for escapeway raises is based on the Contractor development rate plus historical costs for explosives, ground support and ladders.

21.1.3 Infrastructure The main infrastructure cost items is the construction of the Splitters Creek Evaporation Facility that is required to manage the projected underground dewatering volumes as well as the planned embankment raises on both the Brunswick TSF and Bombay TSF. All associated costs are based on tendered unit rates for the specific design for each of these major infrastructure projects.

All remaining infrastructure costs have been based on Mandalay’s recent experience with similar installations.

21.1.4 Mobile Plant Mobile plant purchases include an underground haul truck, additional production drill, transit mixer and auxiliary equipment required to meet production targets.

Mobile plant cost estimates have been based on recent quotations or agreements from appropriate mobile plant suppliers.



21.1.5 Processing Plant Mandalay has identified and estimated the costs associated with the maintenance of the Brunswick Processing Plant and other mill site related initiatives including:

• Brunswick TSF Embankment Raise;

• Bombay TSF Embankment Raise;

• Electrical upgrades;

• Mobile loader and mobile crushing plant replacement;

• Refurbishment of existing plant and key components;

• Purchase of critical spares; and

• Miscellaneous surface facilities upgrades.

21.1.6 Drilling Drilling includes the direct cost of geotechnical, exploration infill and brownfields drilling. The estimated costs for drilling has been developed by Mandalay based on historical Augusta and Cuffley drilling costs for both surface and underground. The costs include set up, teardown and drilling costs of the current contractor (Starwest Pty Ltd) as well as the contractor’s indirect fees and profit.

Geotechnical drilling reflects the requirement to specifically target areas of anticipated poor ground conditions so that the decline path can be designed to minimise the impact of these areas on the decline advance rate and corresponding excavation and support cost.

21.1.7 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 tailings storage facility. 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 are summarised in Table 21-2 and Table 21-3 and described further in the following sections.

Table 21-2: Operating Cost Inputs


Mining

Jumbo Development AUD/m 4,664

Vertical RAR AUD/m 4,849

Vertical Ladderway AUD/m 1,300

Air Leg Development AUD/m 1,283

Stoping AUD/t 74

Supervision AUD/mth 337,238

Geology AUD/mth 150,860

Services AUD/mth 567,559

Processing Cost AUD/t milled 63

Site Services AUD/day 5,939

General and Administration AUD/day 11,023

Selling Cost AUD/t con 213

Table 21-3: Costerfield Operation - Operating Cost Summary

Description Operating Cost

AUD M AUD/t USD M USD / t

Mining 73 169 62 144

Processing 27 63 23 53

Site Services , General and Administration 18 42 15 26

Total 118 274 100 223

21.2.1 Lateral Development The estimated unit cost for lateral development has been developed from Augusta and Cuffley historical costs for labour, equipment, consumables, services, as well as achieved productivities. An allowance for the haulage to surface has also been included.

The lateral development (operating) for Augusta and Cuffley will continue to be undertaken as an ‘owner operator’. Using 3D mine designs the required development is summarised in Table 21-4.

Table 21-4: Summary of Development requirements


Jumbo Capital metres 2,115

Vertical Ladderway metres 126

Vertical RAR metres 90

Handheld Operating metres 11,753



The direct operating costs related to lateral development include:

• Direct labour;

• 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 bag and electrical cables; and

• Miscellaneous materials required to support development activities.

21.2.2 Labour Costs Labour costs have been estimated and include superannuation, workers compensation, payroll tax and partial allowances for leave accrual.

21.2.3 Production Stoping The direct costs related to production stoping has been developed from Augusta and Cuffley historical costs for labour, consumable material and equipment operating and maintenance, as well as achieved productivities associated with the following:

• Installation of secondary ground support;

• Drilling, loading, and blasting long-holes by Mandalay employees;

• Production from the stope with an LHD and tramming to a stockpile or truck loading area;

• Loading haul trucks from stockpile (if required); and

• Backfill preparation and CRF placement.

21.2.4 Augusta to Brunswick ROM Pad trucking 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 AUD4.00 per tonne delivered to the Brunswick ROM pad.

21.2.5 Processing Brunswick processing costs have been estimated from Mandalay actual 2014 processing costs.

21.3 Concentrate 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.



21.4 Royalties and Compensation 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 and is dependent on metal prices and exchange rates at the time of sale



22 Economic Analysis 22.1 Introduction

The Costerfield Operation Technical-Economic model was developed by Mandalay based on the production schedule and assumptions described in the earlier sections. All costs are in 2015 AUD with no provision for inflation or escalation.

The annual cash flow projections were estimated over the project life based on capital expenditures, operating costs and revenue assumptions. The financial indicators examined included pre-tax cash flow, net present value (NPV), internal rate of return (IRR) and payback period.

22.2 Principal Assumptions The key project criteria and assumptions used in preparation of the cash flow analysis have been listed in Table 22-1.

Table 22-1: Project Criteria


Proposed Mill Feed

Tonnes (kt) 431

Gold grade (g/t) 8.1

Antimony grade (%) 3.6

Project Life months 33

Average Production Rate t/mth 12,300

Maximum Mining Rate t/mth 13,800

Metallurgical Recovery* Gold (%) 85–90

Antimony (%) 94

Gravity Gold % 42

Concentrate Grade Gold (g/t) 71

Antimony (%) 53

Concentrate Selling Expenses AUD/dmt 213

Exchange Rate AUD: USD 0.85

Commodity Prices Gold USD/oz 1,200

Antimony USD/t 10,000

* Gold Recovery is 90% in 2015, 86% in 2016 and 85% in 2017

22.2.1 Metal Sale Prices Sale prices of metals are based on analysis of metal price predictions and the review of current and historical prices. Sensitivity analysis demonstrates the expected financial returns at a range of gold and antimony prices. Further information regarding the selected metal sale prices is provided in Section 19 of this Technical Report.



22.2.2 Concentrate & Gold Sales The TEM assumes that concentrate shipments and gold sales are made at the end of each time period, monthly. The payables of the shipments and gold sales as well as associated selling expenses are assumed to occur at these same time periods within the economic model.

The payable metal terms adopted in the economic model are consistent with the current sales contract terms for the gold and antimony concentrate grades and quality. It is also assumed that forty two per cent of the gold recovered reports gravity while the remainder reports to the concentrate.

22.2.3 Exchange Rate The economic model has assumed an exchange rate of AUD:USD 0.85 for the entire project life.

22.2.4 Taxes The Australian Government taxes on Mandalay’s Costerfield Operations include:

• A Goods and Services Tax (GST) at a rate of 10% is levied by the Federal Government on purchases by individuals and corporations on non-exempt goods and services. Businesses can claim back GST on most business inputs. It is assumed that all of the product sales will be to overseas customers, so no GST is applicable; and

• Company tax (30%) which is calculated on the profits generated by the operation.

As at the end of December 2014, Mandalay’s Costerfield Operations had carried forward tax losses of AUD20.6M.

22.2.5 Royalties/Agreements Under the Mineral Resources Development (Mining) Amendment Regulations 2010 of the Victorian State Government, royalties apply to the production of antimony. This royalty is applied at 2.75% of the revenue realised from the sale of antimony produced, less the selling costs.

There is no royalty payable on gold production.

There are compensation agreements in place with land holder owners and neighbouring residents which are affected by the Costerfield Operations. It has been assumed that the current agreements will remain in place for the remaining project life and that no new agreements will be required as the Augusta and Brunswick site footprint will remain largely unchanged.

Both the royalties and agreements have been factored in the financial model as an indirect cost and calculated on a daily basis.

22.2.6 Reclamation Possible salvage value on plant and equipment or profits from the sale of assets has not been included in the TEM. It has been assumed that cash flow and existing rehabilitation bonds will be used to pay for mine closure as well as any additional reclamation required.



22.2.7 Project Financing No assumptions have been made about the project financing in the TEM.

22.3 Economic Summary A summary of the economic factors associated with the project are presented in Table 22-2.

Table 22-2: Project Economics

Description Units Quantity Units Quantity

Tonnes Milled Tonnes 430,725 Tonnes 430,725

Recovered Gold Ounces 98,039 Ounces 98,039

Recovered Antimony Tonnes 14,730 Tonnes 14,730

Operating cost AUD M 118.0 USD M 100

Operating Cost per Payable ounce AUD / Oz Eq1 728 USD/oz eq 727.5

Capital cost AUD M 19.4 USD M 11.3

Payable Gold Ounces 84,990 Ounces 84,990

Payable Antimony Tonnes 9,280 Tonnes 9,280

Payable (Saleable) Metal – Au Eq Oz Eq 162,322 Oz Eq 162,322

Net Revenue (less selling expenses) AUD M 220 USD M 187

After Tax Cash Flow AUD M 81 USD M 69

1 Oz Eq – Gold Ounces + (Antimony Price / Gold price) * Antimony Tonnes.

Tonnes and ounces rounded to nearest thousand.

Million dollars rounded to nearest hundred thousand.

22.3.1 Cash Flow Forecast The estimated cash flow forecast has been provided in Table 22-3.



Table 22-3: Estimated Pre-Tax Cash Flow Summary



22.3.2 NPV The estimated after-tax NPV discounted at 5% interest has been calculated at AUD77.4 M.

22.3.3 Sensitivity The after-tax cash flow sensitivities (+20% and ‐20%) to gold price, antimony price, exchange rate (AUD:USD), metallurgical gold recovery, metallurgical antimony recovery, mill feed gold grade, mill feed antimony grade, capital costs and operating costs has been completed and presented in Table 22-4 and Figure 22-1.

Noting that the capital cost estimates used for the sensitivity analysis excludes the capital development as the cost driver for this is the unit operating costs. Capital development is included in the operating cost sensitivity.

Cashflow is most sensitive to exchange rate, mill feed tonnes, operating costs, gold price then gold recovery

Table 22-4: Sensitivity at 5% NPV After-Tax

-20% (AUD M)

-10% (AUD M)

Base (AUD M)

10% (AUD M)

20% (AUD M)

Gold Price 57 68 77 86 95

Antimony Price 59 69 77 85 93

Mill Feed Tonnes 47 63 77 90 102

Exchange Rate 36 58 77 94 109

Metallurgy Gold Recovery 57 68 77 86 95

Metallurgy Antimony Recovery 60 69 77 85 92

Mill Feed Gold Grade 57 68 77 86 95

Mill Feed Antimony Grade 60 69 77 85 92

Capital Cost 79 78 77 77 76

Operating Cost 96 87 77 67 55

Rounded to nearest million.



Figure 22-1: Sensitivity Analysis



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 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 (EL3316, EL5490 and EL5406), as shown in Figure 23-1, are at an early stage and not relevant to discuss further in relation to this Technical Report.

Figure 23-1: Augusta Mine Adjacent Properties (Source DSDBI Geovic, 2015)

The ownership and status of each of the surrounding exploration leases is shown in Table 23-1.

Table 23-1: Augusta Mine Adjacent Properties (Source DSDBI Geovic, 2015)

Title Owner Status First Granted Expiry

EL5406 Victoria Mining Exploration Pty Ltd Granted 11/01/2012 31/10/2017

EL5490 Iron Mountain Mining Granted 23/08/2013 5/12/2018

EL3316 Nagambie Mining Granted 19/12/1992 19/12/2014

MIN4465 Nagambie Mining Granted 12/11/1993 22/01/2014

The Costerfield Operation is situated 35 to 51 km from other significant central Victorian mining operations; Table 23-2 gives distance from Augusta Mine site to mines in Central Victoria.

Table 23-2: Distance from the Augusta Mine Site to Significant Central Victoria Mining Landmarks



Mine Owner Distance (km) General Direction

Nagambie Mine Nagambie Mining Ltd 40 east-northeast

Fosterville Mine Crocodile Gold Corp 35 northwest

Kangaroo Flat Mine Unity Mining Limited 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, the remnant mineralisation has not been included in the Technical 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 25.1 Geology

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. The core recovery is a potential problem for early stage lode definition only. No lodes are mined or classified as Measured Resources until they have been driven on and face samples obtained to establish the local grades with much more accuracy than the drill core.

Data comparisons show the unweighted average gold and antimony grade being higher within the face sample data than the drill holes. This is attributed to two factors:

• 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. There for 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.

• 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 invariable 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 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.


The 2014 Blank assay results show similar performance problems to those seen in 2013 with 92 of 345 results (26%) that are above the acceptable limit of 0.02 g/t. As above, the unacceptable assays were mostly caused by blanks following high grade assays to remove contamination. These results occurred during the first part of the year and combination of enforcing the procedure of random blank insertion and closer ongoing co-operation with Onsite appears to be remedying the situation. Given the very Gold high 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.

The reproducibility between laboratories of gold grades for 2014 is similar to that which was found in earlier check assay programmes from 2013 and 2012. In contrast, the reproducibility of antimony assays across laboratories decreased markedly from 92% RPD to 68% RPD. This reason for this difference is currently under investigation with a larger sample population being prepared for a three way check assay programme in early 2015. Preliminary investigations suggest that the apparent overcall of antimony grades by Onsite may be more reliable as their acid digest method has been optimised for high grade antimony through the addition of tartaric acid to prevent antimony precipitation out of the solution.

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 5% lower than produced by the plant. The tonnes of Antimony predicted by the model were 1% lower than produced by the plant. The 2013 models show poorer correlation with Gold 16% lower and the Antimony 6% lower.

The tonnes predicted by the model were 8% lower than the tonnes processed by the mill which is largely attributed to additional tonnes from unplanned dilution. The most likely reason for the metal undercall by the model is that some of the dilution is mineralised. It is generally agreed that this is due to thin (<5 cm), discontinuous high grade veins in the hanging wall and footwall that are not included in the block model due to their erratic nature which makes them very difficult to model.

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 increased by 202,000 tonnes of ore with a contained metal increase of 4,000 oz gold and 3,500 t antimony. These increases equate to an overall drop in gold grade from 9.3 g/t to 7.5 g/t and antimony grade drop of 4.1% to 3.6%.

The resource cut-off has dropped slightly from 3.9 AuEq (g/t) to 3.8 AuEq (g/t). The gold equivalent factor used has risen slightly from 1.99 to 2.03. These two changes have the net effect of increasing the tonnes whilst reducing grade for an overall increase in contained metal.

Higher grade areas within E and W Lodes were depleted reducing resource tonnes and grade of remaining ore. Some high-grade remnant ore in to upper levels of E lode also became serialised due to geotechnical instability during the year. This was taken out of the resource.

Drilling and development along the N Lodes brought about an increase of inferred tonnes and a drop in measured and indicated tonnes.

Infill drilling and development on the Cuffley deposit doubled the measured and indicated tonnage whilst halving the inferred tonnes. The grade of the Cuffley deposit dropped due to mining in high grade areas of the deposit and converting lower grade material from inferred to indicated during the year. Through mining, high grade gold intercepts were also found to be part of a parallel structure CE Lode rather than CM Lode as previously modelled. The result of this reinterpretation was a drop in gold grade within the CM Lode.

25.2 Mining SRK makes the following observations regarding the mining operations:

• 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 form 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 continuing to increase and improve the geological information and thus mine designs and planning.




26 Recommendations 26.1 Geology

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 programmes over and above the existing operational plans.

26.2 Mining 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 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.

Compiled by

Mr Peter Fairfield

Principal Consultant (Project Evaluations)

Peer Reviewed by

Ms Anne-Marie Ebbels

Principal Consultant (Mining)


27 References SRK 2013, Mandalay Resources Corporation, Costerfield Operation, Victoria, Australia; Preliminary

Economic Assessment, September 2013.

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Edwards et al., 1998. Heathcote and Parts of Woodlawn & Echuca 1:100,000 Map Area Geological Report, Geological Survey Report No. 108.

Evans-Wheeler J, 2007. Expert Tenement report. Memo to AGD from Tenements Administration Services.

Fredericksen D, 2009. Costerfield Gold and Antimony Project, Augusta and Brunswick Deposits. Fredericksen Geological Solutions Pty Ltd.

Fredericksen D, 2011. Augusta Project Mineral Resource Estimate for Mandalay Resources – Costerfield Operations. Fredericksen Geological Solutions Pty Ltd.

Haines Surveys, 2005; Job 0599 Costerfield Gravity Survey AGD Operations Pty Ltd.

Hanson N, February (1995) Costerfield Project Brunswick Reef Mine Resource Modelling and Estimations, Imago, Unpub. Report for Australian Gold Development NL.

Huxtable D, August 1972. Ore Potential of Brunswick Reef, Internal Mid-East Minerals NL report.

Kitch R, 2001. Independent Fairness Valuation of Costerfield Joint Venture Rob. Kitch and Associates Pty Ltd Unpublished Report for AGD Mining Ltd.

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Snowden, 2012. Mandalay Resource Corporation: Costerfield (Augusta) Gold-Antimony Mine: Mineral Resource and Mineral Reserve Estimate. Project No. 03151. NI 43-101 Technical Report.

Stock E & Zaki N, 1972; Antimony Dispersion Patterns at Costerfield. Mining and Geological Journal Vol. 7 No. 2 (1972) Geological Survey of Victoria.

Stockton I, August 1998. Brunswick Shear Project Prefeasibility Study.

Stoker PT, 2006. Newmont Australia Technical Services Sampling Notes. AMC Report. January 2006, 03-05pp.

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Thomas DE, 1937. Some notes on the Silurian Rocks of the Heathcote Area, Mining and Geological Journal Vol. 1 No. 1 Geological Survey of Victoria.

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UTS, 2008. UTS Geophysics Logistics Report for a Detailed Magnetic, Radiometric and Digital Terrain Survey for the Costerfield Project, carried out on behalf of AGD Operations Pty Ltd. (UTS Job #B054).



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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