Mandalay Resources Corporation Costerfield Operation ......Brett Muller, SRK Associate, Principal Consulting Engineer (Metallurgy), BEng (Minerals Engineering and Extractive Metallurgy),BCom

Mandalay Resources Corporation Costerfield Operation Victoria, Australia

Preliminary Economic Assessment

Prepared by

Authors Peter Fairfield, Principal Consultant (Project Evaluation), BEng (Mining), FAusIMM (No: 106754)

Brett Muller, SRK Associate, Principal Consulting Engineer (Metallurgy), BEng (Minerals Engineering and Extractive Metallurgy),BCom (Finance)

Dr Andrew Fowler, AMC Consultants Pty Ltd, Senior Geologist, PhD (Structural Geology), BSc (Hons), MAIG, MAusIMM CP (Geo)

Date of Report: 25 September 2013 Effective Date: 31 July 2013

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Mandalay Resources Corporation Costerfield Operation Victoria, Australia Preliminary Economic Assessment

Mandalay Resources Corporation 76 Richmond Street Toronto ONTARIO MSC1P1

SRK Project Number: PLI010

SRK Consulting (Australasia) Pty Ltd Level 8, 365 Queen Street MELBOURNE VIC 3000

Project Manager: Peter Fairfield

Date of Report: 25 September 2013 Effective Date: 31 July 2013

Compiled by

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

Authors:

Peter Fairfield, Brett Muller; Dr Andrew Fowler

FAIR/EBBE/hpowe PLI010_NI43-101_Costerfield_Operations_Report_Rev5 9 October 2013

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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 SRK’s 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

2 Introduction Peter Fairfield

3 Reliance on Experts Peter Fairfield

4 Property Description and Location Peter Fairfield Shannon Green

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

Peter Fairfield Shannon Green

6 History Dr Andrew Fowler Brian Cuffley

7 Geological Setting and Mineralisation Dr Andrew Fowler Brian Cuffley

8 Deposit Types Dr Andrew Fowler Brian Cuffley

9 Exploration Dr Andrew Fowler Luke Olson

10 Drilling Dr Andrew Fowler Luke Olson

11 Sampling Preparation, Analyses and Security Dr Andrew Fowler Chris Davis

12 Data Verification Dr Andrew Fowler

13 Mineral Processing and Metallurgical Testing Brett Muller Damon Buchanan

14 Mineral Resource Estimates Dr Andrew Fowler

15 Mineral Reserve Estimates Peter Fairfield Shannon Ainslie

16 Mining Methods Peter Fairfield Shannon Ainslie, Dan Lynch

17 Recovery Methods Brett Muller Damon Buchanan

18 Project Infrastructure Peter Fairfield Shannon Green

19 Market Studies and Contracts Peter Fairfield Manfred Ruff

20 Environmental Studies, Permitting and Social, or Community Impact Peter Fairfield Andrew Mattiske

21 Capital and Operating Costs Peter Fairfield

22 Economic Analysis Peter Fairfield

23 Adjacent Properties Peter Fairfield

24 Other Relevant Data and Information Peter Fairfield

25 Interpretation and Conclusions Peter Fairfield Dr Andrew Fowler

26 Recommendations Peter Fairfield Dr Andrew Fowler

27 References Peter Fairfield


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Executive Summary Introduction SRK Consulting (Australasia) Pty Ltd (SRK) has prepared the Costerfield Operation Preliminary Economic Assessment (PEA) to re-evaluate the viability of mining and processing mineralisation from continuing operations at Augusta Mine in connection with the addition of possible mineralisation sourced from the newly discovered Cuffley Lode which revised production scenario may have a material impact on the proposed production scenario disclosed in SRK Consulting (Australasia) Pty Ltd. March 2013, NI 43-101 report filed March 28, 2013.

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. The Cuffley lode located approximately 500 m to the north of the existing Augusta mine workings.is proposed to supplement the current Augusta Mine and extend the life of the Costerfield Operations.

Property Description and Location The Costerfield Operation encompasses the underground Augusta Mine, the Brunswick Processing Plant, the Cuffley lode deposit and associated infrastructure, all of which is located within mining licence MIN4644. The mining licence is located with exploration lease EL3310 which is held 100% by AGD.

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 planned to be 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 90,000 people, located 50 km to the west-northwest. Costerfield Operation is a residential operation with personnel residing throughout central Victoria as well as Melbourne. Local infrastructure and services are available in Heathcote, the 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 Industry (DPI) as State Forest. The Puckapunyal Military Area is located on the eastern boundary of the Project Area.


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The area has 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 1850’s 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 1860’s 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 ozs 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. This Inferred Resource was further increased in May 2012.

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 Lodes comprise sub-parallel, west-dipping, quartz-stibnite veins denoted (from west to east) ‘N Lode’, ‘W Lode’, ‘E Lode’ as well as a number of additional veins, all of which are located in the west-dipping Costerfield Siltstone host-rock. Mining has demonstrated that mineralised splay veins can also be economic 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 Augusta Mine is currently extracting ore from the W Lode, E Lode and N Lode.


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The Alison-Cuffley lodes are located 500 m northwest of the Augusta Mine. West and East Reefs were mined in the Alison Mine down to a flat-fault at a depth of 130 m. Drilling in 2011 revealed a displaced lode below the fault, now named the Cuffley lode. The Cuffley lode is a steep east-dipping, quartz-stibnite-gold lode similar to the Augusta Lodes and is open at depth and to the south.

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 that 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 échelon 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 Costerfield property includes the Augusta and Brunswick deposits, and the Cuffley lode. The Augusta deposit comprises the E, W, NE, C, NW, and P lodes. Mining at the Augusta deposit is ongoing with production during 2013 of approximately 9,500 tonnes per month at 9.4 g/t Au and 4.3% Sb. Mining has not commenced on the Brunswick deposit or the Cuffley lode.

The Augusta and Brunswick deposit Mineral Resources are stated here with at an effective date of 30 April 2013. The Cuffley lode Mineral Resource estimate is updated in this report with newer drillhole data and is stated with an effective date of 30 June 2013. Readers are cautioned that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Augusta and Brunswick deposits consist of a combined Measured and Indicated Mineral Resource of 506,000 tonnes at 6.6 g/t gold and 3.2% antimony, and an Inferred Mineral Resource of 407,000 tonnes at 3.7 g/t gold and 2.2% antimony. This Mineral Resource includes the run-of-mine (ROM) stockpile, and is depleted for mining up to the 30 April 2013. The figures also exclude remnant parts of the Resource that are no longer accessible, and have been sterilised by mining.

The Cuffley lode consists of an Indicated Mineral Resource of 133,000 tonnes at 16.9 g/t gold and 5.4% antimony, and an Inferred Mineral Resource of 273,000 tonnes at 10.4 g/t gold and 3.2% antimony.

The Mineral Resources are reported at a cut-off grade of 3.6 g/t gold equivalent (AuEq), with a minimum mining width of 1.8 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) + 2.22 x Sb (%)


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The cut-off grade has decreased from 4.7 g/t AuEq, and the relative value attributed to antimony has increased from 2.02 since the 31 December 2012 Mineral Resource estimate (SRK, 2013). The changes in the cut-off grade and the AuEq formula are based on updated costs, recoveries and assumptions.

The Mineral Resource was estimated using a block model and a 2D accumulation method. All relevant diamond drillhole and underground face samples in the Costerfield Property, available as of 31 December 2012 for the E, W, NE, NW, and P lodes, and as of 30 June 2013 for the Cuffley lode were used to inform the estimate. The estimation methodology followed the approach used by AMC in the March 2012 Resource estimate reported in a Snowden NI 43-101 Technical Report on the Property (Snowden, 2012).

Mineral Resources for C Lode and 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 C Lode and Brunswick Mineral Resources are restated here for consistency with the other Costerfield lodes with an updated cut-off grade of 3.6 g/t AuEq and a 1.8 m minimum mining width criterion. The effective date for the C lode and Brunswick Mineral Resources is 30 April 2013. AMC has reviewed the work carried out in 2009 and Dr A Fowler takes Qualified Person responsibility for the restated Mineral Resources.

The estimation method involves the gold and antimony grades and thickness being estimated into a two dimensional (2D) block model. This requires that gold and antimony are multiplied by true thickness (called gold accumulation and antimony accumulation), to correctly assign weights to composites of different lengths during estimation. The estimation method is ordinary kriging, where there are sufficient sample pairs for meaningful variography. Otherwise, inverse distance squared is used. The estimated grade is then back-calculated by dividing estimated gold accumulation and estimated antimony accumulation by estimated true thickness.

The ROM stockpile on 30 April 2013 is included in the Augusta Mineral Resource statement. The tonnes of the stockpile are based on survey information. Stockpile grades were averaged from face samples that were sampled from the areas of the mine that contributed to the stockpile.

A summary of the Augusta and Brunswick Mineral Resources is stated in Table ES-1. The Cuffley lode Mineral Resource is stated in Table ES-2.

Table ES-1: Augusta and Brunswick Mineral Resource Summary on 30 April 2013

Resource Category Tonnes Au (g/t) Sb (%) AuEq (g/t)

Measured 184,000 7.1 4.4 16.9

Indicated 323,000 6.3 2.5 11.8

Inferred 407,000 3.7 2.2 8.6

Notes to Table ES-1: 1. Mineral Resource stated as of 30 April 2013. 2. Mineral Resource used relevant sample data available as of 31 December 2012. 3. Mineral Resource stated according to CIM guidelines. 4. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. 5. A 3.6 g/t AuEq cut-off grade is applied where AuEq is calculated at a gold price of USD1,400 per troy ounce and an

antimony price of USD12,000 per tonne. 6. Tonnes have been rounded to the nearest 1000 t.


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Table ES-2: Cuffley lode Mineral Resource on 30 June 2013


Indicated 133,000 16.9 5.4 29.0

Inferred 273,000 10.4 3.2 17.4

Notes to Table ES-2: 1. Mineral Resource stated as of 30 June 2013. 2. Mineral Resource used relevant sample data available as of 30 June 2013. 3. Mineral Resource stated according to CIM guidelines. 4. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. 5. A 3.6 g/t AuEq cut-off grade is applied where AuEq is calculated at a gold price of USD1,400 per troy ounce and an


Mining Methods Mandalay plans to mine the extension of the Augusta Underground Mineral Resources to the 856 mRL and the Cuffley lode, refer Figure ES-1.

Figure ES 1: Schematic of August and Proposed Cuffley Underground Design

The mine incorporating the Augusta Lodes and Cuffley lode is an underground antimony-gold mine currently producing approximately 100 ktpa from a variety of mining methods to a depth of over 310 m below surface. The peak trucking rate to achieve the production schedule proposed is 108 ktpa and this has been determined to be well within the capability of the current trucking fleet.

The Mine is serviced by a decline haulage system developed from a portal within a box-cut with dimensions of 4.8 m high by 4.5 m wide at a gradient of 1:7 down. The Mine employs predominantly air leg longhole stoping methods as well as longitudinal uphole retreat working a bottom up sequence. These mining methods have been utilised throughout 2010, 2011 and 2012. CRF is placed into stoping voids to maximise extraction and assist with mine stability.

Mandalay proposes to develop the Cuffley lode whose southern Inferred Resource boundary is located approximately 250 m to the northwest of the Augusta Mine. Access to the Cuffley lode will be via a single decline that will connect to the existing Augusta decline at the 1030 mRL. The Cuffley lode is scheduled to produce, in conjunction with the Augusta Mine, at a maximum rate of 108 ktpa. Three different mining methods; full face development, longhole air leg CRF stoping and half upper air leg stoping will be utilised. All mined material will be transported to the Augusta box-cut before being hauled to either the Brunswick ROM Pad or Augusta Waste Rock Storage

North


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Facility. The mining operation will continue to utilise proven methods and equipment that have been in use at the operation for the past four years.

The combined Costerfield production schedule presented in Figure ES-2, shows the reduction in production from the Augusta deposit as the Cuffley lode is developed. A breakdown of the proposed mill feed, at a cut-off grade of 4.7 g/t AuEq, from August and Cuffley is presented in Table ES-3. The proposed schedule is subject to further confirmatory work (ie drilling).

Table ES-3: Costerfield Proposed Mill Feed

Proposed Mill Feed (tonnes)

Gold Grade (g/t)

Antimony Grade (%)

Inferred Material (%)

Augusta 163,000 7.1 3.7 28

Cuffley 345,000 12.1 3.8 64

Costerfield 509,000 10.5 3.8 55

Note: Proposed Mill Feed rounded to the nearest thousands (rounding error in summations)

Figure ES-2: Costerfield Production Schedule (Proposed Mill Feed)

Cuffley Mine Design

The proposed mining schedule follows a bottom-up sequence, mining from the northern and southern extents retreating toward the central access. This sequence enables a consistent production profile to be maintained as it allows for dual development headings on each level. The current Augusta mining methods and extraction sequence which is indicative of the development and stoping sequence that is planned at Cuffley. Refer Figure ES-3.


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Figure ES-3: Isometric of Cuffley Development

Metallurgy and Recovery Methods All proposed mill feed is planned to be processed at the existing Brunswick Processing Plant at Costerfield, near 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.

Ore from the Augusta underground mine 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 the last six years 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 April 2013 has increased, with the mobile crushing plant increasing throughput, has resulted in current recoveries of 96% for antimony and 89% for gold.

Mineralogical and metallurgical testwork 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. The calculated gold recovery for the Cuffley mineralisation is estimated to be 2.6% greater than the 89% achieved for the Augusta ore, resulting in approximately 92%.

For the purposes of this Technical Report, the historical metallurgical recoveries achieved from processing ore for the Augusta Mine have been used, 96% for antimony and 89% for gold.

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

North


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Trials with a portable crushing plant during September 2012 were completed to assess the impact it would have on the throughput of the existing processing plant. Results proved that this additional crushing capacity was able to improve throughput to 10,000 t/mth. A larger capacity unit (Finlay Terex) was employed on a hire basis and eventually purchased, and has been successful in increasing the throughput tonnage. The current mill capacity is capable of handling greater than 10,000 t/mth, which was achieved from October 2012 to December 2012. From January 2013 to April 2013 the processed tonnes were less than 10,000 tonnes a month due to the availability of mineralised material from the mine.

Project Infrastructure The Costerfield Operations surface facilities are representative of a modern antimony-gold mining operation. The Augusta Mine site location, refer to Figure ES-4, comprises the Infrastructure listed below. The proposal Cuffley lode development is established from within the Augusta mine workings, and comprises:

• 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 Raise;

• Ventilation Intake Raise; and

• Water storage dam to manage rainfall run-off and mine dewatering.

The Brunswick site, refer to Figure ES-5, comprises of the following:

• Antimony-gold processing plant and associated facilities;

• Central administration complex;

• Process plant workshop;

• Tailings storage facilities;

• ROM stockpiles;

• Previously mined Brunswick Open Pit; and

• Core farm and core processing facility.


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Figure ES-4: Augusta Mine Site


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Figure ES-5: Brunswick Site Area

Underground Infrastructure The current underground infrastructure consists of the mine workings associated with the Augusta Mine accessed via the box-cut, all associated underground dewatering system and underground ventilation fans.

Services All power is provided directly from the national electricity grid. The chosen load growth scenario has indicated that an upgrade to the Powercor distribution infrastructure will be required to meet a maximum forecasted demand growth to 3,200 kVA at Augusta/ Cuffley and 1,400 kVA at Brunswick. This increase in demand is based on extending the Augusta Mine to the 860 mRL, accessing and mining the current known extents of the Cuffley lode Inferred/ Indicated Resource as well as expansion the Brunswick Processing Plant.


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The upgrade will result in the existing supply point for Augusta remaining unchanged, two new 22 kV supply points being installed at Brunswick and Cuffley. The existing low voltage supply point at Brunswick will be removed.

Powercor has provided a budget estimate of AUD5.7M for the high voltage upgrade. It has been assumed that the 22 kV upgrade will not be commissioned until December 2014. In the meantime, any shortfalls in power will require the use of diesel generated power until the high voltage upgrade is complete.

Contracts At present, there is an agreement in place between Mandalay and the Hunan Zhongan Antimony and Tungsten Trading Co Ltd for the sale of the antimony-gold concentrate produced from Augusta. The agreement is valid until the end of December 2014. Mandalay receives payment based on the concentration of the antimony and gold within the concentrate.

Transportation from Melbourne to the smelter in Chenzhou, shipment documentation, freight administration and assay exchange/ returns are conducted by Minalysis Pty Ltd. The marketing of the concentrate is conducted through Penfold Marketing Pty Ltd who markets the product to the customer and provides market information to Mandalay.

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 establishment of new ventilation shaft for Cuffley and the construction of new groundwater evaporation facilities will require a Work Plan Variation (WPV) to be approved. The DPI facilitates this approval process and will engage with relevant referral authorities, as required. The DPI 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 the regulatory authorities, and any conditions or proposed amendments requested to the WPV are generally negotiated to the satisfaction of both parties.

Capital and Operating Costs All cost estimates are estimates and based on 2013 Australian dollars (AUD). 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 has been estimated using the AUD:USD exchange rate of 0.90.


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This preliminary economic assessment is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves. There is no certainty that the preliminary economic assessment will be realised.

The total capital required for the Costerfield Operation is summarised in Table ES-4.

Table ES-4: Costerfield Operation – Capital Cost Estimate

Item Cost (AUD M)

Cost (USD M)

Sub-total Plant 3.7 3.3 Sub-total Admin 0.4 0.3 Sub-total Enviro 4.5 4.0 Sub-total OH&S 0.0 0.0 Sub-total Ops Geology 1.3 1.2 Sub-total Exploration 0.0 0.0 Sub-total Mining 3.2 5.5 Total - Plant and Equipment 16.1 14.5 Capital Development 15.2 13.7 Total Capital Cost estimate 31.3 28.1

The Life of Mine (LoM) operating costs will 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-5.

Table ES-5: Costerfield Operation – Operating Cost Estimate

Description Operating Cost

AUD M AUD/ t USD M USD / t

Mining 122 240 110 216 Processing 31 60 27 54 Site Services 9 17 8 15 General and Administration 21 40 19 36 Total 182 358 164 322

Million dollars rounded to nearest million “/ t” rounded to the nearest dollar

Offsite Costs

Concentrate selling costs of AUD195/ dmt of concentrate has been included to cover third party transport and shipping costs as well as associated marketing and broker costs. The concentrate selling costs have been estimated using Mandalay historical data from 2012 inclusive.

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.


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Economic Analysis The key project criteria and assumptions used in preparation of the cash flow analysis have been listed in Table ES-6. A summary of the project economics are presented in Table ES-7. Figure ES-6 shows the monthly and cumulative cashflow.

The results are preliminary in nature and include inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realised.

The NPV shown in Table ES-7 is calculated after tax. The after tax estimate is based on a company tax rate of 30% company tax rate, straight line depreciation, an opening book value of AUD43M and AUD42M of tax losses carried forward.

Table ES-6: Project Criteria

Description Units Quantity

Proposed Mill Feed

Tonnes 509,000

Gold grade (g/t) 10.5

Antimony grade (%) 3.8

Project Life months 47

Average Production Rate t/mth 10,800

Maximum Mining Rate t/mth 13,000

Metallurgical Recovery Gold (%) 89

Antimony (%) 96

Gravity Gold % 34

Concentrate Grade Gold (g/t) 71

Antimony (%) 53

Concentrate Selling Expenses AUD/dmt 195

Payable

Gold in Con (%) 78.5

Gravity Gold (%) 98

Antimony (%) 63

Exchange Rate AUD:USD 0.90

Commodity Prices Gold USD/oz 1,300

Antimony USD/t 9,500


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Table ES-7: Project Economics

Description Units Quantity Units Quantity

Tonnes Milled Tonnes 509,000 Tonnes 509,000

Recovered Gold Ounces 152,000 Ounces 152,000

Recovered Antimony Tonnes 18,000 Tonnes 18,000

Operating cost AUD M 182 USD M 164

Operating Cost per Payable ounce AUD / Oz Eq1 851 USD/oz eq 766

Capital cost AUD M 31 USD M 28

Total Cost (Operating + Capital) AUD M 213 USD M 192

Total Cost per Payable Ounce AUD / Oz Eq 998 USD/oz eq 898

Payable Gold Ounces 129,000 Ounces 129,000

Payable Antimony Tonnes 12,000 Tonnes 12,000

Payable (Saleable) Metal – Au Eq Oz Eq 214,000 Tonnes 214,000

Net Revenue (less selling expenses) AUD M 299 USD M 269

After Tax Profit AUD M 84 USD M 76

After Tax NPV5 AUD M 74 USD M 67

IRR % 3309 % 3309

Max Negative Cashflow AUD M -2 USD M -2

Max Negative Cashflow Mth Mar 2014 Mth Mar 2014 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-6: Undiscounted AUD Before-tax Cashflow Profile

The economic model has assumed an exchange rate of AUD:USD 0.90 for the entire project life. SRK notes that over the period that this Technical Report was prepared the exchange rate has fallen and is trading at a new level. As will be discussed in the sensitivity section the project financials are most sensitive to a change in exchange rate. A chart showing the project sensitivity is presented in

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Figure ES-7. 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.

A change in the AUD:USD exchange rate of 0.05 results in a revenue change of AUD16M and an NPV change of AUD14M.

Figure ES-7: Sensitivity Analysis

Adjacent Properties The Augusta Mining Lease (MIN4644) is completely enveloped by exploration leases held by Mandalay (through AGD). In the immediate area of the Augusta Mine there are no advanced projects, nor are there any other Augusta-style antimony-gold operations in production in the Costerfield district. Exploration on adjacent prospects (EL3316, EL5310 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 1049 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.


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Interpretation and Conclusions The drilling, development and mining of the Costerfield operations has demonstrated continuity of grade, mineralisation and geologic structure to support the definition of a reasonable prospect of economic extraction defined by CIM standards for the Augusta and Cuffley Measured, Indicated and Inferred Mineral Resource classifications.

The high level PEA conducted for the combined Mineral Resources of the Augusta Deposit and Cuffley lode indicate the potential for a robust business case based on fair and reasonable production and cost assumptions. Ongoing geological drilling and development of these deposits and advancement of the engineering studies to develop a more detailed LoM Plan is considered to be appropriate.

Further advancement of geological drilling to better define the Mineral Resource in conjunction with revisions to the mine design and mining schedule to better smooth the development and production profile is warranted.

Comments and Recommendations

The overall positive results of this Technical Report indicate the technical and financial feasibility of the project, indicating that additional study work is warranted to advance the project.

It is considered there is opportunity to improve the cashflow profile by undertaking the following actions:

• Review the Mineral Resource to assess potential to identify additional tonnes closer to Augusta;

• Undertaking more detailed scheduling work to smooth the development and production profiles; and

• Continue to refine and review mining methods and capital and operating cost estimates.


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Table of Contents

Important Notice ........................................................................................................................................... ii

Executive Summary .................................................................................................................................... iv

2 Introduction and Terms of Reference ......................................................................... 1

2.1 Scope of Work ..................................................................................................................................... 1

2.1.1 Mineral Resource Report – Costerfield Operation Deposits ................................................... 2 2.2 Work Programme ................................................................................................................................ 2

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

2.4 Qualifications of SRK and SRK Team ................................................................................................ 3

2.5 Acknowledgement ............................................................................................................................... 5

2.6 Declaration .......................................................................................................................................... 5

3 Reliance on Other Experts ........................................................................................... 6

3.1 Marketing ............................................................................................................................................ 6

4 Property Description and Location ............................................................................. 7

4.1 Property Location ................................................................................................................................ 7

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

4.3 Underlying Agreements .................................................................................................................... 11

4.4 Environmental Liability ...................................................................................................................... 11

4.5 Royalties ........................................................................................................................................... 11 4.6 Taxes ................................................................................................................................................ 11

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

5 Accessibility, Climate, Local Resources, Infrastructure and Physiography ......... 13

5.1 Accessibility ....................................................................................................................................... 13

5.2 Land Use ........................................................................................................................................... 13

5.3 Topography ....................................................................................................................................... 13

5.4 Climate .............................................................................................................................................. 13 5.5 Infrastructure and Local Resources .................................................................................................. 14

6 History ......................................................................................................................... 17

6.1 Introduction ....................................................................................................................................... 17

6.2 Ownership and Exploration Work ..................................................................................................... 17

6.2.1 Mid East Minerals (1968 – 1971) .......................................................................................... 17

6.2.2 Metals Investment Holdings (1971) ....................................................................................... 18

6.2.3 Victorian Mines Department (1975 – 1981) .......................................................................... 18 6.2.4 Federation Resources NL (1983 – 2000) .............................................................................. 18

6.2.5 Australian Gold Development NL/ Planet resource JV (AGD) (1987 – 1988) ...................... 18

6.2.6 Australian Gold Development NL (AGD) (1987 – 1997) ....................................................... 18

6.2.7 AGD Operations Pty Ltd (2001 – 2009) ................................................................................ 18

7 Geological Setting and Mineralisation ...................................................................... 22


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7.1 Regional Geology.............................................................................................................................. 22

7.2 Property Geology .............................................................................................................................. 24 7.3 Property Mineralisation ..................................................................................................................... 26

7.4 Deposit Mineralisation ....................................................................................................................... 27

8 Deposit Types ............................................................................................................. 29

9 Exploration .................................................................................................................. 30

9.1 Costean/ Trenching ........................................................................................................................... 30

9.2 Petrophysical Analysis ...................................................................................................................... 30

9.3 Geophysics ....................................................................................................................................... 30

9.3.1 Ground Geophysics ............................................................................................................... 30 9.3.2 Airborne Geophysics ............................................................................................................. 31

9.4 Geochemistry .................................................................................................................................... 31

9.4.1 Mobile Metal Ion (MMI) .......................................................................................................... 31

9.4.2 Bedrock Geochemistry .......................................................................................................... 31

9.5 Aerial Photogrammetry Survey ......................................................................................................... 35

9.6 Underground Face Sampling ............................................................................................................ 35

10 Drilling ......................................................................................................................... 36

10.1 Mandalay Resources (2009 – Present) ............................................................................................ 36

10.2 2009/ 2010 ........................................................................................................................................ 36

10.3 2010/ 2011 ........................................................................................................................................ 36

10.4 2011/ 2012 ........................................................................................................................................ 36

10.5 2013 Cuffley lode Drilling .................................................................................................................. 37 10.6 Drilling Methods ................................................................................................................................ 38

10.7 Collar Surveys ................................................................................................................................... 38

10.8 Downhole Surveys ............................................................................................................................ 40

10.9 Logging Procedures .......................................................................................................................... 40

10.10 Drilling Pattern and Quality ............................................................................................................... 41

10.10.1 Augusta ............................................................................................................................. 41 10.10.2 Cuffley ............................................................................................................................... 41

10.11 Interpretation of Drilling Results ........................................................................................................ 41

10.12 Factors that could Materially Impact the Accuracy of Results .......................................................... 42

11 Sample Preparation, Analyses and Security ............................................................ 44

11.1 Sampling Techniques........................................................................................................................ 44

11.1.1 Diamond Core Sampling ....................................................................................................... 44

11.2 Data Spacing and Distribution .......................................................................................................... 44 11.3 Testing Laboratories ......................................................................................................................... 44

11.4 Sample Preparation .......................................................................................................................... 45

11.5 Sample Analysis ................................................................................................................................ 45

11.6 Laboratory Reviews .......................................................................................................................... 46

11.7 Assay Quality Assurance and Quality Control .................................................................................. 46


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11.7.1 Standard Reference Material ................................................................................................ 46

11.7.2 Blank Material ........................................................................................................................ 49 11.7.3 Duplicate Assay Statistics ..................................................................................................... 50

11.7.4 Check Assay Program – Sample Pulps ................................................................................ 51

11.8 Sample Transport and Security ........................................................................................................ 54

11.9 Conclusions ....................................................................................................................................... 54

12 Data Verification ......................................................................................................... 56

13 Mineral Processing and Metallurgical Testing ......................................................... 57

13.1 Metallurgical Testing – Augusta ........................................................................................................ 57

13.1.1 Historical Testwork ................................................................................................................ 57 13.1.2 Recent Testwork ................................................................................................................... 60

13.1.3 Test Samples ........................................................................................................................ 62

13.1.4 Cuffley Lode Mineralisation ................................................................................................... 63

13.2 Stibnite Liberation ............................................................................................................................. 64

13.3 Gold Liberation .................................................................................................................................. 65

13.3.1 Cuffley Lode Hardness Comparison ..................................................................................... 66 13.3.2 Gravity and Flotation Recovery of the Cuffley Lode ............................................................. 67

13.4 Recovery Data on Augusta Orebody ................................................................................................ 68

13.4.1 Mineralogical Study ............................................................................................................... 68

13.5 Mineral Processing............................................................................................................................ 68

14 Mineral Resource Estimates ...................................................................................... 70

14.1 Introduction ....................................................................................................................................... 70 14.2 Diamond Drillhole and Underground Face Sample Statistics ........................................................... 70

14.3 Grade Capping .................................................................................................................................. 71

14.4 Data Interpretation and Compositing ................................................................................................ 72

14.5 Dip – Dip Direction Domains ............................................................................................................. 75

14.6 Bulk Density Determinations ............................................................................................................. 79

14.7 Variography ....................................................................................................................................... 80 14.8 Estimation Domain Boundaries ......................................................................................................... 81

14.9 Block Model Estimation ..................................................................................................................... 82

14.10 Estimation Parameters ...................................................................................................................... 83

14.11 Block Model Validation ...................................................................................................................... 85

14.12 Mineral Resource Classification ........................................................................................................ 91

14.13 Mineral Resource .............................................................................................................................. 91 14.14 Cut-off Grade Calculations ................................................................................................................ 98

14.15 Reconciliation .................................................................................................................................... 98

14.16 Comparison with Previous Mineral Resource Estimate .................................................................. 100

14.17 Other Material Factors .................................................................................................................... 100

15 Mineral Reserve Estimates ...................................................................................... 102

16 Mining Methods ........................................................................................................ 103


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16.1 Introduction ..................................................................................................................................... 103

16.2 Geotechnical ................................................................................................................................... 105 16.2.1 Augusta Overview ............................................................................................................... 105

16.2.2 Cuffley Overview ................................................................................................................. 105

16.2.3 Rock stress .......................................................................................................................... 107

16.2.4 Stope Design Criteria .......................................................................................................... 107

16.2.5 Ground Support Requirements ........................................................................................... 108

16.2.6 Decline Location .................................................................................................................. 109 16.2.7 References .......................................................................................................................... 109

16.3 Augusta Mine Design ...................................................................................................................... 110

16.3.1 Method Selection ................................................................................................................. 110

16.3.2 Method Description ............................................................................................................. 110

16.3.3 Materials Handling ............................................................................................................... 113

16.4 Augusta Mine Design Guidelines .................................................................................................... 113 16.4.1 Design Parameters .............................................................................................................. 113

16.4.2 Mining Sequence ................................................................................................................. 114

16.4.3 Decline Development .......................................................................................................... 115

16.4.4 Level Development .............................................................................................................. 115

16.4.5 Vertical Development .......................................................................................................... 116

16.4.6 Augusta Design Inventory ................................................................................................... 117 16.5 Cuffley Mine Design ........................................................................................................................ 117

16.5.1 Mine Fill ............................................................................................................................... 118

16.5.2 Ventilation Shaft .................................................................................................................. 118

16.5.3 Additional Geotechnical Work ............................................................................................. 118

16.6 Cuffley Mine Design Guidelines ...................................................................................................... 118 16.6.1 Design Parameters .............................................................................................................. 118

16.6.2 Mining Sequence ................................................................................................................. 118

16.6.3 Decline Development .......................................................................................................... 119

16.6.4 Level Development .............................................................................................................. 119

16.6.5 Vertical Development .......................................................................................................... 119

16.6.6 Stoping ................................................................................................................................ 120 16.6.7 Materials Handling ............................................................................................................... 120

16.6.8 Cuffley Proposed Design Inventory ..................................................................................... 121

16.7 Ventilation ....................................................................................................................................... 121

16.7.1 Existing Circuit ..................................................................................................................... 121

16.7.2 Proposed Ventilation Circuit ................................................................................................ 122

16.7.3 Airflow Requirements .......................................................................................................... 123 16.7.4 Cuffley LoM Ventilation ....................................................................................................... 124

16.7.5 Rise Sizes ........................................................................................................................... 124

16.7.6 Fan Duties ........................................................................................................................... 124

16.7.7 Raise Sizes ......................................................................................................................... 125


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16.7.8 References .......................................................................................................................... 125

16.8 Backfill 125 16.9 Mine Services and Infrastructure .................................................................................................... 126

16.9.1 Electrical and Communications ........................................................................................... 126

16.9.2 Compressed Air ................................................................................................................... 126

16.9.3 Process and Potable Water................................................................................................. 127

16.9.4 Explosives and Magazine.................................................................................................... 127

16.10 Augusta Hydrogeology / Dewatering .............................................................................................. 128 16.10.1 Hydrogeology .................................................................................................................. 128

16.10.2 Dewatering ...................................................................................................................... 129

16.11 Cuffley Hydrogeology / Dewatering ................................................................................................ 130

16.11.1 Hydrogeology .................................................................................................................. 130

16.11.2 Dewatering ...................................................................................................................... 131

16.12 Modifying Factors ............................................................................................................................ 131 16.12.1 Mining Dilution and Recovery ......................................................................................... 131

16.12.2 Cut-off Grade .................................................................................................................. 133

16.13 Life of Mine Schedule...................................................................................................................... 133

16.13.1 Development Schedule ................................................................................................... 133

16.13.2 Production Profile ............................................................................................................ 134

16.13.3 Equipment Requirements ............................................................................................... 135 16.13.4 Personnel ........................................................................................................................ 136

16.13.5 Labour Costs ................................................................................................................... 137

17 Recovery Methods .................................................................................................... 138

17.1 Brunswick Processing Plant ............................................................................................................ 138

17.1.1 Crushing and Screening Circuit .......................................................................................... 138

17.1.2 Milling Circuit ....................................................................................................................... 138 17.1.3 Flotation Circuit ................................................................................................................... 139

17.1.4 Concentrate Thickening Filtration ....................................................................................... 139

17.1.5 Tailings and Circuit .............................................................................................................. 139

17.1.6 Recovery ............................................................................................................................. 139

17.1.7 Concentrate Grade .............................................................................................................. 139

17.2 Plant Upgrade ................................................................................................................................. 142 17.2.1 Crushing and Screening Circuit .......................................................................................... 142

17.2.2 Milling Circuit ....................................................................................................................... 142

17.2.3 Flotation Circuit ................................................................................................................... 142

17.2.4 Concentrate Thickening and Filtration ................................................................................ 143

17.2.5 Tailings Circuit ..................................................................................................................... 143

17.2.6 Reagent Mixing and Storage ............................................................................................... 143 17.2.7 Recovery ............................................................................................................................. 143

17.2.8 Services ............................................................................................................................... 143


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18 Project Infrastructure ............................................................................................... 146

18.1 Surface Infrastructure...................................................................................................................... 146

18.2 Augusta Mine Underground Infrastructure ...................................................................................... 148 18.3 Tailings Storage .............................................................................................................................. 148

18.4 Power Supply .................................................................................................................................. 148

18.5 Power Reticulation .......................................................................................................................... 149

18.6 Water Supply ................................................................................................................................... 150

18.7 Waste Rock Storage ....................................................................................................................... 150

18.8 Water Management......................................................................................................................... 151 18.9 Augusta to Brunswick ROM Pad Transport .................................................................................... 151

18.10 Diesel Storage ................................................................................................................................. 151

18.11 Explosive Storage ........................................................................................................................... 152

18.12 Maintenance Facilities .................................................................................................................... 152

18.13 Housing and Land ........................................................................................................................... 152

19 Market Studies and Contracts ................................................................................. 153

19.1 Concentrate Transport .................................................................................................................... 153 19.2 Marketing ........................................................................................................................................ 153

19.3 Contracts ......................................................................................................................................... 155

20 Environmental Studies, Permitting, and Social or Community Impact ............... 156

20.1 Environment and Social Aspects .................................................................................................... 156

20.1.1 Increased Processing Rates ............................................................................................... 156

20.1.2 Mine Ventilation ................................................................................................................... 156 20.1.3 Water Disposal .................................................................................................................... 156

20.1.4 Waste Rock ......................................................................................................................... 157

20.1.5 Tailings Disposal ................................................................................................................. 158

20.2 Impacts ............................................................................................................................................ 158

20.2.1 Air Quality ............................................................................................................................ 158

20.2.2 Groundwater ........................................................................................................................ 159 20.2.3 Noise ................................................................................................................................... 160

20.2.4 Blasting and Vibration ......................................................................................................... 160

20.2.5 Native Vegetation ................................................................................................................ 160

20.2.6 Visual Amenity ..................................................................................................................... 160

20.2.7 Heritage ............................................................................................................................... 161

20.2.8 Community .......................................................................................................................... 161 20.2.9 Mine Closure and Revegetation .......................................................................................... 162

20.3 Regulatory Approvals ...................................................................................................................... 162

20.3.1 Work Plan Variation ............................................................................................................. 162

20.3.2 Other Permitting .................................................................................................................. 162

20.4 References ...................................................................................................................................... 163

21 Capital and Operating Costs ................................................................................... 164


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21.1 Capital Costs ................................................................................................................................... 164

21.1.1 Capital Lateral Development ............................................................................................... 164 21.1.2 Vertical Development .......................................................................................................... 165

21.1.3 Infrastructure ....................................................................................................................... 165

21.1.4 Mobile Plant ......................................................................................................................... 165

21.1.5 Processing Plant ................................................................................................................. 165

21.1.6 Drilling.................................................................................................................................. 166

21.1.7 Closure ................................................................................................................................ 166 21.2 Operating Costs .............................................................................................................................. 166

21.2.1 Lateral Development ........................................................................................................... 167

21.2.2 Production Stoping .............................................................................................................. 167

21.2.3 Augusta to Brunswick ROM Pad trucking ........................................................................... 167

21.2.4 Processing ........................................................................................................................... 167

21.3 Concentrate Selling Expenses ........................................................................................................ 168 21.4 Royalties and Compensation .......................................................................................................... 168

22 Economic Analysis ................................................................................................... 169

22.1 Introduction ..................................................................................................................................... 169

22.2 Principal Assumptions ..................................................................................................................... 169

22.2.1 Metal Sale Prices ................................................................................................................ 169

22.2.2 Concentrate Sales ............................................................................................................... 170

22.2.3 Exchange Rate .................................................................................................................... 170 22.2.4 Taxes ................................................................................................................................... 170

22.2.5 Royalties/ Agreements ........................................................................................................ 171

22.2.6 Reclamation ........................................................................................................................ 171

22.2.7 Project Financing ................................................................................................................. 171

22.3 Economic Summary ........................................................................................................................ 171 22.4 Sensitivities ..................................................................................................................................... 172

22.5 After Tax Cashflow .......................................................................................................................... 173

22.6 Cash Flow Forecast ........................................................................................................................ 175

23 Adjacent Properties .................................................................................................. 179

23.1 General Statement about Adjacent Properties ............................................................................... 179

24 Other Relevant Data and Information ..................................................................... 181

24.1 Remnant Mining .............................................................................................................................. 181

25 Interpretation and Conclusions............................................................................... 182

25.1 Geology ........................................................................................................................................... 182

26 Comments and Recommendations ......................................................................... 184

26.1 Geology ........................................................................................................................................... 184

26.2 Mining 184

27 References ................................................................................................................ 185


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

Table 4-1: Granted Tenement Details ........................................................................................................... 8

Table 6-1: Historical Drilling Statistics for the Costerfield Property ............................................................. 17

Table 10-1: Drillhole Summary ...................................................................................................................... 36

Table 10-2: Significant Cuffley lode Drillhole Intercepts: January to July 2013 ............................................ 38 Table 11-1: Summary of Onsite Duplicate Statistics ..................................................................................... 50

Table 11-2: Summary of Onsite versus ALS Gold Duplicate Statistics ......................................................... 52

Table 11-3: Summary of Onsite versus ALS Antimony Duplicate Statistics ................................................. 52

Table 13-1: Ore Characterisation testwork vs. plant reconciled recovery results for Antimony.................... 58

Table 13-2: Ore Characterisation testwork vs. plant reconciled recovery results for Gold ........................... 58

Table 13-3: Ore Characterisation testwork size range tested per test .......................................................... 60 Table 13-4: Testwork Schedule and Grades ................................................................................................. 60

Table 13-5: Amdel Qemscan Stibnite Liberation Table................................................................................. 64

Table 13-6: Amdel Qemscan Gold Liberation Table ..................................................................................... 65

Table 13-7: Antimony Recovery Comparison Cuffley Lode vs. Augusta Orebody ....................................... 68

Table 13-8: Gold Recovery Comparison Cuffley Lode vs. Augusta Orebody ............................................... 68

Table 13-9: Plant Recovery Data on Existing Augusta Orebody .................................................................. 69 Table 14-1: Summary Statistics .................................................................................................................... 70

Table 14-2: Percentage of Samples Affected by 150 g/t Au Grade Cap ...................................................... 72

Table 14-3: Gold Statistics Before and After Grade Capping ....................................................................... 72

Table 14-4: Mineralisation Coding ................................................................................................................. 73

Table 14-5: Summary Composite Statistics .................................................................................................. 74 Table 14-6: E Lode Mean Dip and Dip Direction per Domain ....................................................................... 78

Table 14-7: W Lode Mean Dip and Dip Direction per Domain ...................................................................... 78

Table 14-8: NE Lode Mean Dip and Dip Direction per Domain .................................................................... 79

Table 14-9: Face Sample Fitted Model Variogram Parameters .................................................................... 81

Table 14-10: Face Sample Block Model Dimensions ..................................................................................... 83

Table 14-11: Drillhole Sample Block Model Dimensions ................................................................................ 83 Table 14-12: Mine Planning Regularised Block Model Dimensions ............................................................... 83

Table 14-13: Search Parameters .................................................................................................................... 84

Table 14-14: Augusta and Brunswick Mineral Resource Summary ................................................................ 93

Table 14-15: Cuffley Lode Mineral Resource on 30 June 2013 ...................................................................... 93

Table 14-16: E, W, NE and Cuffley lode Current Estimate Compared with the 31 December 2012 Mineral Resource Estimate ................................................................................................................... 101

Table 16-1: Costerfield Proposed Mill Feed ................................................................................................ 105

Table 16-2: Main Mine Design Parameters ................................................................................................. 108

Table 16-3: Mine Design Parameters.......................................................................................................... 113

Table 16-4: Augusta Design Inventory ........................................................................................................ 117

Table 16-5: Mine Design Parameters.......................................................................................................... 118


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Table 16-6: Cuffley Stope and Development Inventory .............................................................................. 121

Table 16-7: Proposed Cuffley Diesel Fleet and ventilation requirements ................................................... 122 Table 16-8: Augusta Primary Flow to Match Diesel Flow ........................................................................... 124

Table 16-9: Cuffley LoM Ventilation Operational Summary ........................................................................ 124

Table 16-10: Cuffley Rise sizes ..................................................................................................................... 124

Table 16-11: Current Augusta Licence Maximum Quantities and Types of Explosives ............................... 127

Table 16-12: Augusta Recovery and Dilution Assumptions .......................................................................... 132

Table 16-13: Jumbo Production Rates .......................................................................................................... 134 Table 16-14: Hand Held Production Rates ................................................................................................... 134

Table 16-15: Augusta Underground Mobile Equipment Fleet ....................................................................... 135

Table 16-16: Personnel on Payroll by Department ....................................................................................... 136

Table 16-17: Underground Shift Mining Personnel ....................................................................................... 137

Table 17-1: Brunswick Mill production results for Feb 2012 to Apr 2013 .................................................... 141

Table 17-2: Brunswick Mill concentrate production Feb 2012 to Apr 2013 ................................................ 141 Table 17-3: Current power requirements at Brunswick Processing Plant................................................... 144

Table 18-1: Current Augusta Licence Maximum Quantities and Types of Explosives ............................... 152

Table 20-1: Permit Requirements ............................................................................................................... 163

Table 21-1: Summary of Capital Costs ....................................................................................................... 164

Table 21-2: Operating Cost Inputs .............................................................................................................. 166

Table 21-3: Costerfield Operation - Operating Cost Summary ................................................................... 167 Table 22-1: Project Criteria ......................................................................................................................... 169

Table 22-2: Project Economics ................................................................................................................... 171

Table 22-3: Sensitivity at 5% NPV .............................................................................................................. 172

Table 22-4: Depreciation Schedule ............................................................................................................. 173

Table 22-5: After Tax Profit & Loss and Valuation (AUD) ........................................................................... 174 Table 22-6: After tax Profit & Loss and Valuation (USD) ............................................................................ 175

Table 22-7: Proposed Production Schedule ............................................................................................... 176

Table 22-8: Estimated Pre-Tax Cash Flow Summary in AUD .................................................................... 177

Table 22-9: Estimated Pre-Tax Cash Flow Summary in USD .................................................................... 178

Table 23-1: Augusta Mine Adjacent Properties (DPL, 2012) ...................................................................... 179

Table 23-2: Distance from the Augusta Mine Site to Significant Central Victoria Mining Landmarks ........ 180


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List of Figures Figure 4-1: Costerfield Operation Location .................................................................................................... 7

Figure 4-2: Plan of Area – Mining Licence No: 4644 ...................................................................................... 9

Figure 4-3: Current AGD Mining and Exploration Lease Boundaries (Mandalay, 2012) ............................. 10

Figure 5-1: Monthly Average Temperature and Rainfall .............................................................................. 14

Figure 5-2: Augusta Box-cut, Portal and Workshop ..................................................................................... 15 Figure 5-3: Aerial View of Brunswick Processing Plant and Brunswick Open Pit ........................................ 15

Figure 7-1: Geological Map of the Heathcote – Colbinabbin – Nagambie Region ...................................... 23

Figure 7-2: Costerfield Property Geology and Old Workings ....................................................................... 25

Figure 7-3: Two View of E Lode 1070 Level South, with Annotation on Right-Hand side ........................... 26

Figure 7-4: Close-up of Mineralisation in E Lode 1070 Level South ............................................................ 27

Figure 7-5: Augusta and Cuffley Schematic Geological Cross Section at 4350N, Costerfield Mine Grid ... 28 Figure 9-1: Augusta South Aircore Programme Results with Lode Locations ............................................. 32

Figure 9-2: Augusta South (Margaret Zone) – Auger Drill 2011 Bedrock Geochemistry: Antimony Contours33

Figure 9-3: Auger Geochemistry Results Displayed as Antimony Contours ................................................ 34

Figure 10-1: Collar Locations near Augusta and Cuffley lodes ...................................................................... 39

Figure 10-2: Cross Section at 4430N through the Augusta Deposit .............................................................. 42

Figure 11-1: G901-8 Gold Standard Reference Material – Assay Results January 2012 to December 201247 Figure 11-2: G907-6 Gold Standard Reference Material – Assay Results January 2012 to December 201248

Figure 11-3: AGD08-01 Antimony Standard Reference Material – Assay Results January 2012 to December 2012 .......................................................................................................................... 48

Figure 11-4: AGD08-02 Antimony Standard Reference Material – Assay Results January 2012 to December 2012 .......................................................................................................................... 49

Figure 11-5: Blank Gold Results for Onsite .................................................................................................... 49

Figure 11-6: Scatter Plot for Onsite Gold Duplicates (g/t) .............................................................................. 50

Figure 11-7: Relative Paired Difference Plot for Onsite Gold Duplicates (g/t) ............................................... 51

Figure 11-8: Scatter Plot for Onsite versus ALS Gold Duplicates (g/t) .......................................................... 52

Figure 11-9: Scatter Plot for Onsite versus ALS Antimony Duplicates (%) .................................................... 53

Figure 11-10: Relative Paired Difference Plot for Onsite versus ALS Gold Duplicates (g/t) ............................ 53 Figure 11-11: Relative Paired Difference Plot for Onsite versus ALS Antimony Duplicates (%) ..................... 54

Figure 13-1: Ore Characterisation testwork (M2610) vs plant reconciled recovery ....................................... 59

Figure 13-2: Pre-Characterisation testwork flow sheet .................................................................................. 61

Figure 13-3: Longitudinal Section of the Cuffley Deposit intercepts as at January 2013 ............................... 62

Figure 13-4: Longitudinal Section of the Cuffley Deposit intercepts as at January 2013 ............................... 63

Figure 13-5: Amdel Qemscan Mineral Abundance ........................................................................................ 64 Figure 13-6: Amdel Qemscan Stibnite Liberation Visual Representation ...................................................... 65

Figure 13-7: Amdel Qemscan Gold Liberation Visual Representation ........................................................... 66

Figure 13-8: Grind Establishment Testwork Comparisons at 75 µm.............................................................. 66

Figure 13-9: Grind Establishment Testwork Comparisons at 53 µm.............................................................. 67

Figure 14-1: Log Probability Plot of Gold – Face Samples ............................................................................ 71

Figure 14-2: Log Probability Plot of Gold – Drillhole Samples ....................................................................... 71


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Figure 14-3: E Lode Dip – Dip Direction Domains Coloured by Dip .............................................................. 75

Figure 14-4: E Lode Dip – Dip Direction Domains Coloured by Dip Direction ............................................... 76 Figure 14-5: W Lode Dip – Dip Direction Domains Coloured by Dip ............................................................. 76

Figure 14-6: W Lode Dip – Dip Direction Domains Coloured by Dip Direction .............................................. 77

Figure 14-7: NE Lode Dip – Dip Direction Domains Coloured by Dip ............................................................ 77

Figure 14-8: NE Lode Dip – Dip Direction Domains Coloured by Dip Direction ............................................ 78

Figure 14-9: Bulk Density Determinations ...................................................................................................... 80

Figure 14-10: W Lode Estimation Domain Boundaries .................................................................................... 81 Figure 14-11: NE Lode Estimation Domain Boundaries Showing Three Domains .......................................... 82

Figure 14-12: Cuffley lode Estimation Domain Boundaries Showing Four Domains ....................................... 82

Figure 14-13: Swath Plot Comparing E Lode Face Samples with Face Sample Block Model Estimated True Thickness by Northing ................................................................................................................ 85

Figure 14-14: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated True Thickness by Elevation ..................................................................................... 86

Figure 14-15: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Northing ............................................................................................ 86

Figure 14-16: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Elevation ........................................................................................... 87

Figure 14-17: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Northing ..................................................................................... 87

Figure 14-18: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Elevation ................................................................................... 88

Figure 14-19: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated True Thickness by Northing ...................................................................................... 88

Figure 14-20: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated True Thickness by Elevation ..................................................................................... 89

Figure 14-21: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Northing ............................................................................................ 89

Figure 14-22: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Elevation ........................................................................................... 90

Figure 14-23: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Northing ..................................................................................... 90

Figure 14-24: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Elevation ................................................................................... 91

Figure 14-25: E Lode Mineral Resource Estimate Longitudinal Projection ...................................................... 94

Figure 14-26: W Lode Mineral Resource Estimate Longitudinal Projection ..................................................... 94

Figure 14-27: NE Lode Mineral Resource Estimate Longitudinal Projection ................................................... 95 Figure 14-28: NW Lode Mineral Resource Estimate Longitudinal Projection .................................................. 95

Figure 14-29: P Lode Mineral Resource Estimate Longitudinal Projection ...................................................... 96

Figure 14-30: Cuffley Lode Mineral Resource Estimate Longitudinal Projection ............................................. 96

Figure 14-31: C Lode 2009 Mineral Resource Estimate Longitudinal Projection ............................................ 97

Figure 14-32: Brunswick 2009 Mineral Resource Estimate Longitudinal Projection ........................................ 97 Figure 14-33: Mill Reconciliation – Dry Tonnes ................................................................................................ 98

Figure 14-34: Mill Reconciliation – Gold (g/t) ................................................................................................... 99

Figure 14-35: Mill Reconciliation – Antimony (%) ............................................................................................. 99


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Figure 16-1: Schematic of August and Proposed Cuffley Underground Design .......................................... 103

Figure 16-2: Combined Augusta and Cuffley Production Profile .................................................................. 104 Figure 16-3: Geotechnical Domains – Section 5010 mN ............................................................................. 106

Figure 16-4: Augusta Longhole CRF Stoping Method (after Potvin, Thomas, Fourie; 2005) ...................... 111

Figure 16-5: Augusta Half Upper Mining Method (after Marshall, 2010) ...................................................... 112

Figure 16-6: Augusta Flat Back Mining Method ........................................................................................... 113

Figure 16-7: Long section of Augusta stoping between 1015 mRL and 989 mRL ....................................... 114

Figure 16-8: Typical Plan View of Augusta Decline and Level Access Development .................................. 115 Figure 16-9: Typical Section Looking North through Augusta Level Development ...................................... 116

Figure 16-10: Isometric of Cuffley Development ............................................................................................ 119

Figure 16-11: Long Section of Cuffley stoping areas ..................................................................................... 120

Figure 16-12: Proposed Cuffley Ventilation Circuit ........................................................................................ 123

Figure 16-13: Conceptual cross section of the groundwater flow paths ........................................................ 128

Figure 16-14: Groundwater elevation contour map of the areas surrounding the Augusta Mine .................. 129 Figure 16-15: Capital Development ................................................................................................................ 133

Figure 16-16: Costerfield Production Profile (proposed Mill Feed) ................................................................ 135

Figure 17-1: Brunswick Processing Plant Flow Sheet .................................................................................. 140

Figure 17-2: Brunswick Processing Plant Proposed Upgrade Flow Sheet .................................................. 145

Figure 18-1: Augusta Mine Site .................................................................................................................... 146

Figure 18-2: Brunswick Site Area ................................................................................................................. 147 Figure 18-3: Augusta Waste Storage Forecast ............................................................................................ 150

Figure 19-1: World Antimony production and consumption, 2000 – 2011 (t Antimony) ............................... 153

Figure 19-2: World Antimony demand and supply forecast, 2000 – 2016 (t Antimony) ............................... 154

Figure 19-3: China trends in domestic FOB prices for Antimony metal and trioxide ................................... 155

Figure 20-1: Groundwater elevation contour map of the areas surrounding the Augusta Mine .................. 159 Figure 22-1: AUD:USD Exchange Rate Chart ............................................................................................. 170

Figure 22-2: Undiscounted Cashflow (AUD) ................................................................................................ 172

Figure 22-3: Sensitivity Analysis ................................................................................................................... 173

Figure 23-1: Augusta Mine Adjacent Properties ........................................................................................... 179


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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 DPI Department of Environment and Primary Industries DSE Department of Sustainability and Environment DTM Digital terrain model EM Electro-magnetic EPA Environmental Protection Agency EVC’s Ecological Vegetation Classes FAR Fresh air rise FAusIMM Fellow of the Australian Institute of Mining and Metallurgy Federation Federation Resources N.L GDip Graduate Diploma GEF Gold and Exploration Finance Company of Australia GMA/WMC Gold Mines of Australia/Western Mining Corporation GPS Global positioning system GST Goods and services tax g/t grams per tonne HBr Hydrobromic acid HCl Hydrochloric acid HR Hydraulic radius ICP - AES Inductively couple plasma atomic emission spectroscopy ID2 Inverse distance squared ID3 Inverse distance cubed IP Induced polarisation


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Abbreviation Meaning

IRR Internal rate of return JORC Joint Ore Reserves Committee kg kilogram kL Kilolitre km kilometre Koz Kilo ounces kt kilo tonne ktpa kilo tonnes per annum Ktpm Kilo tonnes per month kV kilo volt kVA kilo volt ampere kW kilo watt kWh kilo watt 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 Australian Institute of Mining and Metallurgy Metcon Metcon Laboratories MEM Mid-East Minerals N.L mg/kg Milligram 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 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


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Abbreviation Meaning

PEA Preliminary Economic Assessment PhD Doctor of Philosophy Planet Planet Resources Group N.L 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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2 Introduction and Terms of Reference SRK Consulting (Australasia) Pty Ltd (SRK) has prepared the Costerfield Operation Preliminary Economic Assessment (PEA) to re-evaluate the viability of mining and processing mineralisation from continuing operations at Augusta Mine in connection with the addition of possible mineralisation sourced from the newly discovered Cuffley Lode which revised production scenario may have a material impact on the proposed production scenario disclosed in SRK Consulting (Australasia) Pty Ltd. March 2013, NI 43-101 report filed March 28, 2013.

The Costerfield Operation is located within the Costerfield mining district, approximately 10 km north east 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. The Cuffley lode is planned to supplement production from the current Augusta Mine and is located approximately 500 m to the north of the existing Augusta Mine workings.

The Costerfield Operation is contained within Mining Lease MIN4644 and comprises the following:

• Underground mine (Augusta) with production from three mineralised structures, E, W and N Lode;

• A conventional flotation mill (Brunswick) with a current name plate capacity of approximately 85,000 t/year; 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.

This report has been prepared a combination of publicly available and confidential information. Assistance and guidance has been provided by a number of independent consultants.

SRK was engaged to prepare the consolidated PEA for the Costerfield Operation, including a review of mining and processing operation and infrastructure requirements.

AMC Consultants Pty Ltd (AMC) was engaged to prepare a Mineral Resource Report for the Augusta deposit and Cuffley lode.

Consultants from AMC will act as Qualified Persons (QP) for reporting of the Mineral Resources.

The documentation reviewed, and other sources of information, are listed at the end of this report in Section 27, References.

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 on 15 May 2013 and subsequent discussions between Mandalay and SRK, includes the review of documents provided by Mandalay and prepared documentation as required for the PEA Technical Report on the Costerfield Operation in compliance with NI 43-101 and Form 43-101F1 guidelines.



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

• Review and comment of the mining method and mining schedules for Augusta;

• Update the mine design and prepare mining schedules for Cuffley lode;

• 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.

2.1.1 Mineral Resource Report – Costerfield Operation Deposits Mandalay engaged AMC to complete an Independent Mineral Resource Estimate and report for the Augusta deposit and Cuffley lode, suitable for NI 43-101 compliant reporting.

Mandalay provided the exploration data used by AMC and undertook the data collation, interpretation and preliminary modelling. Most of the exploration data consisted of recent drilling information which AMC has reviewed. Geological interpretation and modelling were reviewed by AMC and the Resource Estimation completed by AMC.

2.2 Work Programme The Costerfield PEA Report herein is a collaborative effort between Mandalay, SRK and AMC.

SRK and AMC have independently reviewed the work completed by Mandalay.

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

The Technical Report was assembled in Melbourne during the months of June to September 2013.

2.3 Basis of Technical Report The purpose of this Technical Report is to present the findings of a PEA for both the Augusta deposits and Cuffley lode.

This report is based on information provided by Mandalay to SRK and verified during a site visits during 2012 and 2013 and any additional information provided by Mandalay throughout the course of SRK’s investigations. SRK has no reason to doubt the reliability of the information provided by Mandalay. The most recent site visit was undertaken on 22 May and 4 June 2013.

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

• Discussions with Mandalay personnel;

• Inspection of Mandalay’s Costerfield Operation; and

• Additional information and studies provided by Mandalay.



2.4 Qualifications of SRK and SRK Team The SRK Group comprises over 1,600 professionals, offering expertise in a wide range of Resource engineering disciplines. The SRK Group’s independence is ensured by the fact that it holds no equity in any project and that its ownership rests solely with its staff. This fact permits SRK to provide its clients with conflict-free and objective recommendations on crucial judgment issues. SRK has a demonstrated track record in undertaking independent assessments of Mineral Resources and Mineral Reserves, project evaluations and audits, technical reports and independent feasibility evaluations to bankable standards on behalf of exploration and mining companies and financial institutions worldwide. The SRK Group has also worked with a large number of major international mining companies and their projects, providing mining industry consultancy service inputs.

The compilation of this Technical Report was completed by Peter Fairfield, Principal Consultant (Project Evaluation), BEng (Mining), FAusIMM (No 106754). 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.

Brett Muller, SRK Associate Principal Metallurgist, undertook a review of the metallurgical and mineral processing aspects of the project. Brett Muller, Principal Consulting Engineer (Metallurgy), BEng (Minerals Engineering and Extractive Metallurgy), BCom (Finance), member of The Institute Engineers Australia (Chemical College No 3442998), SRK Associate, 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.

Dr Andrew Fowler, Senior Geologist, AMC Consultants, undertook a review of the Resource aspect of the project. Dr Andrew Fowler, Senior Geologist, (AMC Mining Consultants), PhD (Structural Geology), MAIG, MAusIMM, CP (Geo), 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.

The lead author of the Technical Report is Peter Fairfield.

Internal SRK Peer Review of this Technical Report was completed by Anne-Marie Ebbels, Principal Consultant (Mining), who conducted a peer review of non-geological aspects of this Technical report. Anne-Marie Ebbels, BEng (Mining), MAusIMM (CP) (No 111006) 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.



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 Evaluation) SRK 4 June 2013 BEng (Mining),

FAusIMM Sections 1 -6, 15,and

Sections 18 - 27

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

(Computer Studies), MAusIMM (CP)

SRK Peer Review

Andrew Fowler Senior Geology Consultant AMC Consultants 17 August 2012

PhD (Structural Geology, BSc (Hons), MAIG, MAusIMM CP

(Geo)

Resources Sections 7 to 12, 14

Brett Muller Managing Director SRK Associate 9 March 2010 BEng (Metallurgy), BCom (Finance), M IEAust

Metallurgy and Processing Sections 13 and 17



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

2.6 Declaration SRK’s opinion contained herein and effective 30 June 2013 is based on information collected by SRK throughout the course of SRK’s investigations, which in turn reflect various technical and economic conditions at the time of writing. Given the nature of the mining business, these conditions can change significantly over relatively short periods of time. Consequently, actual results may be significantly more or less favourable.

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

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

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.



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 is located at latitude of 36o 52’ 27” south and longitude 144o 47’ 38” east. The Brunswick Processing Plant is located approximately 2 km north west of Augusta. The Cuffley lode is located approximately 500 m north-northwest of the Augusta workings and access is planned via an underground decline from Augusta.

Figure 4-1: Costerfield Operation Location Source: Google maps

4.2 Land Tenure AGD holds Mining Licence MIN4644, issued by the Victorian State Government under the Mineral Resources (Sustainable Development) Act 1990 (MRSD Act). This licence was due to expire on 30 June 2012, and a renewal of the mining licence to 30 June 2014 was submitted on 15 May 2012 and is currently being processed. This licence covers the current and future planned mining activity.

Mandalay has been advised in writing by the DPI that the tenement is currently being reviewed by the Mining Warden and that the license continues in operation and Mandalay is able to continue working the licence.

Tenure information for the single Mining Licence and three Exploration Licences is shown in Table 4-1.

N



Table 4-1: Granted Tenement Details

Tenement Name Status Company Area 2 Grant Date Expiry Date

MIN4644 Costerfield Granted 1 AGD Operations P/L 1219.3 Ha 25/02/1986 30/06/2012 EL3310 Costerfield Granted AGD Operations P/L 59.0 GRATS 17/09/1993 17/09/2013 EL4848 Costerfield Granted AGD Operations P/L 18.0 GRATS 28/01/2005 27/01/2014 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

Application 04/02/13

Mandalay Resources Costerfield

Operations Pty Ltd

0.96 Ha

EL5452 Antimony Creek

Application 20/11/12

Mandalay Resources Costerfield

Operations Pty Ltd

2.07 Ha

1 MIN4644 pending renewal 2 1 GRATS is equivalent to 1 km2

Mandalay, through its subsidiary AGD, manages the Costerfield Operation and holds a 100% interest in MIN4644, MIN5567, EL3310, EL4848 and EL5432, EL5464 as shown in Figure 4-2 and Figure 4-3. All granted tenements are in good standing as of the Effective Date of this Report.



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



Figure 4-3: Current AGD Mining and Exploration Lease Boundaries (Mandalay, 2012)



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 DPI, requires that negotiation of access and/ or compensation agreements with landowners 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 MIN4644.

4.4 Environmental Liability The rehabilitation bond is currently set at AUD2.511M and is reviewed by the DPI every two years or when a Variation to the Work Plan is approved.

There is a further AUD10,000 bond for each of the following, EL3310, EL4848 and with Vic Roads for licenses for pipelines crossing roads.

The total bond is AUD2.541M.

Rehabilitation is undertaken progressively at Costerfield Operation, with the environmental bond only being reduced when rehabilitation of an area or site has been deemed successful by the DPI. 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 DPI. 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 DPI 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 2012, there is approximately AUD31M 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 MIN4644 has a series of licence conditions that must be met and are the controlling conditions upon which all associated Work Plan Variations (WPV) are filed with the regulatory authority.



Apart from the primary mining legislation which consists of the MRSD Act, operations on MIN4644 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 DPI and are registered against the tenement.

There are no Native Title issues relevant to MIN4644.

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 Bendigo 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 Sustainability and Environment (DSE) 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 lode is located beneath unrestricted Crown land that consists of sparse woodland, with numerous abandoned shafts and workings along the historic 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 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 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 south west 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 90,000 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 propose Cuffley operations will 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 100,000 tonnes per annum (tpa) gravity-flotation-CIP (carbon-in-pulp) 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 a Chinese smelter. 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.

Water supply, infrastructure and management and Tailing Storage are discussed in Section 18. 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 which is reticulated back underground once it has flowed through a series of settling ponds.

Potable water is trucked in from Heathcote, while sewage is captured in sewage tanks before being trucked off site by a local contractor.



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 3676.2 Metals Investment Holdings 1971 12 1760.8 Victoria Mines Department 1975 – 1981 32 3213.0 Federation Resources N.L 1983 – 2000 27 2398.3 AGD/Planet Resources JV 1987 – 1988 23 1349.2 AGD N.L 1987 – 1988 14 1680.8 1994 – 1995 142 1368.5 5536.0 1996 59 195.5 2310.0 1997 23 725.0 AGD Operations 2001 27 3361.1 2002 7 907.5 2003 30 1522.0 2004 27 3159.9 2005 31 4793.4 2006 – 2007 67 4763.4 2007 – 2008 11 2207.2 2008 – 2009 8 1785.8 Subtotal Surface 573 32714.4 13999.3

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 USD 0.45 per pound to USD 1.70. 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 programs. These were conducted across the south Costerfield area encompassing the Alison Mine and south towards Margaret’s, encompassing both the Cuffley lode and the Augusta Mine areas. Diamond drilling for MEM was most successful at the Brunswick Mine. Dropping antimony prices in 1971 caused MEM to abandon their 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) program 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. Antimony 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 programs 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 AGD’s drilling of the MH Zone commenced on 5 April 2001. In total, 27 holes were completed for 3301.1 m. All holes were drilled with an initial PQ or HQ collar to approximately 25 m and then finished with NQ-sized core, 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 program. Much softer selvages to the Mineralisation were successfully recovered during this program but core loss was still estimated to be up to 15%



within the Mineralised zones. All holes were downhole surveyed and orientated during drilling. Collar locations were surveyed by Cummins & Associates from Bendigo.

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

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 metres 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 1514 m were drilled by AGD as part of an infill and extension program 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 program when compared to previous drilling programs, even though the majority of drillholes were drilled in the weathered zone.

In addition to the infill and extension program, 14 RC drillholes were drilled as part of a metallurgical test work program. 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 program 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 program 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 AGD’s drilling activities throughout 2006 and 2007 comprised grid drilling of the Brunswick Deposit and drilling of the periphery of the Augusta Deposit for a total of 7,562 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 1813 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 m by 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.

2007/ 2008 AGD’s drilling activities throughout 2007 and 2008 comprised reconnaissance drilling of the Tin Pot Gully Prospect and drilling along strike and down-dip of the existing Augusta Deposit. A total of 3395.6 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 metres, 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 Au and 16.19 to 47.20% Sb;

• 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 Au and 3.25 to 32.2% Sb; 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 Au and 6.81 to 27% Sb.

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 2207.2 m that were drilled on the northern section of the Augusta Deposit, particularly from 4411 mN to 4602 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.



2008/ 2009 AGD’s drilling activities throughout 2008 and 2009 comprised drilling along strike and down-dip from the existing Augusta resource. A total of 2585.95 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.



7 Geological Setting and Mineralisation 7.1 Regional Geology

The Costerfield gold-antimony field 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). This is shown in Figure 7-1.

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 field is centred on the Costerfield Dome, with its core of poorly-exposed lower Silurian Costerfield Siltstone. 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 south-west 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 two to six metres of alluvium which was deposited at the meandering Mountain Creek/ Wappentake Creek confluence.



Figure 7-2: Costerfield Property Geology and Old Workings Source: Mandalay, 2012.



7.3 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 kms. Individual lodes can persist for up to 800 m in 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.

The fault-hosted narrow lodes have been formed in multiple phases. The earliest phase was the bedding parallel laminated quartz veins, which are barren. The next was the quartz-pyrite-arsenopyrite phase with erratic coarse gold. The last was the stibnite-gold phase, which hosts evenly-distributed, fine-grained gold. The Costerfield lodes are typical anastomosing, en échelon style, narrow-vein systems, dipping from 25° to 70° west to steep east (70° to 90°East). Mineralised shoots are observed to plunge to the north in some lodes and south in other lodes.

The mineralisation occurs as single lodes and vein stockworks associated with brittle fault zones. These bedding and cleavage parallel faults range from sharp breaks of less than 1 mm to dilated shears 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.

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 7-3 is a photograph of typical mineralisation as shown in E Lode on the 1070 m RL and shows the features of the deposit discussed above. A close-up photograph of the mineralised lode is shown in Figure 7-4.

Figure 7-3: Two View of E Lode 1070 Level South, with Annotation on Right-Hand side Notes: 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 7-4: Close-up of Mineralisation in E Lode 1070 Level South Notes: For scale, the lens cap is 67 mm. Source: SRK Consulting, 2010.

7.4 Deposit Mineralisation The Augusta deposit currently comprises five economic lodes: E Lode, W Lode, NE Lode, NW Lode, and P Lode (Figure 7-5). The Cuffley lode is an additional economic structure, which is currently not included in the Augusta deposit due to its distance from the underground decline access (Figure 7-5). Some of the other potentially economic lodes have also been identified including C Lode and Brunswick 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 1100 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.

NE Lode lies approximately 90 m to the west of E Lode. The lode is subvertical and can be interpreted in drilling over a strike length of approximately 700 m with a down-dip extent of approximately 340 m. The true thickness ranges from less than 0.1 m up to 2.7 m, with an average width of 0.29 m where it has been mined. At present, the lode is open at depth.

NW Lode lies approximately 110 m to the west of E Lode. The lode dips at approximately 55˚ to the south-east, 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 NE Lode. The true thickness ranges from less than 0.1 m up to 2.4 m, with an average width of 0.39 m where it has been mined.



P 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 60 m, with a down-dip extent of approximately 60 m. It is interpreted as a small transfer structure between W Lode and NE Lode. The true thickness ranges from less than 0.1 m up to 1.9 m, with an average width of 0.39 m where it has been mined.

The Cuffley lode lies approximately 200 m to the west of E 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 220 m. It has an average true thickness of approximately 0.53 m. At present, the lode is open at depth.

C Lode lies approximately 80 m to the west of E Lode. The lode dips at approximately 60˚ to the west and occurs in two pods along strike from each other that have a combined strike length of approximately 300 m, and a down-dip extent of approximately 90 m. It has an average true thickness of approximately 0.42 m.

The Brunswick Lode lies approximately 600 m west and 1300 m north of E Lode. The lode is approximately vertical and occurs over a strike length of approximately 350 m, with a down-dip extent of approximately 200 m. It has an average true thickness of approximately 1.28 m. At present, the lode is open at depth.

Figure 7-5: Augusta and Cuffley Schematic Geological Cross Section at 4350N, Costerfield Mine Grid

Note: NE lode is annotated as N Lode on the figure. NW and P Lodes are too small to display. Source: Mandalay, 2012.



8 Deposit Types The Costerfield field is part of a broad gold-antimony province mainly confined within 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, un-mineralised, puggy, sheared fault structures such as the King Cobra Fault. Exploration in the property is guided by predictions of where these fault/ lode intersections might be, using structural mapping combined with aero-magnetic images.

Large, flat, west and northwest-dipping reverse faults have displaced the lodes in the Costerfield Mine at the northern end of the field. It has been recognised that such faults occur throughout the field. At the Alison Mine, production stopped in 1922 because the lodes were ‘lost’ on a flat west-dipping fault, since named the Adder Fault. Drilling in 2011 successfully targeted a displaced lode below the fault, now known as the Cuffley lode.



9 Exploration Exploration work 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, with such methods leading to the discovery of the Augusta Deposit.

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 the 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

− Wet bulk density

• Inductive properties:

− Magnetic susceptibility

− Diamagnetic susceptibility

− Electromagnetic conductivity

• Galvanic properties:

− Galvanic resistivity

− Chargeability.

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

9.3 Geophysics

9.3.1 Ground Geophysics Based on the results of the petrophysical testing program, a limited program of ground magnetic, gravity, and IP profiling, with optimal measurement parameters, was carried out across the Augusta Deposit. None of the techniques were found to be effective 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 (MIN4644) The effectiveness of bedrock geochemistry was demonstrated by Mid-East Minerals N.L. (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 program, the aircore bit was replaced by a reverse circulation (RC) bit. A total of 104 holes were drilled for a cumulative total of 547 m and an average hole depth of 5.2 m.

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



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

Plots of the MEM 1970 bedrock geochemistry on the Augusta Lode plans illustrate that antimony mineralisation within 50 m of surface (or top-of-bedrock) can be accurately detected by close-spaced (15 m) sampling on 100 m traverses across strike (Figure 9-1). The antimony halo is subdued where the high-grade lode is greater than 50 m below top-of-bedrock. For example, on cross-section 4550N, M032 drillhole intersected E Lode (55% antimony, 59 g/t gold) 35 m below surface and MH041 drillhole intersected E Lode up-dip (353% 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 ppm – 400 ppm 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

Augusta South Auger Bedrock Geochemistry Drilling (MIN4644) Mandalay engaged Statewide Drilling to undertake the Augusta South auger geochemistry program with an Edson 200 auger rig. The three traverse programme (3600N, 3500N, 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 program determined when bedrock was intersected by observing the cuttings and ensured that the driller penetrated at least 1 m into bedrock. A 3 kg – 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 program. Using a spear device to subsample introduces potential bias to the resultant assay, as particles do not have an equal opportunity of being sampled. However, it 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



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



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.

9.6 Underground Face Sampling Approximately 60% 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:

1 The face is marked out by the sampler to show the limits of the lode and the bedding angle;

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

3 The face and lengths of the sample are measured;

4 The face is photographed (a chalkboard is used to display the date, name of the face, and northing);

5 Each sample is collected as a channel sample using pick or pneumatic chisel and placed into a sample bag;

6 Care is taken to obtain a representative sample;

7 Where there are two or more lode structures in the face, samples are also taken of the intervening siltstone;

8 Samples are between 0.5 kg and 2 kg in weight;

9 The sample length can vary from 5 cm to 1.5 m across the structures;

10 Each sample bag is assigned a unique sample ID;

11 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;

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

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

14 Face samples are taken at the appropriate orientation to the mineralisation and are representative of the mine heading being sampled.



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 (Table 10-1).

Table 10-1: Drillhole Summary

Year No. of Holes

Metres (Diamond)

Metres (Percussion / Auger)

Surface Drilling

2009 – 2010 106 126.4 547.0 2010 – 2011 224 2430.8 732.0 2011– 2012

2013 2646

7 11017.9 2386.7

7295.6

Total 2983 15961.8 8574.6

Underground Drilling

2009 – 2010 11 332.5 2010 – 2011 24 8191.9 2011– 2012 34 7563.5

Total 69 16087.9

10.2 2009/ 2010 Drilling through 2009 and 2010 mainly comprised drilling along strike and down-dip from the existing Augusta Resource. In total, 332.5 m of diamond coring was undertaken.

In addition, 547 m of bedrock geochemistry aircore drilling was completed within MIN4644 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, 24 holes for 8,192 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 NE 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. NE 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 30,107 m of HQ / NQ diamond coring, 1222 m of LTK48 diamond drilling, and 11,097 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 rig, two LM90 diamond rigs, one Kempe underground diamond 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 NE Lodes Inferred and Indicated Mineral Resource and give confidence of the structural continuity of W and NE 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 program began in July 2011 (AD series of holes), following the MB007 discovery. As a follow-up program, 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.

From July 2011 to December 2012, 34 holes have been completed, totalling 13,937 m. A portion of this 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 mRL700, 490 m below the surface. This has provided confidence in the depth continuity of the lode, thus drilling will be ongoing in 2013, to Inferred Resource category initially, to ‘bound’ the extent of the lode.

10.5 2013 Cuffley lode Drilling From January 2013 to July 2013, seven drillholes targeting the Cuffley lode were drilled 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: January to July 2013

Hole ID

Collar Easting (Mine Grid) (m)

Collar Northing

(Mine Grid) (m)

Collar RL

(Mine Grid) (m)

Azimuth (degrees)

Dip (degrees)

Intercept depth below

surface (m)

True Width

(m)

Gold Grade (g/t)

Antimony Grade

(%)

AD030 15284.6 4970.0 1182.0 294.2 60.3 1121.7 0.37 39.7 25.6

AD031 15284.6 4970.0 1182.0 280.4 54.9 1127.1 0.30 39.1 34.6

AD032 15284.6 4970.0 1182.0 266.6 60.2 1121.8 1.02 106.2 17.4

AD039 15348.0 4973.0 1182.5 277.5 59.1 1122.9 0.11 32.8 33.0

AD040 15306.0 4916.0 1183.0 290.1 67.6 1114.4 0.51 7.0 1.9

MH335 15445.8 4574.9 1020.2 292.1 24.8 1157.2 0.18 2.4 0.0

MH336 15445.8 4574.9 1019.8 292.8 35.8 1146.2 0.99 0.8 2.5

10.6 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. Assay information gained from this drilling is not used as input to mineral resource estimation. 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.

10.7 Collar Surveys Since 2006, hole collars have been surveyed according to the current Costerfield Mine Grid, either by Mandalay surveyors or by GWB Survey Pty Ltd. Between 2006 and 2011, Adrian Cummins & Associates provided surveying 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. Drill 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.

Figure 10-1 illustrates the known drillhole collar locations for both Augusta and Cuffley.

Figure 10-1: Collar Locations near Augusta and Cuffley lodes Note: Black dots are drillhole collars used to estimate Mineral Resources. Mineralised lodes at 1025 m RL are projected to surface. Mandalay mine grid.



10.8 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.

10.9 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 on to 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.



10.10 Drilling Pattern and Quality

10.10.1 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 1000 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 NE Lodes can vary greatly. Where increased geological confidence is required, infill holes specifically targeting NE or E Lodes have been drilled. 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.

10.10.2 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 detailed infill drilling between the 930 mRL and 960 mRL of the Cuffley lode, it is intended that the ‘W’ pattern will again be used so as to maximize strike information. This is as opposed to depth information, which is gained by drilling on section. It is intended that this infill drilling will be conducted on an approximate 30 m by 40 m grid.

10.11 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 space. Figure 10-2 illustrates a typical cross-sectional interpretation completed for the Augusta deposit.



Figure 10-2: Cross Section at 4430N through the Augusta Deposit Source: Mandalay, 2012

10.12 Factors that could Materially Impact the 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 at 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 of a repeatable nature, with appropriate QA/QC methodologies employed. The assay method used is also considered to be appropriate for this style of mineralisation.

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



The nature of 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.

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.

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 block model estimates come from both drill core and channel sampling along the ore development drives, while the Cuffley lode estimate was informed using only diamond drill core samples.

11.1.1 Diamond Core Sampling Much of the drill core produced from the Costerfield area is composed of barren siltstone. Thus, not all diamond drill core is 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:

• All stibnite-bearing veins are sampled;

• A waste sample is taken either side of the mineralised vein;

• 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 representative sample, the diamond drill core is cut in half with a diamond saw along the top or bottom mark of the orientated core.

Sampling intervals for drill core used for resource estimation purposes are no smaller than 5 cm in length and no greater than 1.5 m in length. The average sample length for drill core samples within the Augusta drill program is 61 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. 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 5 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.

11.2 Data Spacing and Distribution Within the Augusta Deposit, 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 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:

1 Sample material is placed into a calico bag previously marked with a sample number;

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

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

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

5 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:

1 Samples are received and checked against the submission sheet;

2 A job number is assigned and worksheets as well as sample bags prepared;

3 Samples are placed in an oven and dried overnight at 80°C;

4 The entire sample (up to 2 kg) is jaw-crushed to approximately 2 mm;

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

6 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:

1 50 g of pulp is fused with 180 g of flux (silver);

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

3 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;

4 1.4 mL of hydrochloric acid (HCl) is added;

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

6 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:

1 0.2 g of sample is added to a flask of distilled water (20 mL);

2 30 mL of 50% nitric acid is added;

3 20 mL of tartaric acid is added;

4 80 mL of 50% HCl is added and allowed to stand for 40 minutes;

5 5 mL of hydrobromic acid (HBr) is added;

6 The solution is mixed for one hour and left to stand overnight until fuming ceases;



7 Sample is heated until colour changes to light yellow and white precipitate dissolves;

8 When cool, the sample is diluted to 200 mL with distilled water; and

9 Antimony content is determined by AAS.

11.6 Laboratory Reviews A mining consulting company commissioned by Mandalay visited the Onsite Laboratory in Bendigo on the 22 December 2011 to conduct an inspection of the facility. A tour of the laboratory was completed in the presence of Onsite’s Laboratory Manager, Mr Rob Robinson.

The mining consulting company found that the goods inwards, sample preparation, fire assay, and sample storage sections of the laboratory are effectively located in one large open building/ warehouse. A large curtain partly separates the fire assay from the sample preparation sections. The open nature of the laboratory building poses some risk of contamination via dust transfer throughout the open area. There are large open doors to the front and rear of the building that allow for airflow through the entire building space. This; however, also allows for dust transfer around the building. This situation is far from optimal, though the mining consulting company found no other causes for serious concern during its inspection.

Mandalay undertakes regular check assay programmes. The most recent of these was undertaken in January 2013. 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 Mandalay has implemented a number of projects to improve the confidence of assay data. This included a re-assay programme which demonstrated the reliability of the original dataset. A recompilation and analysis of the QA/QC dataset collected for site standard reference material and gold duplicates has been completed.

Confidence with the sample collection procedures and assay results is also provided by the good reconciliation observed between the resource block model tonnes and grade, the grade control model tonnes and grade, and the tonnes and grade received at the processing plant during 2011 and 2012.

11.7.1 Standard Reference Material A recompilation and analysis of the QA/QC dataset collected for site standard reference material and gold duplicates has been completed.

Two standards that are currently in use have been made from material collected underground at Augusta (AGD08-01 and AGD08-02), and are routinely submitted to Onsite. Additionally, Mandalay routinely sends two commercially available, gold-only standards to Onsite (G901-8 and G907-6). A standard is sent with each batch of exploration samples and twice a week with the underground face samples.

Results from January 2012 to December 2012 are displayed in Figure 11-1 to Figure 11-4. 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.

A review of the results shows that:

• G901-8 gold displays good compliance for the period, with three exceptions. There is also a low bias displayed for most of the period (Figure 11-1);

• G907-6 gold displays good compliance for the period, with three exceptions. There is also a slight low bias displayed for most of the period (Figure 11-2);

• AGD08-01 antimony displays good compliance for the period, with one exception. There is a slight low bias for part of the period (Figure 11-3); and

• AGD08-02 antimony displays good compliance for the period, with five exceptions. There is also a high bias displayed for most of the period (Figure 11-4).

AMC considers that the level of compliance and bias displayed by the standards is acceptable and demonstrates the reliability of the gold and antimony grades used to inform the block model estimate.

AMC recommends that Mandalay obtain an antimony standard that is closer to the lode antimony grade (approximately 18% Sb).

Figure 11-1: G901-8 Gold Standard Reference Material – Assay Results January 2012 to December 2012



Figure 11-2: G907-6 Gold Standard Reference Material – Assay Results January 2012 to December 2012

Figure 11-3: AGD08-01 Antimony Standard Reference Material – Assay Results January 2012 to December 2012



Figure 11-4: AGD08-02 Antimony Standard Reference Material – Assay Results January 2012 to December 2012

11.7.2 Blank Material Mandalay sends uncrushed samples of basalt as blank material 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 of blank material exceeded the 0.02 g/t Au threshold (Figure 11-5). This result is unacceptable and Mandalay is currently working with Onsite to resolve the issue. AMC notes, however, that the highest blank value obtained (0.81 g/t Au) is still <3% of the average lode grade of 32.7 g/t Au. Therefore, AMC considers that the high blank values are not material for the Mineral Resource estimate.

Figure 11-5: Blank Gold Results for Onsite Note: The 0.02 g/t threshold is marked with a red line



11.7.3 Duplicate Assay Statistics A summary of laboratory duplicate statistics assayed by Onsite for original gold assays versus duplicate gold assays for the life of the project is presented in Table 11-1. The duplicates are assayed on separate aliquots of the same sample pulp. A scatter plot of this data is presented in Figure 11-6, 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 11-7. 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 Augusta duplicate dataset achieves 76% of pairs at less than 10% RPD, which demonstrates adequate precision in the gold assays by Onsite.

Table 11-1: Summary of Onsite Duplicate Statistics

Au Original Duplicate

Number of samples 885 885

Mean 16.51 17.00

Maximum 369.90 462.60

Minimum 0.01 0.01

Pop Std Dev. 36.21 38.84

CV 2.19 2.28

Bias -2.97%

Cor Coeff 0.99

Percent of samples < 10% RPD 76.03

Figure 11-6: Scatter Plot for Onsite Gold Duplicates (g/t)



Figure 11-7: Relative Paired Difference Plot for Onsite Gold Duplicates (g/t)

11.7.4 Check Assay Program – Sample Pulps Duplicate statistics resulting from a gold and antimony check assay program comparing Onsite and ALS Minerals (ALS) are presented in Table 11-2 and Table 11-3 respectively. The check assay program was undertaken by Mandalay in February 2013. The duplicates are assayed on separate aliquots of the same sample pulp.

The gold scatter plot of this data (Table 11-2) shows significant bias (31%) between Onsite and ALS, with ALS returning higher grades than Onsite. However, the antimony scatter plot shows some bias (5%), with Onsite returning higher grades than ALS (Figure 11-9). The gold RPD plot shows that 62% of duplicate pairs show less than 20% RPD (Figure 11-10). This demonstrates poor reproducibility across laboratories. The antimony RPD plot shows that approximately 92% of duplicate pairs are less than 20% RPD (Figure 11-11), which demonstrates good reproducibility across laboratories.

AMC considers that although the number of samples included in the check assay program is low, the poor reproducibility observed in the gold assays is an issue that requires further investigation. The results suggest that Onsite may be undercalling higher-grade assays. It is worth noting that this is also observed in the gold certified standard results (Figure 11-1), but not to the same extent. If Onsite is undercalling higher-grade assays, there is a chance of underestimating the gold resource.



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

Au Onsite ALS

Number of samples 39 39 Mean 34.42 44.86 Maximum 124.60 165.00 Minimum 1.06 1.10 Pop Std Dev. 32.54 43.52 CV 0.95 0.97 Bias -30.35% Cor Coeff 0.85 Percent of samples < 10% RPD 61.54

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

Sb Onsite ALS

Number of samples 39 39

Mean 22.92 21.68

Maximum 55.10 54.50

Minimum 0.81 0.85

Pop Std Dev. 13.21 12.81

CV 0.58 0.59

Bias 5.41%

Cor Coeff 0.99

Percent of samples < 10% RPD 92.31

Figure 11-8: Scatter Plot for Onsite versus ALS Gold Duplicates (g/t)



Figure 11-9: Scatter Plot for Onsite versus ALS Antimony Duplicates (%)

Figure 11-10: Relative Paired Difference Plot for Onsite versus ALS Gold Duplicates (g/t)



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

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.

11.9 Conclusions AMC 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 50% of blank samples exceeded the acceptable threshold is of concern and is being investigated. However, AMC does not believe that the high blank values are material for the Mineral Resource estimate because the highest blank value obtained (0.81 g/t Au) is still <3% of the average lode grade of 32.7 g/t Au;

• The poor reproducibility observed in the gold assays for duplicate pulps assayed by Onsite and ALS suggests that Onsite may be undercalling higher-grade assays. It is worth noting that this is this is also observed in the gold standard reference material results, but not to the same extent;



• Security of samples being transported to the laboratory is adequate. However, the storage of returned pulps needs to be reviewed and rectified;

• AMC recommends that Mandalay obtains an antimony standard that is closer to the lode antimony grade (approximately 18% Sb);

• AMC recommends that Mandalay investigates the poor reproducibility of the gold assays between Onsite and ALS; and

• AMC recommends that Mandalay undertakes quarterly laboratory inspections and reviews of Onsite, and institutes check assay programs on a quarterly basis.



12 Data Verification On 17 August 2012, AMC full-time employee Dr A Fowler (QP for sections 7 to 12 and Section 14) visited the Augusta Mine site and Brunswick exploration core shed. All drillcore 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

− Sample storage

− QA/QC

− Data validation procedures

− Underground mapping procedures

− Survey procedures

− Geological interpretation

− Exploration strategy

• A review of underground face photos and back mapping;

• An inspection of the core shed and drill core intersections from Augusta and Cuffley; and

• Random cross-checks of 100 assay results in the database with original assay results returned from Onsite Laboratories.

AMC makes the following observations based on the data verification undertaken:

• Site geologists are appropriately trained and are conscious of the specific sampling requirements of narrow-vein, high-grade deposits;

• Considerable attention is paid to the collection and analysis of structural information from drill core and underground mapping. This structural analysis is used to successfully guide exploration and mining; and

• Cross-checking the database with the original assay results uncovered one transposition error.

AMC considers that one error out of one hundred is an acceptable level of accuracy. However, it recommends that Mandalay conducts its own regular, random checks of the database in future.

In AMC’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 – Augusta

13.1.1 Historical Testwork From 2004 to May 2012 Augusta samples have been tested to determine the amenability of the ore to metallurgical recovery. The following companies were involved with various aspects of the metallurgical evaluation:

• ALS Ammtec, Brookvale NSW (Previously known as Metcon Laboratories) (Metcon); and

• AMDEL LIMITED Mineral Services Laboratory, Torrensville Plaza, South Australia (Amdel).

Historically, Amdel tested drill core samples in 2004 prior to the processing of the Augusta ore through the Brunswick processing plant. Amdel tested the ore 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 testwork since 2008, the aim of this testwork has been to compare the recovery of the laboratory tested samples to the mill reconciled recovery. In total, ten separate ore characterisation tests have been completed on Augusta ore. Samples were selected via face sampling of ore drives or mill feed belt cuts to coincide with what was being processed at the time in 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 testwork 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.

Table 13-1 and Table 13-2 outline the plant performance recovery for antimony and gold separately, compared to the historical testwork for the Augusta mined and processed ore.

For the first four ore characterisation tests (M1592, M1609, M1654 and M2091), there were no accurate flotation feed sizing’s for comparison, hence the percentage difference between the plant versus testwork recovery cannot be calculated. This is also why there is a range given for the test work recovery as the metallurgical testing was completed at different size fractions.

Tests M2115 and M2139 were completed only at the p80 106 µm size range, which is not comparable to the actual plant flotation feed size range that was p80 76 µm and p80 60 µm at the time.

The remaining three tests, M2305, M2421 and M2610, were tested at similar flotation feed size ranges to the actual plant feeds size at the time and showed good correlation for antimony recovery. However, the gold recovery results are inconsistent when compared to the plant gold recoveries.



Table 13-1: Ore Characterisation testwork vs. plant reconciled recovery results for Antimony

Ore Characterisation Sb Results

Laboratory Testwork

Recovery %

Reconciled Plant Recovery

% Plant flotation

feed Plant vs

Laboratory Testwork

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 @106um 98 90 76 Flotation feed size difference too great

for direct comparison

2010 July M2139 @106um 73 88 60

2011 January M2266 @75um 94 83 60 11

2011 February M2305 @53um 94 92 47 2

2011 July M2421 @53um 97 96 47 1

2012 May M2610 @53um 97 96 47 1

Table 13-2: Ore Characterisation testwork vs. plant reconciled recovery results for Gold

Ore Characterisation

Au Results

Laboratory Testwork

Recovery % Reconciled Plant

Recovery % Plant flotation

feed Plant vs Laboratory

Testwork

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

Figure 13-1 illustrates the flow sheet for the Metcon ore characterisation testwork, specifically for test M2610. As outlined previously, the gold deportment test work on the flotation tailings has been conducted since July 2011 (i.e. test M2421). Not all size fractions have been tested for every ore characterisation test.

Table 13-3 shows what size ranges have been tested for each ore characterisation test.



Figure 13-1: Ore Characterisation testwork (M2610) vs plant reconciled recovery



Table 13-3: Ore Characterisation testwork size range tested per test

Ore Characterisation Tests

Flotation Feed p80 Range (µm)

150 106 100 75 53 38

2008 January M1592

2008 April M1609

2008 May M1654

2010 May M2091 2010 June M2115 2010 July M2139 2011 January M2266 2011 February M2305 2011 July M2421 2012 May M2610

In summary, considering the recent improvements in plant recoveries on the existing Augusta ore, there is good correlation between the plant recovery and testwork recovery results for antimony, however, there is inconsistency with the gold recovery comparisons with the plant data.

To determine the potential gold and antimony recoveries of the existing Augusta orebody, the current operational recoveries provides the best estimate of future performance, as discussed in Section 13.2.

13.1.2 Recent Testwork Extensive metallurgical testwork was completed to determine the metallurgical characteristics of the Cuffley lode. The aim of this test work was to highlight any potential processing issues with regards to the Cuffley mineralisation as well as enable a processing comparison between Augusta and Cuffley. The following companies 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).

To determine the potential differences with regards to processing between the currently processed Augusta ore and potential Cuffley mineralisation, three different Cuffley samples and one Augusta orebody sample, labelled 2012 May Mill Feed, were sent to Metcon for ore characterisation testwork. A summary of this testwork is shown in Table 13-4.

Table 13-4: Testwork Schedule and Grades

Cuffley Testwork Schedule and Grades


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



The Metcon ore characterisation testwork involves testing specifically targeted grind sizes and the ore amenability at these sizes to gravity concentration, flotation, cyanide recovery and gold deportment on the flotation tailings, to determine the percentage of cyanide soluble, sulphide locked and silicate locked gold. The flowsheet for the Metcon ore characterisation testwork is shown in Figure 13-2.

Figure 13-2: Pre-Characterisation testwork flow sheet



13.1.3 Test Samples Sample Selection for Cuffley #1 Sample Approximately 3 m of ¼ core was sampled for each drillhole with a total of seven drillholes sampled to obtain a 35 kg representative sample of the Cuffley lode. The length of sample was calculated using the angle between the lode structure and each individual drillhole with the intent to achieve a sample that would represent a 2.2 m wide mining excavation, which in turn would represent the head grade delivered to the Brunswick processing plant. The drill intercepts used are roughly evenly spaced over an area of 200 m by 130 m, as shown in Figure 13-3. The drillholes included are MB007, AD001, AD002, AD005, AD006, AD007 and AD008. The Cuffley #1 Sample is considered to be representative of the high grade core of the Cuffley lode.

Figure 13-3: Longitudinal Section of the Cuffley Deposit intercepts as at January 2013

Sample Selection for Cuffley #2 Sample Approximately 3 m of ¼ core was sampled from each drillhole with a total of six drillholes sampled to obtain a 30 kg sample of lower grade material from the Cuffley lode. The length of sample was calculated using the angle between the lode structure and each individual drillhole with the intent to achieve a sample that would represent a 2.2 m wide mining excavation, which in turn would represent the head grade delivered to the Brunswick processing plant for this bounding area. The drill intercepts used bound the area in which the Cuffley #1 Sample was taken, as shown in Figure 13-4. The drillholes included are AD009, AD0010, AD0011, AD0012, AD013 and AD016. The Cuffley #2 Sample is considered to be representative of what is interpreted as the lower grade halo material from of the Cuffley lode.

MB007

AD001 AD002

AD005

AD007 AD008 AD006



Sample Selection for Cuffley #3 Sample – Ball Mill Work Index Sample Approximately 2 m of ½ core was sampled from three holes. One metre of core was sampled either side of the Cuffley #2 Sample from holes, AD009, AD013 and AD016, as shown in Figure 13-4. The Cuffley #3 Sample is representative of the siltstone host rock and is representative of the expected Cuffley mineralisation hardness.

Figure 13-4: Longitudinal Section of the Cuffley Deposit intercepts as at January 2013

13.1.4 Cuffley 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, which is further discussed in Section 17.

The mineral abundance as shown in Figure 13-5 shows 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 gold and 3.55% antimony).

The similarities in the mineralisation and the recoveries achieved from the ore characterisation testwork suggest that there is no significant difference in mineralisation that will have an adverse effect on the Cuffley mineralisation processing recovery.

AD009 AD011 AD013

AD016 AD012 AD010



Figure 13-5: Amdel Qemscan Mineral Abundance

13.2 Stibnite Liberation Antimony liberation was slightly better for the Cuffley #1 sample when compared to the Mill Feed 2012 Sample. A liberation of 94.2% of the antimony was achieved in the Cuffley #1 Sample compared to the Mill Feed May 2012 Sample, which achieved 89.4% liberation.

These results show that the Cuffley mineralisation potentially has better liberation characteristics than the Augusta ore, however, the effect of feed grade differences may be a contributing factor with these liberation percentages. Overall, there does not seem to be any liberation issues with the Cuffley mineralisation. Antimony liberation is shown below in Table 13-5 and Figure 13-6.

Table 13-5: Amdel Qemscan Stibnite Liberation Table

Mass %

Cuffley #1 Sample

Mill Feed May 2012

=100 83.9 78.1 Liberated

90-100 10.3 11.3

80-90 0.9 1.7

High Middling 70-80 0.7 1.0

60-70 0.4 0.7

50-60 0.4 0.6

Low Middling 40-50 0.4 0.9

30-40 0.6 1.0

20-30 0.7 1.4

Locked 10-20 0.7 2.0

0-10 0.9 1.3

Total 100.00 100.00



Figure 13-6: Amdel Qemscan Stibnite Liberation Visual Representation

13.3 Gold Liberation Gold liberation was significantly better for the Cuffley #1 Sample when compared to the Mill Feed May 2012 Sample, with 95.5% of the sample in the liberated/ high middling liberation group, compared to 67.8% for the Mill Feed May 2012 Sample as shown in Table 13-6. This is also supported by the ore characterisation test work that demonstrated a small increase in recovery for all three size ranges tested when comparing Cuffley #1 sample test work recovery to the Augusta ore (Mill Feed May 2012) results. The increase in total gold recovery ranged from 2.6% to 4.1% with an average of 3.2%, as shown in Table 13-8. However, the significant differences in feed grade cannot be ignored and may be a contributing factor in the liberation differences. Gold liberation is shown below in Table 13-6 and Figure 13-7.

Table 13-6: Amdel Qemscan Gold Liberation Table

Mass %

Cuffley #1 Sample

Mill Feed May 2012

=100 58.2 55.5 Liberated

90-100 0.0 0.0

80-90 37.3 0.0

High Middling 70-80 0.0 12.3

60-70 0.0 0.0

50-60 1.5 0.0

Low Middling 40-50 0.0 3.7

30-40 0.0 0.0

20-30 0.0 0.0

Locked 10-20 0.0 4.0

0-10 3.0 24.5

Total 100.00 100.00



Figure 13-7: Amdel Qemscan Gold Liberation Visual Representation

13.3.1 Cuffley Lode 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. The Metcon ore characterisation testwork included laboratory grind determinations which determine the appropriate grind time to get the correct feed size for the testwork. Figure 13-8 and Figure 13-9 show the difference in hardness between the two deposits. From these results, it can be assumed that there is little difference in the hardness of the two deposits.

Figure 13-8: Grind Establishment Testwork Comparisons at 75 µm

0

5

10

15

20

25

30

Grin

ding

tim

e (m

in)

Deposit

Cuffley vs. Augusta p80 = 75µm Testwork Grind Establishment



Figure 13-9: Grind Establishment Testwork Comparisons at 53 µm

13.3.2 Gravity and Flotation Recovery of the Cuffley Lode As discussed previously in Section 13.1.1, the results of the recently completed Augusta testwork closely match plant performance. The summary tables of completed testwork, Table 13-7 and Table 13-8, show the differences in laboratory performance between the Augusta ore (represented by Mill Feed May 2012 sample) and Cuffley mineralisation. The Cuffley #2 Sample has been left out of this recovery calculation, due to the low head grade of this sample. Since the feed rate has increased due to the use of the mobile crushing plant, the average flotation feed p80 has equalled 60 µm, comparable to the 53 µm laboratory test size range. (Start of October 2012 to end of April 2013). As will be discussed in Section 13.4 ‘Recovery Data on Augusta Orebody’, the current plant performance since September 2012 when tonnages increased on the Augusta ore equals 96% for antimony recovery and 89% for gold recovery. Considering the current flotation feed is approximately P80 53 µm, the calculated antimony recovery for the Cuffley mineralisation is estimated to be 0.7% less than the 96% achieved for the Augusta ore, as per Table 13-7, resulting in approximately 95%. The calculated gold recovery for the Cuffley mineralisation is estimated to be 2.6% greater than the 89% achieved for the Augusta ore, as per Table 13-8 resulting in approximately 92%.

For the purposes of this Technical Report, the historical recoveries achieved at processing ore for the Augusta Mine have been used, 96% for antimony and 89% for gold.

Deposit



Table 13-7: 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 13-8: 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

13.4 Recovery Data on Augusta Orebody Table 13-9 demonstrates the Brunswick processing plant performance on the currently mined and processed Augusta orebody from February 2012 to April 2013. During this time, the throughput has increased due to the mobile crushing plant. Increasing throughput has resulted in current recoveries of 96% for antimony and 89% for gold.

13.4.1 Mineralogical Study Amdel Laboratories have completed mineralogical testwork on a split of the sample sent for ore characterisation testwork. This was completed to compare the potentially different mineralisation between the two deposits.

13.5 Mineral Processing Antimony recovery and repeatability with the existing Brunswick processing plant is excellent. Gold recovery is not as high as the antimony recovery due to the gold component locked in sulphides in the flotation tailings ranges from 20 – 42% from Metcon testwork data.



Table 13-9: Plant Recovery Data on Existing Augusta Orebody

2012 2013 Feb 12 – Aug 12

Sep 12 – Apr 13

Feb Mar Apr May Jun Jul Aug Sep Oct Nov Dec Jan Feb Mar Apr Avg Avg

DMT Milled 5,929 7,392 7,269 6,648 6,894 6,833 6,979 9,299 10,151 10,242 11,523 9,984 8,243 8,351 9,906 6,849 9.712

Antimony recovery (%) 96 96 96 96 97 98 96 96 96 96 94 96 96 96 96 96 96

Gold recovery (%) 90 88 88 90 91 91 90 89 90 90 86 89 90 91 92 90 89

Antimony feed grade (%)

4.39 4.29 4.60 5.40 4.32 4.77 3.76 4.16 4.58 4.37 3.33 4.38 5.37 4.65 3.72 4.50 4.32

Gold feed grade (g/t) 9.10 8.25 8.22 8.78 8.03 8.81 7.09 7.09 9.24 9.40 6.27 8.85 9.59 9.96 10.32 8.32 8.84

Test period Powerscreen 31 Aug 31 to 11 Oct

Finlay Terex crusher in service from 12 Oct 12



14 Mineral Resource Estimates 14.1 Introduction

AMC has estimated gold grade, antimony grade and lode thickness using the two dimensional (2D) accumulation method. The 2D accumulation method requires that gold and antimony are multiplied by true thickness (called gold accumulation and antimony accumulation), to correctly assign weights to composites of different lengths during estimation. The interpolation method is ordinary kriging, where there are sufficient sample pairs for meaningful variography. Otherwise, inverse distance squared is used. The estimated grade is then back-calculated by dividing estimated gold accumulation and estimated antimony accumulation by estimated true thickness.

14.2 Diamond Drillhole and Underground Face Sample Statistics Statistics for gold and antimony, weighted by true thickness for the underground face samples, and diamond drillhole samples, are presented in Table 14-1. The results show that the face samples have generally captured significantly higher gold and antimony grades than the drillhole samples. This is partly a result of the face samples sampling the higher grade parts of the veins, but might also suggest that the face sample is more selective than the logged drillhole intersections.

Table 14-1: Summary Statistics

Lode Sample type Variable No. of Samples Min Max Mean Coeff. Var.

E Lode Drillhole

Au (g/t) 134 0.01 104.8 17.0 1.20 Sb (%) 134 0 59.02 10.0 1.23

Face Au (g/t) 1470 0 413.1 37.8 0.97 Sb (%) 1470 0 67.5 21.1 0.78

W Lode Drillhole

Au (g/t) 189 0 347.3 28.6 1.41 Sb (%) 189 0 63.4 13.5 1.22

Face Au (g/t) 1620 0 4650 51.6 1.80 Sb (%) 1620 0 67.5 27.6 0.73

C Lode Drillhole Au (g/t) 103 0.01 352 30.6 2.08 Sb (%) 104 0.01 45.6 8.6 1.20

NE Lode Drillhole

Au (g/t) 146 0 378.2 30.4 2.08 Sb (%) 146 0 59.9 11.3 1.43

Face Au (g/t) 199 0 437.7 40.0 1.22 Sb (%) 199 0 61.6 26.6 0.87

NW Lode Drillhole

Au (g/t) 1 39.1 39.1 39.1 0.00 Sb (%) 1 42.2 42.2 42.2 0.00

Face Au (g/t) 94 0 1433 40.0 4.06 Sb (%) 94 0 62.9 14.3 1.50

P Lode Drillhole

Au (g/t) 2 1.44 14.1 10.0 0.59 Sb (%) 2 1.22 1.84 1.4 0.20

Face Au (g/t) 71 0.04 194.8 30.7 1.51 Sb (%) 71 0 52.3 16.9 1.07

Cuffley lode Drillhole

Au (g/t) 120 0 601.2 28.8 1.99 Sb (%) 120 0 65.9 10.7 1.38

Brunswick Drillhole Au (g/t) 345 0 129.75 4.9 1.94 Sb (%) 345 0 150 2.3 4.27



14.3 Grade Capping Statistical analysis of gold grades has been undertaken previously and resulted in a gold grade cap of 150 g/t before accumulation (Figure 14-1 and Figure 14-2). This grade cap has been in operation since 2009 and has not been changed due to the generally good agreement observed between the estimated gold in the block model and the gold produced (Figure 14-32). The percentage of samples affected by the gold grade caps for each lode is summarised in Table 14-2. Statistics before and after grade capping are presented in Table 14-3. No grade cap was applied to antimony.

Grade caps were not applied in the 2009 resource estimates for C lode or Brunswick (Fredericksen, 2009).

Figure 14-1: Log Probability Plot of Gold – Face Samples Note: Orange = E Lode, Blue = W Lode, Green = NE Lode, Pink = NW Lode, Lt. Blue = P Lode

Figure 14-2: Log Probability Plot of Gold – Drillhole Samples Note: Orange = E Lode, Blue = W Lode, Green = NE Lode, Pink = Cuffley Lode



Table 14-2: Percentage of Samples Affected by 150 g/t Au Grade Cap

Lode Face samples Drillhole samples

E 2% 0% W 5% 3% NE 5% 4% NW 7% N/A P 5% N/A Cuffley N/A 3%

Table 14-3: Gold Statistics Before and After Grade Capping

Lode Sample type Variable No. of

Samples Min Max Mean Coeff. Var.

E Lode Drillhole

Au (g/t) 134 0.01 104.8 17.0 1.20 Grade capped Au (g/t) 134 0.01 104.8 17.0 1.20

Face Au (g/t) 1470 0 413.1 37.8 0.97 Grade capped Au (g/t) 1470 0 150 37.1 0.89

W Lode Drillhole

Au (g/t) 189 0 347.3 28.6 1.41 Grade capped Au (g/t) 189 0 150 27.5 1.22

Face Au (g/t) 1620 0 4650 51.6 1.80 Grade capped Au (g/t) 1620 0 150 47.0 0.92

NE Lode

Drillhole Au (g/t) 146 0 378.2 30.4 2.08 Grade capped Au (g/t) 146 0 150 24.7 1.58

Face Au (g/t) 199 0 437.7 40.0 1.22 Grade capped Au (g/t) 199 0 150 38.7 1.10

NW Lode

Drillhole Au (g/t) 1 39.1 39.1 39.1 0.00 Grade capped Au (g/t) 1 39.1 39.1 39.1 0.00

Face Au (g/t) 94 0 1433 40.0 4.06 Grade capped Au (g/t) 94 0 150 22.4 1.74

P Lode Drillhole

Au (g/t) 2 1.44 14.1 10.0 0.59 Grade capped Au (g/t) 2 1.44 14.1 10.0 0.59

Face Au (g/t) 71 0.04 194.8 30.7 1.51 Grade capped Au (g/t) 71 0.04 150 28.6 1.37

Cuffley lode Drillhole

Au (g/t) 120 0 601.2 28.8 1.99 Grade capped Au (g/t) 120 0 150 25.5 1.63

Note: Statistics are length-weighted

14.4 Data Interpretation and Compositing Mandalay coded the significant intersections according to its interpretation of the lode geometry. AMC reviewed Mandalay’s coding and suggested minor adjustments where necessary to avoid overestimation of lode thickness and contained metal. The coding system is given in Table 14-4. After the intersection coding was adjusted in consultation with Mandalay, the drillhole and underground face samples were composited over the full width of the coded intersections. The composite statistics are presented in Table 14-5.



Table 14-4: Mineralisation Coding

Mineralised Structure Zone Code

E Lode 10 E Lode Waste 11 W Lode 20 W Lode Waste 21 W Foot wall 1 22 W Foot wall 2 23 W Foot wall 3 24 W Foot wall Splay 25 W Hanging Wall 1 26 W Hanging Wall 2 27 W Hanging Wall 3 28 W Hanging Wall Splay 29 C Lode 30 C Lode Waste 31 NE Lode 40 NE Lode Waste 41 NW Lode 45 D Lode 50 P Lode 55 K Lode 60 J Lode 70 I Lode 80 Un-interpreted Significant Interval 99

Cuffley lode 110 Brunswick 200



Table 14-5: Summary Composite Statistics

Lode Sample type Variable No. of

Samples Min Max Mean Coeff. Var.

E Lode

Drillhole Au (g/t) 101 0.0 104.8 16.7 1.11 Sb (%) 101 0.0 59.0 9.8 1.04 True Thickness (m) 101 0.1 4.7 0.44 1.34

Face Au (g/t) 1111 0.0 150.0 35.1 0.85 Sb (%) 1111 0.0 67.5 21.5 0.72 True Thickness (m) 1111 0.0 2.8 0.34 0.94

W Lode



C Lode Drillhole Au (g/t) 75 0.0 352.0 28.8 1.89 Sb (%) 75 0.0 45.6 8.4 1.15 True Thickness (m) 75 0.0 2.6 0.42 1.16

NE Lode



NW Lode



P Lode



Cuffley lode Drillhole Au (g/t) 34 0.0 138.1 33.2 1.21 Sb (%) 34 0.0 35.0 10.4 1.00 True Thickness (m) 34 0.1 1.8 0.54 0.72

Brunswick Drillhole Au (g/t) 71 0.0 27.1 5.9 0.89 Sb (%) 71 0.0 37.7 2.4 1.76 True Thickness (m) 71 0.0 4.3 1.28 0.76



14.5 Dip – Dip Direction Domains To estimate true thickness from the drillhole intersections, and convert the two dimensional (2D) tonnes and grade estimates to three-dimensional (3D) tonnes and grade estimates, dip and dip-direction domains were interpreted in long section. The method to achieve this was as follows:

• Generate a surface of each lode by wireframing the centre points of each composite using the face and drillhole samples;

• Calculate the orientation of each vein on a 5 m by 5 m grid in long section; and

• Digitise strings that define domains with consistent dip and dip directions.

The domains are numbered and shown in Figure 14-3 to Figure 14-8. The dip and dip-direction statistics for each domain are given in Table 14-6 and Table 14-7. NW Lode is interpreted to have a consistent dip of 57° towards 124°. P Lode is interpreted to have a consistent dip of 85° towards 105°. Cuffley lode is interpreted to have a consistent dip of 85° towards 097°.

C Lode was interpreted to dip 60° towards 260° while Brunswick was interpreted to dip 89° towards 104° (Fredericksen, 2009).

Figure 14-3: E Lode Dip – Dip Direction Domains Coloured by Dip



Figure 14-4: E Lode Dip – Dip Direction Domains Coloured by Dip Direction

Figure 14-5: W Lode Dip – Dip Direction Domains Coloured by Dip



Figure 14-6: W Lode Dip – Dip Direction Domains Coloured by Dip Direction

Figure 14-7: NE Lode Dip – Dip Direction Domains Coloured by Dip



Figure 14-8: NE Lode Dip – Dip Direction Domains Coloured by Dip Direction

Table 14-6: E Lode Mean Dip and Dip Direction per Domain

Domain Dip (°)

Dip Direction (°)

1 55 250 2 65 262 3 59 274 4 68 261 5 74 263 6 60 270 7 67 263 8 62 259 9 75 259

Table 14-7: W Lode Mean Dip and Dip Direction per Domain

Domain Dip (°)

Dip Direction (°)

1 54 260 2 52 270 3 48 282 4 57 272 5 58 280 6 68 282 7 85 277 8 81 284 9 62 279

10 73 258 11 69 272 12 60 275



Table 14-8: NE Lode Mean Dip and Dip Direction per Domain

Domain Dip (°)

Dip Direction (°)

1 85 133 2 84 272 3 86 249 4 86 120 5 80 230 6 82 122 7 83 249 8 80 123 9 80 253

10 66 133 11 83 282 12 71 273 13 84 266 14 50 63 15 81 139 16 76 276

14.6 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%))/2.65))

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 14-9 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 is generally higher than that determined from the regression based on antimony content. However, it important to note that the formulae method has been used since 2005 to assign bulk density to resource estimates. Reconciliation between the resource block models and the processing plant has shown it to be a reliable method for estimating tonnes. Further test work is necessary to understand the difference between the formulae method and the measured bulk density.

For the resource estimate, bulk density was assigned using the formula method in line with previous estimates.



Figure 14-9: Bulk Density Determinations

14.7 Variography Variographic analysis was carried out on true thickness, gold accumulation and antimony accumulation on the composited face sample and drillhole samples. This was carried out to identify the directions of continuity of grade and thickness, and to assist in the selection of estimation ranges. The variography was carried out 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 were selected based on observations from underground mapping and from the continuity observed in the variograms. If the variograms were considered to meaningfully reflect the true spatial continuity in the data, random function models were fitted to the experimental variograms calculated in the direction of best continuity and 90 degrees to that direction in the 2D plane. The face sample data for E Lode, W Lode and NE Lode had sufficient sample pairs to calculate meaningful variograms. The parameters for the fitted models are given in Table 14-9.

There were insufficient sample pairs to compute meaningful variograms for the drillhole database. Therefore, the face sample variogram parameters are also used for resource estimation with the drillhole dataset.

To ensure accurate estimation of grade after dividing the estimated accumulations by the estimated thickness, the same estimation parameters are used for both the accumulations and true thickness.

AMC has estimated the thickness and grade in NW Lode, P Lode and Cuffley lode using the inverse distance estimation method, because there were insufficient sample pairs for meaningful variography in these lodes.

For details of the C Lode and Brunswick Mineral Resource variographic parameters, the reader is referred to the 2009 NI43-101 Technical Report (Fredericksen, 2009).



Table 14-9: Face Sample Fitted Model Variogram Parameters

Variable Lode Nugget Rotations 1st spherical structure 2nd spherical structure

X Y Z Major Semi-Major Minor Spatial

Variance Major Semi-Major Minor Spatial

Variance Thickness E 0 50 0 0 24 10.5 9 0.68 56.5 23 29.5 0.32 Au accum E 0.18 50 0 0 12.5 13.5 6.5 0.52 55 35 25 0.3 Sb accum E 0.19 50 0 0 19.5 12 10.5 0.47 46.5 23.5 25.5 0.34 Thickness W 0 0 0 0 7.5 4.5 3 0.52 36.5 24 23 0.48 Au accum W 0.35 0 0 0 25 7 7 0.4 27.5 17 17 0.25 Sb accum W 0.21 0 0 0 11.5 10 9.5 0.54 50.5 27.5 23 0.25 Thickness NE 0 -30 0 0 33 9.5 6 0.47 45.5 13 13.5 0.53 Au accum NE 0.3 -30 0 0 33 27 2 0.27 43.5 27.5 20 0.43 Sb accum NE 0.1 -30 0 0 28 18 4.5 0.61 54.5 19.5 30.5 0.29

14.8 Estimation Domain Boundaries Interpreted 2D structural boundaries were supplied by Mandalay for W lode and NE lode, to constrain the interpolation of thickness and grade during resource estimation. These boundaries were generated to take into account grade discontinuities observed in underground mapping on W lode and an interpreted fault-lode intersection in NE lode. In addition, AMC generated another estimation domain boundary for NE Lode and split the Cuffley lode into four estimation domains. This was done to limit grade smearing along strike and to improve the agreement between the estimated grades in the block model and the drillhole composite grades.

The W lode, NE Lode and Cuffley lode estimation domain boundaries are hard boundaries as shown in Figure 14-10, Figure 14-11 and Figure 14-12 respectively. The domains were used to code the samples and blocks. Each domain was estimated independently.

Figure 14-10: W Lode Estimation Domain Boundaries Note: Black dots are sample locations, Blue lines are estimation domain boundaries.



Figure 14-11: NE Lode Estimation Domain Boundaries Showing Three Domains Note: Black dots are sample locations, Blue lines are estimation domain boundaries.

Figure 14-12: Cuffley lode Estimation Domain Boundaries Showing Four Domains Note: Black dots are sample locations, Blue lines are estimation domain boundaries.

14.9 Block Model Estimation Grade accumulation and true thickness were estimated into 2D block models, whose cell centroids had an arbitrary Easting of 15,400 mE. The 2D estimates were run independently for the face samples and diamond drillhole samples using different cell sizes, resulting in two models. The cell sizes were selected based on the sample spacing of each data set. Sub-cells were used in the Y and Z directions to better define the mining depletion and domain boundaries. The specifications for each model are given in Table 14-10 and Table 14-11. For details of the C Lode and Brunswick Mineral Resource block model parameters, the reader is referred to the 2009 NI 43-101 Technical Report (Fredericksen, 2009).



Table 14-10: Face Sample Block Model Dimensions

Block model origin Block dimension (m) Number of cells

X 15399.5 1 1 Y 4000 2.5 500 Z 800 5 80

Table 14-11: Drillhole Sample Block Model Dimensions


X 15399.5 1 1 Y 4000 20 55 Z 800 10 40

After the estimation of thickness, gold accumulation, and antimony accumulation took place, the sub-celled models were regularised to the dimensions of the face sample block model, and expanded in the X direction to assist with mine planning. The specifications of the regularised models are given in Table 14-12.

Table 14-12: Mine Planning Regularised Block Model Dimensions


X 15250 10 30 Y 4000 2.5 650 Z 800 5 100

14.10 Estimation Parameters True thickness, gold accumulation, and antimony accumulation were estimated in the 2D vertical plane using ordinary kriging for the E, W and NE Lodes, and inverse distance squared for the NW, P and Cuffley lodes. The search parameters are listed in Table 14-13. For details of the C Lode and Brunswick Mineral Resource estimation parameters, the reader is referred to the 2009 NI43-101 Technical Report (Fredericksen, 2009). 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 15,400 mE;

• The orientation of the long axis of the search ellipsoid for each lode matched the maximum continuity observed in the face sample variography;

• The anisotropy of the search ellipsoid for each lode was guided by the anisotropy observed in the variography. Cuffley lode anisotropy was assessed visually as there were insufficient sample pairs for meaningful variography. The long axis of the search ellipsoid dipped 45° north in domains south of 4900 mN and 45° south in domains north of 4900 mN.

• For the E, W and NE Lodes, two separate estimates were made using channel samples and drillhole samples, using three estimation passes. For NW, P and Cuffley lodes, only one estimate was made using combined drillhole and face sample data in four estimation passes;

• Octants were used in the first three passes and then switched off for subsequent passes;

• Discretization was set to 4 by 4 in the Y and Z directions;



• After estimation, the results of the face sample estimates for the E, W and NE Lodes were overprinted on the drillhole sample estimate for each lode, applying a prioritisation to the face sampling data. Lode volumes were estimated for each block by applying dip and dip direction corrections to take account of the estimate being run in a 2D plane. Lode tonnes were estimated for each block by applying the formula specified in Section 14.6; and

• The resource was based on a minimum true lode width of 1.8 m, which is the practical minimum mining width applied at the Augusta Mine. For blocks with widths less than 1.8 m, diluted grades were estimated by adding a waste envelope with zero grade and 2.65 t/m3 bulk density to the lode.

Table 14-13: Search Parameters

Search Pass Lode

Search distance Min no. samples

Max no. samples

Octant search

Min octants

Min samples

per octant

Max samples

per octant

Long axis Short axis

1

E 40 20 4 16 on 2 1 2

W 40 30 4 16 on 2 1 2

NE 40 20 4 16 on 2 1 2

NW 40 20 4 16 on 2 1 2

P 40 20 4 16 on 2 1 2

Cuffley 100 20 4 16 off - - -

2

E 80 40 4 16 on 2 1 2

W 80 60 4 16 on 2 1 2

NE 80 40 4 16 on 2 1 2

NW 80 40 4 16 on 2 1 2

P 80 40 4 16 on 2 1 2

Cuffley 200 40 4 16 off - - -

3

E 120 60 2 16 on 2 1 2

W 120 90 2 16 on 2 1 2

NE 120 60 2 16 on 2 1 2

NW 120 60 2 16 on 2 1 2

P 120 60 2 16 on 2 1 2

Cuffley 500 100 1 16 off - - -

4

E 40 20 4 16 off - - -

W 40 30 4 16 off - - -

NE 40 20 4 16 off - - -

NW 40 20 4 16 off - - -

P 40 20 4 16 off - - -

5

E 80 40 1 16 off - - -

W 80 60 1 16 off - - -

NE 80 40 1 16 off - - -

NW 80 40 1 16 off - - -

P 80 40 1 16 off - - -



14.11 Block Model Validation The resource estimate was validated by visual comparison of the sample thickness and grades with the estimated grades in long section and by plotting the sample and block estimated true thickness, gold and antimony grades on swath plots. Examples are given in Figure 14-13 to Figure 14-24, which compare the face samples for each lode with the block models that were estimated using only face samples.

Swath plot orientations were chosen as N-S, which is along the strike of the lode and in the vertical direction, which is down-dip. No E-W swath plots were produced due to the restricted nature of the deposit across strike. Preparation of the swath plots included:

• True thickness in the block model being weighted by volume;

• Thickness in the composites being weighted by declustered weights;

• Gold and antimony grades in the face samples and the drillholes being weighted by true thickness; and

• Gold and antimony grades in the block model being weighted by tonnes.

Figure 14-13 to Figure 14-24 demonstrate good agreement between the face samples and block model values estimated using only the face samples for the E and W Lodes.

Figure 14-13: Swath Plot Comparing E Lode Face Samples with Face Sample Block Model Estimated True Thickness by Northing



Figure 14-14: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated True Thickness by Elevation

Figure 14-15: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Northing



Figure 14-16: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Elevation

Figure 14-17: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Northing



Figure 14-18: Swath Plot Comparing E Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Elevation

Figure 14-19: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated True Thickness by Northing



Figure 14-20: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated True Thickness by Elevation

Figure 14-21: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Northing



Figure 14-22: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Gold Grade by Elevation

Figure 14-23: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Northing



Figure 14-24: Swath Plot Comparing W Lode Face Samples with Face Sample Resource Block Model Estimated Antimony Grade by Elevation

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 Brunswick processing plant head grade during 2011 and 2012. The classification criteria are the following:

• The Measured Resource is located in, and is defined by the developed areas of the mine. Generally this criteria means that the estimate is supported by underground channel sampling;

• The Indicated Resource is located where drilling is based on a minimum of 40 m by 40 m in long section; 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 AMC and reported in March 2012 (Snowden 2012). However, there is one exception. Previously, there was a requirement for mining to have begun on a lode before any part of the lode could be classified as Indicated. AMC believes that this criterion can now be revised in light of Mandalay’s accumulated knowledge of the Costerfield mineralisation and structure.

14.13 Mineral Resource The Augusta and Brunswick deposit Mineral Resources are stated here with at an effective date of 30 April 2013. The Cuffley lode Mineral Resource estimate is updated in this report with newer drillhole data and is stated with an effective date of 30 June 2013.



The Augusta and Brunswick deposits consist of a combined Measured and Indicated Mineral Resource of 506,000 tonnes at 6.6 g/t gold and 3.2% antimony, and an Inferred Mineral Resource of 407,000 tonnes at 3.7 g/t gold and 2.2% antimony. This Mineral Resource includes the ROM stockpile, and is depleted for mining up to the 30 April 2013. The figures also exclude remnant parts of the Resource that are no longer accessible, and have been sterilized by mining.

The Cuffley lode consists of an Indicated Mineral Resource of 133,000 tonnes at 16.9 g/t gold and 5.4% antimony, and an Inferred Mineral Resource of 273,000 tonnes at 10.4 g/t gold and 3.2% antimony.

The Mineral Resources are reported at a cut-off grade of 3.6 g/t gold equivalent (AuEq), with a minimum mining width of 1.8 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) + 2.22 x Sb (%)

The cut-off grade has decreased from 4.7 g/t AuEq, and the relative value attributed to antimony has increased from 2.02 since the 31 December 2012 Mineral Resource estimate (SRK, 2013). 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 31 December 2012 for the E, W, NE, NW, and P lodes, and as of 30 June 2013 for the Cuffley lode were used to inform the estimate. The estimation methodology followed the approach used by AMC in the March 2012 Resource estimate reported in a Snowden NI 43-101 Technical Report on the Property (Snowden, 2012).

Mineral Resources for C Lode and 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 C Lode and Brunswick Mineral Resources are restated here for consistency with the other Costerfield lodes with an updated cut-off grade of 3.6 g/t AuEq and a 1.8 m minimum mining width criterion. The effective date for the C lode and Brunswick Mineral Resources is 30 April 2013. AMC has reviewed the work carried out in 2009 and Dr A Fowler takes Qualified Person responsibility for the restated Mineral Resources.

The ROM stockpile on the 30 April 2013 is included in the Augusta Mineral Resource statement. The tonnes of the stockpile are based on survey information. Stockpile grades were averaged from face samples that were sampled from the areas of the mine that contributed to the stockpile.

Details of the Augusta and Brunswick Mineral Resources are stated in Table 14-4. The Cuffley Lode Mineral Resource is stated in Table 14-5. Longitudinal projections are displayed in Figure 14-25 to Figure 14-32 with the same scale and legend to enable comparison.



Table 14-14: Augusta and Brunswick Mineral Resource Summary

Lode Resource Category Tonnes Au (g/t) Sb (%) Au Eq (g/t) E lode Measured 86,000 5.2 3.7 13.5 Indicated 46,000 2.7 1.2 5.4 E lode Total 133,000 4.3 2.9 10.7 W lode Measured 52,000 9.7 5.6 22.1 Indicated 73,000 4.2 1.7 8.1 W lode Total 125,000 6.5 3.4 14.0 NE lode Measured 26,000 7.7 5.1 19.0 Indicated 165,000 7.6 2.9 14.1 NE lode Total 190,000 7.6 3.2 14.8 NW Lode Measured 7,000 7.4 3.8 15.9 P Lode Measured 12,000 7.6 3.3 14.9 Brunswick Indicated 39,000 9.0 3.4 16.6 Other lodes Total 57,000 8.5 3.5 16.2 Stockpiles Measured 1,000 7.3 3.8 14.4 E lode Inferred 36,000 2.5 1.2 5.1 W lode Inferred 45,000 2.7 1.4 5.9 C lode Inferred 50,000 3.9 1.8 7.9 NE lode Inferred 155,000 3.6 2.5 9.1 Brunswick Inferred 122,000 4.6 2.6 10.5 Total Measured + Indicated 506,000 6.6 3.2 13.7 Total Inferred 407,000 3.7 2.2 8.6

Notes to Table 14.14: 1. Mineral Resource stated as of 30 April 2013. 2. Mineral Resource used relevant sample data available as of 31 December 2012 3. Mineral Resource stated according to CIM guidelines. 4. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. 5. A 3.6 g/t AuEq cut-off grade is applied where AuEq is calculated at a gold price of USD1,400 per troy ounce and an


Table 14-15: Cuffley Lode Mineral Resource on 30 June 2013


Indicated 133,000 16.9 5.4 29.0

Inferred 273,000 10.4 3.2 17.4

Notes to Table 14.15: 1. Mineral Resource stated as of 30 June 2013. 2. Mineral Resource used relevant sample data available as of 30 June 2013 3. Mineral Resource stated according to CIM guidelines. 4. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. 5. A 3.6 g/t AuEq cut-off grade is applied where AuEq is calculated at a gold price of USD1,400 per troy ounce and an




Figure 14-25: E Lode Mineral Resource Estimate Longitudinal Projection Note: Black dots represent face and diamond drillhole samples used in the resource estimate. The mining exclusion zone is due to a crown reserve at this location.

Figure 14-26: W Lode Mineral Resource Estimate Longitudinal Projection Note: Black dots represent face and diamond drillhole samples used in the resource estimate.



Figure 14-27: NE Lode Mineral Resource Estimate Longitudinal Projection Note: Black dots represent face and diamond drillhole samples used in the resource estimate.

Figure 14-28: NW Lode Mineral Resource Estimate Longitudinal Projection Note: Black dots represent face and diamond drillhole samples used in the resource estimate.



Figure 14-29: P Lode Mineral Resource Estimate Longitudinal Projection Note: Black dots represent face and diamond drillhole samples used in the resource estimate.

Figure 14-30: Cuffley Lode Mineral Resource Estimate Longitudinal Projection Note: Black dots represent diamond drillhole samples used in the resource estimate.



Figure 14-31: C Lode 2009 Mineral Resource Estimate Longitudinal Projection Note: Black dots represent diamond drillhole and open pit channel samples used in the 2009 Resource Estimate.

Figure 14-32: Brunswick 2009 Mineral Resource Estimate Longitudinal Projection Note: Black dots represent diamond drillhole samples used in the 2009 Resource Estimate.



14.14 Cut-off Grade Calculations The 3.6 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. Further details can be found in Section 15: Mineral Reserve Estimates.

14.15 Reconciliation Reconciliation results for dry tonnes delivered to the Brunswick Processing Plant, gold head grade (g/t), and antimony head grade (%) are plotted in Figure 14-28, Figure 14-29 and Figure 14-30 respectively.

Reconciliation results show reasonable agreement between the resource block model and the processing plant during 2012, considering the unquantified errors that influence this result. Such errors include stockpiling, ore-waste misallocation, and unplanned dilution. Over the period, the ounces of gold predicted by the model were 14% lower than produced by the plant. The tonnes of antimony predicted by the model were 3% lower than produced by the plant. The agreement gives confidence to the sample collection procedures, the quality of the assays and the resource estimation methodology.

Figure 14-33: Mill Reconciliation – Dry Tonnes



Figure 14-34: Mill Reconciliation – Gold (g/t)

Figure 14-35: Mill Reconciliation – Antimony (%)



14.16 Comparison with Previous Mineral Resource Estimate Table 14-15 compares the current estimate with the last estimate, which had an effective date of 31 December 2012. The table shows the combined effects of more level development and changes in the definition of sterilised ground in the period between estimates. AMC makes the following observations:

• Ongoing mining during the period has removed some of the previously stated Mineral Resource in E, W and NE lodes;

• Changes in the definition of sterilized ground has added to the Mineral Resource in E and W lodes;

• The E, W and NE Lode Mineral Resources have not been re-estimated between December 2012 and April 2013, so additional drilling has not influenced the result; and

• Additional infill drilling on the Cuffley lode and subsequent re-estimation has lowered the grade of the Cuffley lode overall, but the Indicated Mineral Resource has increased by approximately 30% at the expense of the Inferred Mineral Resource in terms of contained metal.

Note in Table 14-5 both Mineral Resource estimates are reported according to the 31 December 2012 parameters to enable a direct comparison. These are: using a cut off 4.7 g/t AuEq, an AuEq factor of 2.02, with a minimum mining width of 1.8 m.

14.17 Other Material Factors AMC 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.



Table 14-16: E, W, NE and Cuffley lode Current Estimate Compared with the 31 December 2012 Mineral Resource Estimate

Lode Resource Category Tonnes Dec 12

Tonnes April 13

Percent difference

Au (g/t) Dec 12

Au (g/t) April 13

Percent difference

Sb (%)Dec 12

Sb (%)April 13

Percent difference

E Measured + Indicated 85,000 97,000 12 % 4.4 5.3 17 % 2.9 3.5 18 %

Inferred 20,000 16,000 -25 % 3.1 3.1 -1 % 1.4 1.4 1 %

W Measured + Indicated 125,000 110,000 -14 % 7.5 7.0 -7 % 4.0 3.7 -8 %

Inferred 20,000 25,000 20 % 3.4 3.4 -1 % 1.7 1.7 1 %

NE Measured + Indicated 187,000 179,000 -4 % 8.2 8.0 -2 % 3.4 3.3 -2 %

Inferred 120,000 122,000 2 % 4.1 4.1 0 % 2.9 2.9 -2 %

Cuffley Measured + Indicated 79,000 79,000 41 % 20.8 16.9 -23 % 6.4 5.4 -17 %

Inferred 299,000 249,000 -20 % 10 11 12 % 3.8 3.3 -13 % Note: 1. Both Mineral Resource estimates are reported in this table above a cut off 4.7 g/t AuEq, using an AuEq factor of 2.02, with a minimum mining width of 1.8 m to enable comparison. 2. Tonnes have been rounded to the nearest 1000 t. 3. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.



15 Mineral Reserve Estimates No current Mineral Reserves have been calculated for the Cuffley lode. Previously Mineral Reserves have been published by Mandalay for the Augusta Mine. SRK Consulting (Australasia) Pty Ltd. March 2013, NI 43-101 report filed 28 March 2013.

No additional Mineral Reserves are being presented for Augusta Mine.



16 Mining Methods 16.1 Introduction

Mandalay plans to mine the extension of the Augusta Underground Mineral Resources to the 856 mRL and the Cuffley Lode, refer Figure 16-1.

Figure 16-1: Schematic of August and Proposed Cuffley Underground Design

The Augusta Mine incorporating the Augusta Lodes and Cuffley lode is an underground antimony-gold mine currently producing approximately 100 ktpa of mill feed from a variety of mining methods to a depth of over 310 m below surface. The peak trucking rate to achieve the production schedule proposed is 108 ktpa and this has been determined to be well within the capability of the current trucking fleet.

The Augusta Mine is serviced by a decline haulage system. The Augusta decline has been developed from a portal within a box-cut with 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 Level 2 was completed with a road-header. The decline provides primary access for personnel, equipment and materials to the underground workings.

The decline haulage length is approximately 2,300 m. The lower working levels, namely 960 mRL, 940 mRL, 920 mRL, and 900 mRL, are the most active area of the mine, this will change as mining progresses and sub-levels are closed off and new sub-levels are developed.

The Augusta Mine employs predominantly air leg longhole stoping methods as well as longitudinal uphole retreat working a bottom up sequence. These mining methods have been utilised throughout 2010, 2011 and 2012. CRF is placed into stoping voids to maximise extraction and assist with mine stability.



Mandalay proposes to develop the Cuffley Lode whose southern Inferred Resource boundary is located approximately 250 m to the northwest of the Augusta Mine. Access to the Cuffley lode will be via a single decline that will connect to the existing Augusta decline at the 1030 mRL. The Cuffley Lode is scheduled to produce, in conjunction with the Augusta Mine, at a maximum rate of 108 ktpa. Three different mining methods; full face development, longhole air leg CRF stoping and half upper air leg stoping will be utilised. All mined material will be transported to the Augusta box-cut before being hauled to either the Brunswick ROM Pad or Augusta Waste Rock Storage Facility.

Production from Augusta begins to decline from June 2014 because production from Cuffley increases, refer Figure 1,m 6-2.

Figure 16-2: Combined Augusta and Cuffley Production Profile

The mining schedule created for the Augusta Mine, in this Technical Report, includes Inferred Mineral Resources.

As part of the PEA, the volume of material contained within the designed shapes has been calculated and is referred to as the Proposed Mill Feed. The Proposed Mill Feed, derived from the application of modifying factors (for mining recovery and dilution) to Measured, Indicated and Inferred Resources is presented in Table 16-1.

The proposed mill feed was estimated using a cut-off grade of 4.7 g/t Au Eq.

This Proposed Mill Feed is a subset of the available Mineral Resource.

The assumptions and design basis for the Augusta Mine and Cuffley lode are presented in this section.

-

2,000

4,000

6,000

8,000

10,000

12,000

14,000

16,000

Aug-

13

Oct

-13

Dec-

13

Feb-

14

Apr-

14

Jun-

14

Aug-

14

Oct

-14

Dec-

14

Feb-

15

Apr-

15

Jun-

15

Aug-

15

Oct

-15

Dec-

15

Feb-

16

Apr-

16

Jun-

16

Aug-

16

Oct

-16

Dec-

16

Feb-

17

Apr-

17

Jun-

17

Tonnes per month

Tonnes Milled Cuffley Tonnes Milled - Augusta



Table 16-1: Costerfield Proposed Mill Feed

Proposed Mill Feed (tonnes)

Gold Grade (g/t)

Antimony Grade (%)

Inferred Material (%)

Augusta 163 6.7 3.4 28 Cuffley 345 12.0 3.8 64 Costerfield 509 10.3 3.6 55

Note: Proposed Mill Feed rounded to the nearest thousands (rounding error in summations)

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 air leg (hand held drill);

• Flat back overhand ‘cut-and-fill’ development;

• Floor benching of level development;

• Crown ‘half upper’ stopes;

• Longhole stopes with loose rock filled and varying level interval spacing; and

• Longhole stopes with CRF and fixed sub-level spacing.

The chosen mining method for future stoping is longhole stoping with CRF and fixed sub-level spacing. This is the primary method in use for current levels from the 1040 mRL to the 920 mRL. CRF and extensive cable bolting of the hangingwall on all levels prior to stoping provides the required support of the hangingwall and therefore rib pillars are unnecessary. To ensure medium term stability of the hangingwall, crown/ sill pillars are left in place and extracted on the completion of a stoping panel as half upper stopes utilising hand held mining techniques.

Numerical modelling using the program ‘FLAC3D’ was undertaken in March 2012 to examine the geomechanical performance of the planned longhole 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 hangingwall of the open stopes, was calibrated to the in situ conditions whereby excessive hangingwall 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 hangingwall failure is limited to individual open stopes and extends to a depth of approximately 1 m to 3 m. The CRF pillars are predicted to remain stable throughout the extraction process.

16.2.2 Cuffley Overview The Cuffley lode 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 lode 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 lode. The six geotechnical domains of interest are outlined below and shown in Figure 16-3. 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 positioned;

• 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 5010 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 hangingwall 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 approximately 3 m, is valid; refer to Table 16-2.

The final stage of crown/ sill pillar extraction consists of ‘half upper’ open stope retreat utilising hand held mining techniques.

The early years of mining history at Augusta (above 1075 mRL) has left behind what is termed as ‘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.

A central mine access pillar has been planned for the majority of stopes remaining below the 1049 mRL. The primary reason for the use of a central access pillar is to provide a stable 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). Table 16-2 lists this method as Option A.

A trial stoping area was completed in February 2013. This involved a single development drive advance direction with an increased level interval of 10 m. This spacing has the advantage of minimising stresses on the access pillar because stoping will only be retreated in one direction towards the access. This is opposed to the current stoping methodology that results in a shrinking central pillar due to stoping being retreated in both directions towards the level access. Table 16-2 lists this method as Option B.

Temporary and long-term sill pillar placement is important for minimising stress affects within the local mine environment. Stope dimension design and development level interval spacing is critical to minimise increases in mine stresses in the future.

The term ‘hydraulic radius’ (HR) is used to characterise the size and shape of stope surfaces and is a function of the area over perimeter of a given stope surface. The area and perimeter are directly proportional to the selection of level interval spacing. For example, the levels from 1049 mRL down are spaced approximately 7 m apart (refer to Table 16-2, Option A). This results in a development extraction ratio of approximately 40%. The longhole stope dimensions are designed with an 8 m



strike length, stability empirical rule. The 7 m level interval in conjunction with 3 m-high production drives results in a hangingwall HR of 2.22, as shown by equation one:

Equation One: HR(Hangingwall span 7 + 3) = (10 x 8)/(2x10)+(2 x 8) = 2.22

If the level interval is increased to 10 m and a 7 m strike length, stability empirical rule adopted (Refer to Table 16-2, Option B), a hangingwall HR of 2.28 is calculated, as shown by equation two:

Equation Two: HR(Hangingwall span 10 + 3) = (13 x 7)/(2x13)+(2 x 7) = 2.28

A level spacing of 10 m results in a development extraction sequence ratio of approximately 30% for Option B. This is preferred because it has been observed that ground conditions, through the development process, deteriorate the closer the level interval spacing. One main advantage of the closer level spacing has been the opportunity to support the mid span of stoping blocks.

A longhole stoping performance review and cost analysis from a trial stoping area will be beneficial to provide a sound basis for establishing the final future level spacing design.

Table 16-2: Main Mine Design Parameters

Pillar/ Development Type Dimension

Development level spacing Option A – 7.0 m (8.0 m strike length) 1.8 mW 2.8 mH Option B - 10.0 m (7.0 m strike length) 1.8 mW x 2.8 mH

Crown pillar dimensions and spacing 3.0 mW x 3.0 mH (21.0 m vertical spacing) Dimensions of central access pillar

Option A - shrink retreat 30 m from access centroid Option B - single retreat direction 15 m from access centroid

Rib pillar dimensions 2.5 m true strike length Sill pillar dimensions 2.0 mW x 3.0 mH 3.0 m dia. Ventilation Shaft pillar size 10 – 15 m from capital development Decline stand-off Augusta 50 m Decline stand-off Cuffley 40 m

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.

At Cuffley ground support requirements have been assessed based upon two methods, the empirical Barton and Grimstad design chart and discrete wedge analysis, using the identified hanging wall structures. Based on this assessment, it appears that the geotechnical conditions at Cuffley will be similar to those experienced at Augusta resulting in the current Augusta Mine ground support standards for capital and production drive development being adopted at Cuffley. 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.

It has been common practice at the Augusta Mine to select ‘black steel’ (non-galvanised) ground support products for operating development. This ground support practice will be adopted for all Cuffley production drive development.

16.2.6 Decline Location The current decline dimensions are 4.8 m high x 4.5 m wide with a gradient of 1 in 7 down. In the upper levels of the mine, the decline was mined at 4.5 m high x 4 m to 4.4 m wide. The decline dimensions have been increased to accommodate the installation of ground support by a jumbo as opposed to hand held methods which was the case in the upper levels of the decline.

The current design stand-off ranges from 110 m to 170 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.

Rock bolts consisting of resin, friction and cable bolts as well as mesh, are the recommended ground support for the establishment of access cross cuts. Major structures encountered by the decline are supported by the use of ‘in-cycle’ fibrecrete with cable bolts for additional support.

16.2.7 References The geotechnical content contained within Section 16 is based on work undertaken by Beck Engineering Pty Ltd. The review 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 a PEA; 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 Longhole 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 stoping method requires stoping to commence immediately once level development is completed. 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 air leg mining method.

A flat back mining method will be employed in rare cases such as remnant mining of uncompleted mining areas.

16.3.2 Method Description Augusta mines five individual lodes (W, NE, NW, P and E Lode) which have widths varying from 0.1 m to 1.5 m and dip of 45° to 85° which are favourable conditions for longhole 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 hole lengths. New areas of Augusta north have been developed using 10 m floor to floor level spacing.

The production cycle for longhole CRF stoping, as illustrated Figure 16-4, comprises the following:

• Develop access to the orebody;

• Establish bottom sill drive and upper fill drive;

• Drill air-leg blast-holes 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 obviates 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 Longhole CRF Stoping Method (after Potvin, Thomas, Fourie; 2005)

The half upper mining method cycle comprises the following:

• Drill 2.5 m - 3 m blind air leg longholes 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.

Figure 16-5 illustrates the Augusta half upper mining method. Stope geometry is engineered to negate the need to use remote control loaders for ore mucking.



Figure 16-5: Augusta Half Upper Mining Method (after Marshall, 2010)

The flat back method mining cycle comprise the following:

• Develop initial ore drive;

• Backfill ore drive with waste rock fill;

• Ramp up over the top of the rock fill;

• Strip ore into void between rock fill and ore;

• Extract ore and install ground support; and

• Commence next flat back cut.

Figure 16-6 illustrates the Flat Back mining method. Previously installed ground support is mined back through.



Figure 16-6: Augusta Flat Back Mining Method Source: Hutchinson, Diederichs; 1996

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 within a temporary waste rock storage facility. A portion of suitable material is screened and utilised underground for CRF. The remaining waste is either trucked across to the Brunswick site for capping the TSF or used in the construction of additional TSF lifts. Approximately 25,000 m3 will remain on the Augusta site to meet future rehabilitation commitments.

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 16-3.

Table 16-3: Mine Design Parameters

Item Size Gradient

Decline 4.8 mH x 4.5 mW 1:7 down

Decline sumps 4.8 mH x 4.5 mW 1:6 down

Drilling stockpiles 5.0 mH x 5.0 mW 1:50 up

Level cross-cut 5.0 mH x 5.0 mW 1:50 up

Truck tip 6.0 mH x 4.5 mW 1:50 up

CRF mixing bay 4.8 mH x 4.5 mH 1:10 down



Item Size Gradient

Ore stockpiles 4.8 mH x 4.5 mH 1:50 up

Level access 2.8 mH x 2.0 mW 1:5 up; 1:7 up; 1:7 down

Ore development 1.8 mH x 2.8 mW 1:50 up

Level spacing 21.0 m N/A

Sub-level spacing 7.0 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 the central access. This sequence enables a consistent production profile to be maintained as it allows for dual development headings on each level.

The mining and stoping sequence from 1049 mRL to the 1040 mRL was assessed by Itasca for potential stress problems using 3D elastic modelling. It was found that tensile stress within the central pillar was minimal in relation to the intact rock strength and the shear stress induced along structures parallel to the orebody was less than 2 MPa. This modelling has begun the process to confirm that the adopted mining sequence will ensure local stability within the Augusta Mine.

Stope design parameters used by Mandalay since 2010; (Refer to Table 16-2) have been utilised for designing stope shapes between 980 mRL and 870 mRL. It is recommended that additional stress modelling be undertaken to confirm the stability of the intended design within this region.

Figure 16-7 indicates the current Augusta mining methods and extraction sequence.

Figure 16-7: Long section of Augusta stoping between 1015 mRL and 989 mRL



16.4.3 Decline Development Access to the remaining Augusta Mineral Resources will be via a depth extension of the existing decline. At present, the decline is planned to the 870 mRL, approximately 320 m below surface.

A typical plan view of the Augusta decline and level development is shown by Figure 16-8.

Figure 16-8: Typical Plan View of Augusta Decline and Level Access Development

16.4.4 Level Development Ore drive development is designed to ensure the lode is positioned in the face in order to minimise the hangingwall exposure. Ore drives are excavated and supported by handheld mining methods that have been proven to be generally stable and productive. A typical cross section through the Augusta level development is shown by Figure 16-9.



Figure 16-9: Typical Section Looking North through Augusta Level Development

16.4.5 Vertical Development Ventilation rises of 2.4 m diameter will be excavated and supported between levels to extend the existing ventilation circuit to depth. Mining of these rises will be undertaken using ladder rising at an angle of 65° and on completion will have a 0.9 m diameter modular escape ladder way installed as a means of secondary egress.

In addition, a 3 m diameter raise bore from surface to approximately the 1020 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. This upgrade has enabled the Augusta primary ventilation system to be increased from the current fresh air component of 40 m3/s to approximately 80 m3/s.



16.4.6 Augusta Design Inventory The proposed underground stoping and development for the Augusta deposit is summarised in Table 16-4.

Table 16-4: Augusta Design Inventory

Level Development Stoping Total

Tonnes Gold (g/t)

Antimony (%) Tonnes Gold

(g/t) Antimony

(%) Tonnes Gold (g/t)

Antimony (%)

1 - - - 390 12.8 7.1 390 12.8 7.1

2 - - - 434 29.0 18.2 434 29.0 18.2

3 - - - 165 23.5 13.0 165 23.5 13.0

4 - - - 262 26.9 18.8 262 26.9 18.8

5 - - - 361 15.1 10.3 361 15.1 10.3

6 - - - 130 2.3 3.3 130 2.3 3.3

6 sub - - - 1,389 8.3 4.5 1,389 8.3 4.5

7 - - - 1,783 7.8 4.6 1,783 7.8 4.6

7 sub - - - 1,389 8.3 4.5 1,389 8.3 4.5

1070 - - - 4,732 4.6 3.9 4,732 4.6 3.9

1060 - - - 5,167 5.1 3.6 5,167 5.1 3.6

1040 - - - 3,120 3.4 2.5 3,120 3.4 2.5

1024 4,795 4.2 2.4 5,111 5.6 3.3 9,906 4.9 2.8

1005 1,303 3.2 2.2 6,368 7.9 4.6 7,671 7.1 4.2

984 7,590 5.7 2.4 7,838 8.4 3.6 15,428 7.1 3.0

980 - - - 1,025 4.3 2.6 1,025 4.3 2.6

960 - - - 11,119 10.1 5.1 11,119 10.1 5.1

959 7,279 4.8 2.5 9,354 7.0 3.6 16,633 6.0 3.1

940 4,859 9.1 2.3 12,104 10.4 4.9 16,962 10.0 4.1

920 10,547 7.1 2.7 15,930 8.1 3.4 26,477 7.7 3.1

900 15,268 5.2 2.5 11,362 7.8 3.5 26,631 6.3 2.9

880 19,599 4.2 2.6 14,450 5.8 3.5 34,048 4.9 3.0

Total 71,241 5.4 2.5 113,981 7.8 4.0 185,222 6.8 3.4 Source: 2013_May_PEA_Schedule-just-breakdown-sheet-inventory.xlsx ; cut-off grade 4.6 g/t AuEq

16.5 Cuffley Mine Design A central access pillar recovery sequence such as ‘bottom-up’ (end of mine) is recommended to reduce dilution potential in active stoping areas. The extraction sequence will highlight the formation of all permanent and recoverable crown pillars.

The proximity of the decline to the stoping blocks at Cuffley is approximately 120 m into the hangingwall. The final decline stand-off distance 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 design attempted to keep the access positions to each level as central as possible to aid extraction of the central access pillar.

Once the true extent of the Cuffley lode resource has been defined, further studies to confirm either a centrally located decline or southerly positioned decline are recommended to determine the final decline placement.



16.5.1 Mine Fill The practice of placing CRF in stope voids will be undertaken at Cuffley to improve local ground stability, reduce unplanned dilution and improve mining recoveries.

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.

16.5.2 Ventilation Shaft The proposed Cuffley exhaust shaft has yet to have any geotechnical drilling or analysis. This will be undertaken once decline development commences.

16.5.3 Additional Geotechnical Work Additional geotechnical tasks identified to improve the understanding of the Cuffley geotechnical environment include:

• Geotechnically log all Cuffley diamond drillcore;

• Examine the extent of the ‘East Fault’ across the Cuffley Deposit. Major faults such as this have previously caused significant capital development delays during the history of the Augusta Mine; and

• Obtain an appropriate level of detail of the hydrogeological environment before mining commences.

16.6 Cuffley Mine Design Guidelines

16.6.1 Design Parameters The mine design parameters that have been used in the design of the Cuffley LoM plan is summarised in Table 16-5.

Table 16-5: 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

Drilling stockpiles 5.0 mH x 5.0 mW 1:50 up

Level cross-cut 5.0 mH x 5.0 mW 1:50 up

Truck tip 6.0 mH x 4.5 mW 1:50 up

CRF mixing bay 4.8 mH x 4.5 mW 1:10 down

Production stockpiles 4.8 mH x 4.5 mW 1:50 up

Level access 4.8 mH x 4.5 mW 1:7 up; 1:7 down

Production development 1.8 mH x 2.8 mW 1:50 up

Level spacing 20.0 m N/A

Sub-level spacing 10 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 current Augusta mining methods and extraction sequence which is indicative of the development and stoping sequence that is planned at Cuffley.

16.6.3 Decline Development Pre-production development is proposed to comprise of a single decline being driven from existing Augusta development at the 1030 mRL. The decline is designed to be excavated at 1:7 down to provide initial access to production development and stoping blocks on the 965 mRL. Prior to level development being undertaken it is proposed that all the capital development associated with the Cuffley RAR and 965 mRL infrastructure level is completed.

Once the initial production level accesses are completed, the spiral decline is designed to be extended upwards and downwards from the 943 mRL, thus allowing multiple levels to be opened concurrently.

An isometric view of the Cuffley decline and level access development is shown in Figure 16-10.

Figure 16-10: Isometric of Cuffley Development

16.6.4 Level Development Production drive development is designed to be mined to ensure the lode is positioned in the face in order to minimise the hangingwall exposure. All production development would be 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.

16.6.5 Vertical Development Based on current designs, the Cuffley exhaust shaft is planned to extend vertically from surface to a depth of 215 m and will be 3.0 m in diameter to satisfy primary ventilation requirements. Excavation and support of the Cuffley exhaust shaft is scheduled to be prioritised once the decline reaches shaft bottom in order to establish primary ventilation.

North



Ventilation rises of 2.4 m diameter 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. Mining of these rises will be undertaken using ladder rising at an angle of 65°.

16.6.6 Stoping Stoping of the Cuffley lode is proposed to utilise the same stoping methods employed at Augusta, namely longitudinal longhole air leg stoping in conjunction with CRF and air leg half uppers. These stoping methods are expected to perform well based on the vertical dip of the Cuffley lode, which is similar to E Lode and steeper than W Lode in the Augusta Mine. Figure 16-11 illustrates a long section of the Cuffley production drive development and stoping panels.

Figure 16-11: Long Section of Cuffley stoping areas

16.6.7 Materials Handling All mineralisation and waste is proposed to be trucked from the mine via the internal Cuffley decline from an ultimate depth of over 300 m to the bottom ore development level situated at the 850 mRL. Material would then be trucked to Augusta via the 1030 mRL decline and hauled to surface via the existing Augusta Mine decline. It has been assumed that this will necessitate the addition of one additional haul truck to meet the Cuffley development and production requirements. A private contractor would rehandle the mill feed and transfer it to the Brunswick ROM pad to be stockpiled, screened, blended and crushed prior to be being fed into the Brunswick Processing Plant.

All other waste hauled to surface via the underground truck fleet is proposed to be stockpiled within a temporary waste storage facility. A portion of suitable material will be screened and utilised underground for CRF. The remaining waste is scheduled to be either trucked across to the Brunswick site for capping tailing facilities or used in the construction of additional TSF lifts. Approximately 25,000 m3 will remain on the Augusta site to meet future rehabilitation commitments.



16.6.8 Cuffley Proposed Design Inventory The planned stope and development inventory for Cuffley is summarised in Table 16-6.

Table 16-6: Cuffley Stope and Development Inventory

Level Development Stoping Total

Tonnes Gold (g/t)

Antimony (%) Tonnes Gold

(g/t) Antimony

(%) Tonnes Gold (g/t)

Antimony (%)

1050 347 5.30 2.52 - - - 347 5.30 2.52 1040 2,814 9.68 4.49 4,680 7.46 3.68 7,494 8.30 3.98 1030 3,874 10.38 4.73 6,378 13.66 6.40 10,252 12.42 5.77 1020 3,880 11.03 4.94 6,789 14.93 6.67 10,669 13.51 6.04 1010 4,219 14.58 4.64 6,790 21.27 6.77 11,008 18.71 5.95 1000 6,980 13.78 3.67 8,435 17.65 5.74 15,415 15.90 4.80 990 5,253 13.68 4.01 9,703 17.75 5.09 14,956 16.32 4.71 980 6,973 11.38 3.34 11,256 16.04 4.82 18,229 14.26 4.25 970 6,616 12.07 3.34 11,612 15.73 4.48 18,228 14.40 4.07 960 6,261 12.50 3.25 10,589 17.61 4.67 16,850 15.72 4.14 950 3,489 12.69 3.71 4,656 17.81 4.25 8,145 15.62 4.02 940 5,906 12.69 3.02 10,082 17.55 4.17 15,987 15.75 3.75 930 5,928 13.09 3.27 10,466 17.66 4.40 16,394 16.01 3.99 920 6,269 12.28 3.11 4,197 18.93 4.76 10,467 14.94 3.77 910 6,274 12.38 3.03 11,646 15.82 4.04 17,920 14.61 3.69 900 5,930 12.35 3.12 11,640 15.30 4.02 17,570 14.30 3.72 890 5,915 10.26 2.77 5,324 12.74 3.42 11,238 11.43 3.08 880 6,257 9.44 2.64 12,728 11.60 3.26 18,985 10.89 3.06 870 6,932 7.57 2.33 12,697 10.07 3.05 19,628 9.18 2.80 860 7,278 6.94 2.31 5,427 9.98 3.24 12,705 8.24 2.71 850 6,919 5.61 2.15 12,619 7.52 2.83 19,537 6.84 2.59 840 6,921 5.29 2.21 12,600 7.22 2.87 19,521 6.54 2.63 830 6,231 5.29 2.32 5,129 7.21 3.04 11,360 6.16 2.65 820 5,541 4.85 2.40 10,689 6.73 3.18 16,230 6.08 2.91 810 2,782 3.14 3.21 5,807 5.07 3.77 8,589 4.44 3.59 800 1,744 1.47 3.64 1,752 2.89 4.87 3,496 2.18 4.26

Total 137,530 9.98 3.12 213,688 13.35 4.18 351,217 12.03 3.76

16.7 Ventilation

16.7.1 Existing 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 1020 mRL and then a return air rise that extends to the surface from the 1020 mRL. Fresh air from the decline is force ventilated into the active levels using secondary ventilation fans.

Two single stage 132 kW axial fans that have been built into bulkheads at the 1020 mRL act as the mine’s primary ventilation fans. These fans draw a total of 80 m3/s of air through the mine.



16.7.2 Proposed Ventilation Circuit The proposed Cuffley lode development can be readily ventilated with a conventional shaft exhaust system supported by secondary ventilation for underground development. Based on the current Cuffley lode mine design and increased haulage distance, the planned diesel fleet will be slightly larger than Augusta and will require a minimum primary ventilation flow of 83 m3/s.

Table 16-7 summarises the current and future Cuffley lode diesel fleet requirements and estimated primary airflow requirements.

Table 16-7: Proposed Cuffley Diesel Fleet and ventilation requirements

Task Model Number Operating

Engine Power (kW)

Machine Airflow Requirement

(m3/s)

Task Airflow Requirement

(m3/s)

Development LHD 1700 1 231 11.6 11.6

Development LHD 1700G 1 258 12.9 12.9

Truck TH430 2 293 14.7 29.3

Production LHD LH203 4 72 3.6 14.4

Tool Carrier JLG307 2 75 3.8 7.5

Grader - planned 140H 1 138 6.9 6.9

Totals 1,067 82.6

The Cuffley ventilation system will be established as a three stage process:

• As at 30 April 2013, approximately 230 m of development will be necessary to reach the base of the proposed ventilation shaft. Ducting simulations indicate that this initial access development can be accomplished by using a 1254 mm diameter twin stage 75 kW axial auxiliary fan mounted in the Augusta decline with 1400 mm diameter low friction duct. The predicted development haulage fleet includes one Caterpillar 1700 G loader and one Sandvik TH430 truck. Predicted air flow out the end of the 230 m vent column is 30 m3/s;

• Due to the length of the proposed Cuffley Shaft and the unknown resource potential, the Cuffley Shaft has been designed to be raise-bored at a diameter of 3 m. This will provide a maximum exhaust capacity of up to approximately 140 m3/s; and

• The level connections will be via 3.0 m diameter rises developed by longhole rise methods.

Figure 16-12 shows the planned ventilation infrastructure for Cuffley and the connection to the current Augusta ventilation circuit.



Figure 16-12: 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 be applied to the ventilation calculations.

Table 16-8 summarises the current and future Augusta diesel fleet requirements and estimated primary airflow requirements.



Table 16-8: Augusta Primary Flow to Match Diesel Flow

Task Model Number Operating

Engine Power (kW)

Machine Airflow Requirement

(m3/s)

Task Airflow Requirement

(m3/s)

Development LHD 1700 1 231 11.6 11.6

Development LHD 1700G 1 258 12.9 12.9

Truck TH430 1 293 14.7 14.7

Truck M426 1 278 13.9 13.9

Production LHD 151 2 55 2.8 5.5

Production LHD LH203 2 72 3.6 7.2

Tool Carrier JLG307 2 75 3.8 7.5

Grader - planned 140H 1 138 6.9 6.9

Total

1,602

80.1 LHD – load-haul-dump

16.7.4 Cuffley LoM Ventilation The highest demand ventilation case for both Augusta and Cuffley (i.e. highest fan pressures and lowest air flow) will be when both mines are fully developed and operational. The resultant flows and fan duties of this occurrence are summarised in Table 16-9.

Table 16-9: Cuffley LoM Ventilation Operational Summary

Description Quantity Comments

Augusta Portal Flow 133 m3/s Air velocity 5.5 m/s, must not exceed 6 m/s

Total Underground Flow 159 m3/s Primary fans 448 kW total

Cuffley Total Exhaust 83 m3/s Fixed flow, can be improved by decommissioning Augusta fans

Augusta Fan Pressure 1890 Pa Based on 2 x Clemcorp CC1400 110 kW fans

Cuffley Fan Pressure 2280 Pa Predicted fan pressure using 83 m3/s fixed flow

High pressure fans for Cuffley will need to be sought in order to provide the required primary flows.

16.7.5 Rise Sizes The parameters of the designed Cuffley rises are in Table 16-10.

Table 16-10: Cuffley Rise sizes

Cuffley Rises Diameter

Cuffley RAR (linking levels from 961 mRL to 1017 mRL and 857 mRL) 3.0 m diameter

Cuffley RAR (961 mRL to Surface) 3.0 m diameter

16.7.6 Fan Duties The existing Augusta primary fans comprises two Zitron ZVN-1-14-132/4 twin stage 132 kW axial fans installed in a bulkhead on the 1020 mRL. The operating parameters of each fan are:

• Lower operating fan static pressure of 1000 Pa at 55 m3/s; and

• Higher operating fan static pressure of 2800 Pa at 35 m3/s.

At present, the fans are operating at a measured pressure of 2600 Pa for a combined volume capacity of 80 m3/s.



16.7.7 Raise Sizes The parameters of the existing rises are as follows:

• Augusta Fresh Air Rise (surface to 1020 mRL) 2.4 m diameter;

• Augusta RAR (below the 1020 mRL) 2.4 m diameter; and

• Augusta RAR (1020 mRL to surface) 3.0 m diameter.

Based on the 80 m3/s requirement, the main decline airflow will be at a maximum velocity of 4.2 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 main exhaust rise will have a velocity of approximately 12 m/s, which is at the lower limit that main exhaust shafts should be operated at. Hence, if additional ore is discovered down dip of the existing Mineral Resources, the main Augusta RAR from 980 mRL to surface can accommodate airflows up to 140 m3/s.

16.7.8 References The information presented in Section 16.7 is supported by a review undertaken by AJ Bruce Consulting Pty Ltd as reported in Costerfield Operations Ventilation Update, 9 August 2012.

The main objectives of the review are as follows:

• Determine the primary flow requirements for both Augusta and Cuffley;

• Re-examine the maximum depth that can be achieved with the current Augusta ventilation system;

• Propose improvements that can be made to the Augusta primary ventilation system to increase the maximum depth achievable with the current Augusta primary fans;

• Determine the size and location of a shaft(s) to successfully primary ventilate Cuffley; and

• Determine the size and location of primary fans to adequately ventilate both Cuffley and Augusta.

16.8 Backfill CRF is mixed in batches of 24 t of -150 mm aggregate (stockpiled on surface from waste material screened periodically by cartage contractor and carted to CRF mixing bays as required) with 1,200 kg ordinary Portland cement and 660 L water. The 24 t of aggregate is equivalent to 3 x R1700 loader buckets and is loaded into trucks from the stockpile using an R1700 loader. The truck then conveys the dry aggregate to a CRF mixing bay proximal to the stope been filled (mixing bays are located at each access). The hydrated cement mix is batched on surface using a batching plant supplied on contract by Mawsons Ltd and is conveyed to the mixing bays using a leased agitator truck specifically designed for use underground. Mandalay is investigating options for purchasing a suitable agitator truck.

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 bogger bucket prior to been trammed to the top 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 stope void. Care is taken during placement of the CRF that the rock-tube is not displaced which is secured by chains during 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 aggregate. 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 a 2,353 kVA (2,000 kW) 22 kV supply. The Augusta Mine reached the limit during the fourth quarter of 2012. The upgrade to power supply will be completed by Powercor in the second quarter of 2014, in the interim; a single 750 kVA diesel generator with an equivalent back up unit will be installed on surface to supply additional power requirements.

Power is reticulated between each level through the RAR system as opposed to the decline, thereby reducing the length of cable runs throughout the mine. High voltage (11 kV) power is fed underground via a service hole from surface to a 1.5 MVA 11 kV substation at the 1020 mRL. From this location, an additional service hole has been drilled to the 935 Sub Station Cuddy which will allow the high voltage cable to be lowered and connected to the 1.5 MVA 11 kV sub-station at this location. From these sub-stations, the power is reticulated to the working areas via cable and distribution boxes.

The Augusta Mine utilises a leaky feeder system for communications that will be extended as the mine develops into new production areas. Underground refuge chambers have a direct communications 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 via an existing 4-inch poly line located in the decline. The underground compressed air distribution system consists of 2-inch poly piping on each sub-level. Ring mains are installed where possible to increase the systems efficiency. Compressed air is used to power pneumatic equipment and / or activities including:

• Air leg drills;

• Pneumatic ammonium nitrate-fuel oil (ANFO) loaders;

• Blasthole cleaning for development rounds;

• Pneumatic longhole drills, and

• Longhole cleaning.



16.9.3 Process and Potable Water Process water for the Augusta underground and surface operations is sourced from the Augusta Mine Dam located at the entrance to the site. This dam acts as a holding pond for all the water pumped from underground until the water is pumped to the evaporation/ storage dams. Process water is supplied underground via the existing 4-inch diameter poly pipe located in the exhaust shaft and in the decline. Process water is distributed to the working levels via 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.

16.9.4 Explosives and Magazine All manufacturing, storage, sale, import, transport and use of explosives are conducted in accordance with the WorkSafe Dangerous Goods (Explosives) Regulations 2011.

Mandalay utilises its own licenced personnel and equipment to handle, store, transportation 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 1020 mRL. Table 16-11 states the current Augusta Magazine licence allowances.

Table 16-11: 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

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 1020 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. It is also planned to establish an agreement with a local crane contractor to enable a suitably sized crane to be mobilised to the collar of the Cuffley exhaust shaft for emergency egress purposes.

The ladder way design from Cuffley is stand alone and comprises of an enclosed ladder way within a 1.2 m 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 1040 mRL and down to the 810 mRL. Once personnel evacuate to the 958 mRL, emergency egress will be provided via an emergency pod that will be hoisted to surface by a crane located at the Cuffley exhaust shaft collar.

Refuge Chambers Underground refuge chambers are installed underground in response to hazards posed by irrespirable atmospheres. Augusta has a fresh air base located at Level 3, a 16 man refuge chamber at the 1050 mRL and 965 mRL and a 20 man chamber at the 985 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 16-13.

Figure 16-13: 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 16-14. 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 16-14: 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 is 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 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 and ground water is achieved by a staged pump station system. Currently comprising Weartuff WT084 mono pumps located at 1040 mRL, 1030 mRL and 960 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 Flygt pumps to direct water to the 960 mRL mono pumps. One of the 960 mRL mono pump lifts water to the 1030 mRL mono pump which pumps water to surface via the Level 5 rising main. The other 960 mRL mono pump lifts water to the 1040 mRL mono pump which pumps water to surface via the 1040 mRL rising main. This system will need to be extended as the mine progresses vertically. The mine owns an additional WT084 pump which can be used as a travelling mono pump, allowing it 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. It is thought that future dewatering of the RBA above Cuffley lode can be undertaken either through directly pumping groundwater from the Alison Shaft or targeting the Adder Fault via vertical groundwater bores.

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.

The Cuffley lode will be developed horizontally via access through the Augusta Mine at depth, it is likely to have high initial groundwater inflows during the initial development stage. The high initial groundwater flows are likely to decrease as the hydraulic pressure reduces and dewatering occurs. Where Augusta Mine is dewatered via sumps within the underground drives, Cuffley may have the potential to be dewatered through vertical groundwater bores.

To date, there has been no aquifer tests or groundwater sampling conducted directly within the Cuffley lode. However, groundwater has been sampled from many locations around Cuffley within the RBA. The chemistry is variable, but the salinity is within the range of 5,000 to 8,000 mg/L and pH values being neutral to alkaline at 7 to 8.5. The chemistry is mainly sodium-chloride type and has elevated arsenic and antimony, which is associated with the mineralisation.

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 Cuffley 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 cones of depression will merge between the two mines.



16.11.2 Dewatering Dewatering Cuffley internally is proposed to be via a rising main. It is expected that the Augusta dewatering has already removed a large amount of water and decreased the head pressure at Cuffley. Removal of process water and ground water will be achieved by a sole pump station located at the 960 mRL. The pump station will comprise of dual Weartuff WT084 mono pumps that have a capacity to each pump 240 m of total dynamic head with a total system flow of 25 L/s. The 960 mRL monos pumps will direct water to surface via a planned 960 mRL rising main (which will be situated within the exhaust shaft).

The lower working levels will utilise electric Flygt pumps to direct water to the 960 mRL mono pumps. This system will need to be extended as the mine progresses vertically. This will occur with the assistance of another currently owned WT084 which can be used as a travelling mono pump, allowing it to be positioned where required at either the Augusta or Cuffley deposit.

16.12 Modifying Factors

16.12.1 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 hangingwall rock that has been drilled and blasted to maximise mining recovery and/ or maintain favourable wall geometry for stability.

Due to the narrow width of the Augusta Lode mineralisation, the Mineral Resources includes an element of planned mining dilution as the Mineral Resource is reported to conform to a minimum 1.2 m mining width.

Unplanned dilution includes waste rock (or low grade material) and/ or backfill from outside the planned production drive profile or stope limits that overbreaks or sloughs 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 these figures. The longhole overbreak and dilution factors are considered to be conservative because every attempt is made to mine the stope as close to the lode width as possible.

The percentages of planned/ unplanned dilution for the various Augusta mining methods are shown in Table 16-12.



Table 16-12: Augusta Recovery and Dilution Assumptions

Mining Method Unit Air Leg Development

Air Leg Flat Back

Air Leg Half Uppers Longholes

Planned Height Augusta m 2.8 2.5 2.4 4.2

Planned Height Cuffley m 2.8 2.5 2.4 7.2

Planned Width m 1.8 1.8 1.2 1.2

Overbreak Dilution % 20 20 10 10

Fall Off Dilution % 5 10 10 10

Bogging Dilution % 0 5 10 5

Recovery Factor % 100 95 80 90

An additional dilution percentage is added to the Mineral Resources to reflect the dilution added during the bogging process. Bogging dilution for air leg half uppers has been assumed due to the practice of bogging up against the back of a waste bund. This dilution has been based on the geometry of the stope and what the loader is able to safely reach. Bogging dilution for air leg flat backs reflects the nature of bogging on a waste filled floor. The bogging dilution for air leg stopes is viewed as conservative because it is based on the previous method of firing and bogging against unconsolidated rock fill. The mine wide use of CRF for stope backfill is likely to result in this factor being improved; however, this is yet to be appropriately measured and therefore has not be incorporated into this Technical Report.

Mining Recovery Planned recovery reflects the percentage of the in‐situ Mineral Resource that will be accessed and developed. Reasons that some block model in‐situ Mineral Resource will not be recovered may include:

• The Mineral Resource includes a small volume that is separate from the main mining area and does not justify the cost to develop and mine; and

• Stoping areas within the block model that cannot be mined due to unfavourable economics.

The mining shapes designed are assessed against the block model to give lode tonnes and grade. These numbers are then processed through a resource to reserve converter that calculates all planned dilution, unplanned dilution and recovery factors dependent on mining method to determine the tonnage before entering into the schedule as shown in Table 16-12. In the schedule, 4.76 g/t AuEq cut-off grade has been applied to the Augusta Mineral Resources to identify and evaluate mining areas.

For Cuffley, two recovery factors have been considered in establishing the proposed mill feed used for the PEA.

Planned recovery reflects the percentage of the in‐situ Mineral Resource that will be accessed and developed. The reason that some of the block model in‐situ Mineral Resource will not be recovered includes:

• Stoping areas within the block model that remain as pillars at the end of the LoM. In the case of Cuffley it was assumed that 40% of the sill pillars would be recovered; and

• Bogging, to reflect this is a typical standard mining recovery factor of 90% was applied to the diluted Cuffley mining shapes.



16.12.2 Cut-off Grade Major assumptions made in estimating the design cut-off grade were based on past experience within the Augusta deposit as well as historical performances, both physical and economic, of the extraction and processing methodologies.

Based on initial cost estimates from the current Augusta operations a cut-off grade of 4.76 g/t AuEq was applied to the stope designs to determine zones with acceptable economic grade for 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. As this is a preliminary economic analysis, the cut-off grade that was applied is considered to be suitable for this level of study.

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.

Physical designs for the development and stope openings are applied to the Mineral Resource in a 3D to 2D mining software package (creating a thickness accumulated model). The physical parameters of the openings (length, position, volume, tonnes, ore thickness, density and grade) and Resource categories are exported to a tabulated electronic form in order to apply dilution factors and calculate resultant grade and tonnage.

Scheduling is then performed to evaluate economic values and test the designs. If development areas are not economic in themselves, the value of the associated stopes must justify the total stope development including the production drives.

16.13 Life of Mine Schedule

16.13.1 Development Schedule

All capital development in relation to the Cuffley lode is anticipated to commence in June 2013 and be completed during the December 2016. The final Augusta capital development is scheduled to be completed in June 2013. The capital development schedule is illustrated by Figure 16-15.

Figure 16-15: Capital Development

-

20

40

60

80

100

120

140

160

180

200

Capital Metres per month



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 is shown in Table 16-13.

Table 16-13: Jumbo Production Rates

Headings Available 1 2 3

1 Jumbo (m/mth) 140 160 180

The hand held development rates are based upon the assumptions as shown in Table 16-14. The waste development and production development are undertaken by different crews. The production development production rate is less than the waste development rate as it allows time for face sampling and poorer ground conditions.

Table 16-14: Hand Held 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

16.13.2 Production Profile

The mining strategy for the Costerfield operation is presented in Figure 16-6. This is a coarse schedule and for the purposes of this Technical Report it has not been smoothed. The production profile at the end of the mine life was compressed to eliminate months of negative cashflow resulting from a low production profile. Production in the last six months of the schedule was compressed into the preceding five months to maintain a production rate in the order of 10,000 tpmth. This was less than 5% of the total production and is considered to not be material to the outcomes of the report.



Figure 16-16: Costerfield Production Profile (proposed Mill Feed)

16.13.3 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 replace retired equipment or meet increased production demands.

Planned future equipment additions include a new haulage truck and grader. Planned equipment replacements include two 151 loaders, a production drill as well as a light vehicle per year for the remainder of the operation. The fleet upgrade is indicative only and will be reviewed upon finalisation of scheduling requirements.

The existing mobile equipment fleet is summarised in Table 16-15. It is envisaged that this fleet will largely remain unchanged over the remaining LoM.

Table 16-15: Augusta Underground Mobile Equipment Fleet

Equipment Type Existing Fleet

2-Boom Jumbo 2

Production Drill 2

LHD - 1700 Loader 2

LHD - 151 Loaders 6

Haulage Trucks 2

Charging Rig 1

Telehandler 2

Service Tractor 3

Light Vehicles 10

Total 27



16.13.4 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, 12-hour shift 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 16-16. It is expected the current staffing levels will be sufficient to successfully operate both the Augusta and Cuffley underground operations simultaneously.

Table 16-16: Personnel on Payroll by Department

Classification Number

Executive 3 Mining 111 Technical Services 23 Safety 3 Environmental 3 Administration 14 Exploration 13 Plant 26 Total 196

Table 16-17 lists the mining personnel required for each Augusta underground mining crew on single shift basis.



Table 16-17: Underground Shift Mining Personnel

Classification Number

Shift Supervisor 1

Jumbo Operator 2

Loader Operator 5

Magazine Keeper 1

Truck Driver 2

Longhole Driller 1

Air Leg Operator 6

Service Crew 2

Nipper 1

Charge Hand 1

Total 22

16.13.5 Labour Costs Labour costs have been estimated and include superannuation, workers compensation, payroll tax and partial allowances for leave accrual.



17 Recovery Methods 17.1 Brunswick Processing Plant

The currently operational Brunswick Processing Plant operates 24 hours/ day, 7 days/ week and processes a sulphide gold-antimony ore. Operation of the crushing plant is restricted to 10 hours/ day, to minimise the effects of noise on the surrounding neighbours. There is also only a 5 hour crushing shift on Saturday and no crushing on Sundays.

The surface crushing and screening system processes underground feed down to a particle size range suitable for milling through an existing ball mill circuit on site. With the inclusion of the portable crushing unit, the capacity of the plant has been successfully upgraded to 10,000 t/mth.

The processing plant comprises of a diesel powered mobile crushing plant which has 16 mm feed screen on recirculating load with an impact crusher. The primary crushed product then forms feed for the original two-stage crushing circuit which utilises the two fine ore bins, two milling stages in series with classification and gravity concentration in closed circuit. The flotation circuit comprises of a rougher, scavenger and single stage cleaner circuit for the production of an 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. Figure 17-1 illustrates the current Brunswick Processing Plant flow sheet.

17.1.1 Crushing and Screening Circuit The crushing and screening plant consists of primary stage crushing through a diesel powered impact crusher with 16 mm screen on recirculating load. This crushed ore is conveyed to two fine ore bins in parallel utilising the existing jaw crushing circuit as transportation of ore to these bins, each of 120 t capacity. The previously used two stages of jaw crushing circuit is still available for use for redundancy, the primary jaw set at 45 mm gap and the secondary intermeshing with intermediate screening. Both jaw crushers operate in open circuit. This delivers a product of approximately 50% passing 5 mm. A re-feeder 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 fed from the stockpile directly into the primary ball mill as needed.

17.1.2 Milling Circuit The fine ore is reclaimed from the bins or stockpile via feeders which discharge onto the primary mill feed conveyor belt as feed to the milling circuit. The milling circuit comprises of 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 mill and undersize being fed to a centrifugal gravity concentrator to recover a small mass high grade concentrate that is sent either to the concentrate thickener or to gold room for tabling. The gravity concentrate is sent direct to 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 sent 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 mostly during September 2012 to assess the impact it would have on the throughput of the existing processing plant. Results proved that this additional crushing capacity was able to improve throughput to 10,000 t/mth. A larger capacity unit (Finlay Terex) was employed on a hire



basis and eventually purchased, and has been successful in increasing the throughput tonnage. The current mill capacity is capable of handling +10,000 t/mth, which was achieved from October 2012 to December 2012 (Refer to Table 17-1). From January 2013 to April 2013 the processed tonnes were less than 10,000 tonnes a month due to availability of mill feed from the mine.

17.1.3 Flotation Circuit The flotation circuit is fed from the secondary mill closed circuit cyclone overflow. The cyclone overflow is fed to a conditioning tank where lead nitrate, an activator and potassium amyl xanthate, a collector are added. The conditioning tank feeds two site fabricated tank cells in series. The two tank cell concentrates combine with the cleaner concentrate in the concentrate thickener. The tank cell tailings then flow to the Denver cells.

The rest of the flotation circuit consists of eight Denver square DR100 cells for the remaining rougher and scavenging duties and four Denver square DR15 cells are used for cleaning duties. The concentrate from the rougher flotation cells is pumped to the cleaner flotation cells while the tailing becomes feed for the scavenger flotation cells. The concentrate from the scavenger flotation cells is recycled to the feed of the rougher flotation cells while the tailing is pumped to the tailing thickener. The concentrate from the cleaner flotation cells is pumped to the concentrate thickener while the tailing is recycled to the feed of the rougher flotation cells. Refer to Figure 17-2 for visual representation of this circuit.

17.1.4 Concentrate Thickening Filtration The final concentrate is then pumped from the thickener directly to the plate and frame filter. The moist filter cake is discharged directly into concentrate bags. The filtrate is recycled to the concentrate thickener while the water overflow from the concentrate thickener is recycled through the plant as process water.

17.1.5 Tailings and Circuit The flotation tails are settled out in a thickener, where the water overflow is recycled through the plant and the thickened solids are pumped to a tailings storage facility. The tailings are managed via spigots and the separated water from the tailings is also pumped back and recycled through the plant.

17.1.6 Recovery The current plant recovery on the existing Augusta ore feed is discussed in Section 13. In summary, despite the increase in production, antimony recoveries have been maintained at 96% and gold has lost only 1% from 90% to 89% due to the increase in throughput rates. This was achieved by processing similar grade material. The results are summarised in Table 17-1.

17.1.7 Concentrate Grade The current concentrate antimony grades are reasonably steady after the throughput increase with a drop of 1% to approximately 53% antimony.

Concentrate gold grades are more variable and depend on the feed grade and the gravity recoverable gold component of the feed.

With increased mill throughput since September 2012, the concentrate production has increased accordingly with antimony concentrate grade maintained, refer Table 17-2.



Figure 17-1: Brunswick Processing Plant Flow Sheet



Table 17-1: Brunswick Mill production results for Feb 2012 to Apr 2013

Table 17-2: Brunswick Mill concentrate production Feb 2012 to Apr 2013

Feb-12

Mar-12 Apr-12 May-

12 Jun-12 Jul-12 Aug-12

Sep-12 Oct-12 Nov-12 Dec-

12 Jan-13 Feb-13 Mar-13 Apr-13 Avg Feb-12 -

Aug-12 Avg Sep-12 - Apr-

13 Con DMT 467 578 597 629 529 566 455 691 818 785 673 787 782 716 678 546 741

Antimony Con % 53 53 54 55 55 56 55 54 55 55 54 53 54 52 52 54 53

Gold Con g/t 100 93 88 84 95 96 102 84 92 82 91 85 84 90 94 94 88

Test period Powerscreen Aug 31st to Oct 11th

Finlay Terex crusher in service from Oct 12th

Feb-12 Mar-12 Apr-12 May-12 Jun-12 Jul-

12 Aug-12

Sep-12 Oct-12 Nov-12 Dec-12 Jan-13 Feb-13 Mar-13 Apr-13 Avg Feb-12 -

Aug-12 Avg Sep-12 - Apr-

13 DMT Milled 5,929 7,392 7,269 6,648 6,894 6,833 6,979 9,299 10,151 10,242 11,523 9,984 8,243 8,351 9,906 6849 9712 Antimony recovery (%) 96 96 96 96 97 97 96 96 96 96 94 96 96 96 96 96 96

Gold recovery (%) 90 88 88 90 91 91 90 89 90 90 86 89 90 91 92 90 89 Antimony feed grade (%) 4.39 4.29 4.60 5.40 4.32 4.77 3.76 4.16 4.58 4.37 3.33 4.38 5.37 4.65 3.72 4.50 4.32

Gold feed grade (g/t) 9.10 8.25 8.22 8.78 8.03 8.81 7.09 7.08 9.24 9.40 6.27 8.85 9.59 9.96 10.32 8.32 8.84

Test period Powerscreen Aug 31st to Oct 11th Finlay Terex crusher in service from Oct 12th



17.2 Plant Upgrade Mandalay had undertaken a study with assistance from Sedgman Metal Engineering to determine the areas of the plant that need to be upgraded to increase the Brunswick Processing Plant throughput from a nominal 7,200 t/mth to a nominal 10,000 t/mth. However, due to the success of a portable crushing unit the recommendations therein were mostly discarded. On the basis of recent plant performance at higher throughput, it has been determined that a plant upgrade is to be investigated regarding:

• improvements and capacity increase in flotation,

• provision of redundant pumps; and

• concentrate handling changes.

The remainder of the plant, including milling, should remain unchanged. These potential upgrades are aimed at increasing reliability, efficiency and maintaining recovery at higher feed grades. The gold room is also to be upgraded to increase gravity gold handling capacity. Figure 17-2 illustrates the proposed upgraded processing plant flowsheet.

17.2.1 Crushing and Screening Circuit To achieve 10,000 t/mth, the first area of focus was the crushing and screening circuit. The most cost effective option was the portable crusher unit, even with higher operational cost the determining factors were the initial outlay and construction downtime incurred from going with the fixed plant option. Just by trialling and eventually purchasing a mobile crushing plant, the 10,000 t/mth target was achieved with no major upgrades to the remainder of the plant. By utilising the existing crushing plant and fine ore bins, the crushed product from the portable crushing plant can be transferred at finer feed size into the ore bins twice a shift to provide consistent feed to the existing milling circuit.

The initial trial in September 2012 of a portable crusher was able to establish the improvements in capacity of the mill, which has been maintained at 10,000 t/mth (ore feed dependant) with the larger portable crushing unit.

17.2.2 Milling Circuit The milling circuit will remain basically unchanged. It has been identified during increased throughput trials in December 2012 that the secondary mill recirculation load is the bottleneck of the current milling circuit. This can only be rectified by installing a larger secondary mill if the current flotation feed size distribution is to be maintained. Redundant pumps are to be installed on Ball Mill 1 and 2 discharge pumps, flotation feed and Knelson feed by the third quarter of 2013. This will eliminate the need to shut down to rebuild pumps on a monthly basis.

17.2.3 Flotation Circuit Due to the increased throughput the current flotation circuit has reduced volumetric capacity, and corresponding retention time. Consequently, investigation is taking place to improve the flotation circuit capacity by installing two additional tank cells on the front end of the flotation circuit. This adjusted circuit will still be fed from the secondary mill closed circuit cyclone overflow. The existing surge/ conditioning tank will feed the two new 10 m3 tank cells in series and then the two site fabricated tank cells in series. Denver cells will continue to be used for the remaining rougher, scavenger and cleaning duties and are to be refurbished later in 2014. The cleaner circuit capacity will also be increased from four Denver square DR15 cells to six Denver square DR15 cells. By installing additional tank cells, the rougher/ scavenger DR100 cells and the walls of the



existing tank cells can be refurbished in sequence without the need to shut down the plant for an extended period.

17.2.4 Concentrate Thickening and Filtration Also being investigated is an improved thickening and filtration circuit. The potential concentrate handling circuit is to be upgraded via the installation of an agitated concentrate storage tank with corresponding density gauges constructed adjacent to the thickener. A froth buster will be installed on the feed to the thickener to de-aerate the feed.

17.2.5 Tailings Circuit The tailings thickener has sufficient capacity to handle the increased tonnage rate. The average tails thickener underflow density has been maintained at around 50% (+/- 10%). This has been achieved since September 2012 processing at increased rates. A redundant tails pump is to be installed on the tails thickener during the fourth quarter of 2013. A new tailings facility is to be constructed in quarter two 2013 with completion in quarter three 2013 for capacity till mid-2014. Options for additional tailings storage are currently being evaluated for mid-2014 onwards.

17.2.6 Reagent Mixing and Storage No upgrade work is anticipated for the reagent mixing and storage area.

17.2.7 Recovery All improvements planned to the existing circuit will be to maintain current levels of recovery.

17.2.8 Services Water The water services at the Brunswick Processing Plant consist of the raw water, process water and excess water disposal. The plant operates in a positive water balance situation with excess water being lost to the atmosphere by evaporation which is accelerated by the use of an evaporator at the Brunswick Pit.

The process water supply consists of concentrate thickener overflow, tailing thickener overflow and TSF decant return water. Process water is stored in and distributed from a dedicated tank. TSF decant water is stored in a pond adjacent to the plant. As the site is in a positive water balance, adequate process water supplies are available.

Air The Brunswick Concentrator requires both low and high pressure air supplies. The addition of two new 10 m3 flotation cells will require an upgrade to the low pressure air supply. Currently three separate low pressure blowers supply the rougher, scavenger and cleaner cells, with existing tank cells running off high pressure air. With the purchase of two new blowers of sufficient capacity (with one as standby) all of the flotation cells, including the existing and new tank cells are to be supplied with low pressure air. The high pressure air supply will be upgraded in late 2013 to increase the capacity and availability of high pressure air to the filter press to improve filtration efficiency.



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 approximately 10,000 t/mth. The current calculated peak load is estimated at 900 kVA and occurs when all the primary crushing is operational (if the mobile crusher unit is not operable and the existing open circuit jaw crushing circuit is used), as shown below in Table 17-3.

Table 17-3: Current power requirements at Brunswick Processing Plant

Description Measurement

Current supply capacity 1,000 kVA

Current power usage (peak) 900 kVA

With the advent of the portable crusher and increased tonnage there has been a transfer of power away from the former primary crushing circuit to the diesel powered portable crusher. The average total Brunswick monthly power usage prior to the increased throughput was 858 kVA and decreased to 788 kVA afterwards.

By decreasing the power demand of the primary crushing circuit and only upgrading the back end of the circuit, there is no appreciable increase in electrical power requirements and it has been deemed that the current supply capacity is sufficient to operate the upgraded Brunswick Processing Plant. If there is an issue with insufficient power supply in the future the evaporator system can be taken off mains power and run off a generator, freeing up an additional 58 kVA if needed.

However, an increase of the electrical supply to 1,400 kVA at 22 kV at Brunswick has been included in the Powercor distribution infrastructure upgrade to ensure sufficient electrical capacity is made available to satisfy any unforeseen electrical loads as well as enable future plant upgrade options to be considered.



Figure 17-2: Brunswick Processing Plant Proposed Upgrade Flow Sheet



18 Project Infrastructure 18.1 Surface Infrastructure

The Costerfield Operations surface facilities are representative of a modern antimony-gold mining operation. The Augusta Mine site location, refer to Figure 18-1, which 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 raise;

• Ventilation intake raise; and

• Augusta water storage dam to manage rainfall run-off and mine dewatering.

Figure 18-1: Augusta Mine Site



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

• Antimony-Gold processing plant and associated facilities;

• Central administration complex;

• Process plant workshop;

• Tailings storage facilities;

• Run of Mine stockpiles;

• Previously mined Brunswick Open Pit; and

• Core farm and core processing facility.

Figure 18-2: Brunswick Site Area



18.2 Augusta Mine 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 around 870 mRL, as at the end of April 2013. The gradient of the decline is 1:7.

The mine has primary ventilation infrastructure installed comprising of two single stage 132 kW axial fans located at the 1020 mRL. The return airway consists of a 3 m diameter raise from surface to the 1020 mRL, with a series of short raises progressively extended from the 1020 mRL to depth, that as of April 2013, extend to the 900 mRL or approximately 280 m below surface. Fresh air is supplied via the decline portal with secondary fans positioned in the decline force ventilating levels.

The mine dewatering system consists of a staged pump station system that pumps mine water (groundwater inflows and mine process water) directly to the surface where it flows through a series of silt ponds before being directed to the evaporation/ storage dams adjacent to the Augusta Surface facilities.

An underground crib room exists at the 1085 mRL and the underground magazine exists at the 1020 mRL.

In addition to the fixed plant, Mandalay owns, operates and maintains all mobile mining equipment including jumbo development drills, production drill, loaders, trucks, and ancillary equipment required to undertake mining operations.

18.3 Tailings Storage Since operations began in the 1970’s, 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 TSF dams were constructed based on a turkeys nest type design with earthen embankments.

All tailings are deposited in the Bombay TSF which last had a lift completed in the first quarter of 2012. This lift increased the storage capacity by an additional 120,000 m3, which at the scheduled LOM production rate will extend the life of the facility until July 2013. A further 2 m lift was completed in August 2013 to further extend the life of the facility until the end of the second quarter of 2014 by adding an additional 130,000 m3 of storage capacity.

A new TSF will need to be brought on line by the end of the second quarter of 2014. A further lift on either the Bombay TSF or Brunswick TSF is being considered as is the use of the old Brunswick Open Pit. Use of the Brunswick Open Pit as a potential future tailings storage facility will require comprehensive assessment before being pursued, but there are precedents for this type of tailings storage approach being used in the region Victoria. If the Brunswick Open Pit is not deemed suitable as a long term tailings storage solution, then it is likely another TSF facility will need to be constructed to satisfy long term tailings disposal requirements, a location for this facility is yet to be identified.

18.4 Power Supply The Costerfield Operations purchases electricity directly from the main national electricity grid and has connections at both the Brunswick Processing Plant and Augusta Mine.



The Costerfield Operations has an existing arrangement for electrical supply of 2,000 kW (2,353 kVA) 22kV supply at the Augusta Mine and a 1,000 kVA 240/415 V 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 approached Powercor in September 2011 regarding three projected load growth scenarios. The chosen load growth scenario has indicated that an upgrade to the Powercor distribution infrastructure will be required to meet a maximum forecasted demand growth to 3,200 kVA at Augusta / Cuffley and 1,400 kVA at Brunswick. This increase in demand is based on extending the Augusta Mine to the 860 mRL, accessing and mining the current known extents of the Cuffley lode Inferred / Indicated Resource as well as expansion the Brunswick Processing Plant.

The upgrade will result in the existing supply point for Augusta remaining unchanged, two new 22 kV supply points being installed at Brunswick and Cuffley. The existing low voltage supply point at Brunswick will be removed.

Powercor has provided a budget estimate of AUD5.7 M for the high voltage upgrade. It has been assumed that the 22 kV upgrade will not be commissioned until December 2014. In the meantime, any shortfalls in power will require the use of diesel generated power until the high voltage upgrade is complete.

The diesel generator installation is planned to consist of two 750 kVA generators that will be connected to an independent 10,000 L fully bunded fuel tank.

An upgrade of the electrical infrastructure that supplies the Costerfield Operations will allow Mandalay to continue to take mains power on a take or pay kWhr charge.

18.5 Power Reticulation Power for the Augusta underground operations is fed from a high voltage 22 kV Powercor feed that feeds:

• The surface switch room via a 1 MVA transformer that steps the voltage down to 415 V; and

• The surface switch yard via a 2 MVA transformer that steps the voltage down to 11 kV.

The 415 V supply is stepped up to 1000 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 1000 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 1020 mRL. From this location, the substation steps the voltage down to 1000 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 935 mRL to service mine working load requirements below this level.

The Augusta Mine also 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.

A Power Factor Correction Unit is planned to be installed and commissioned in April/ May 2013 on the 11 kV supply to increase the efficiency of the underground electrical distribution system.



18.6 Water Supply Water for the Augusta underground and surface operations is sourced from the Augusta Mine Dam at the entrance to the site. Excess mine water that is not consumed by operations is pumped from the Augusta Mine Dam to an evaporation pond and storage dam facility adjacent to the Augusta Mine that has a total storage capacity of 140 ML.

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 179 ML per year from the Augusta underground workings, Brunswick bore and South Costerfield Shaft. An application to increase this to 700 ML per annum has been submitted to Goulburn-Murray Water for approval to satisfy the forecasted combined dewatering output for both Augusta and Cuffley.

Mandalay does not have a permit to discharge water from either of its Costerfield Operations sites.

18.7 Waste Rock Storage The waste rock storage facility located within the Augusta Mine footprint is currently at capacity with regards to both size and height, resulting in all underground waste being trucked off site to the Brunswick TSF for capping purposes.

It is estimated that additional waste storage will be needed to accommodate for a forecasted total waste storage requirement of 233,000 m3, as shown by Figure 18-3. The additional waste storage requirement above the 150,000 m3 Augusta Waster Storage Facility will be satisfied by either capping the Brunswick TSF, utilising the material for future TSF lifts, increasing the current August Waste Storage Facility, or establishing an alternative waste storage site.

Either option may require the rehabilitation bond to be reviewed as construction of either storage facility could impact mine closure plan provisions, depending on whether the waste storage facilities would be permitted as permanent or temporary.

Figure 18-3 also identifies a positive balance at the end of the mine of 173,000 m3, which will be reduced to 13,725 m3 by capping both the Bombay TSF and Brunswick TSF.

Figure 18-3: Augusta Waste Storage Forecast

Campaigns of sizing selected portions of the waste material are also conducted to generate screened rock for 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 Figure 18-3.



18.8 Water Management Groundwater is currently pumped to the Augusta evaporation pond as well as two dams designed for water storage, with a total storage capacity of 160 ML.

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, resulting in a total site storage capacity of approximately 318 ML.

Excess process water is currently disposed of via evaporation the Bombay TSF. Total evaporation/ water disposal capacity is currently estimated at approximately 380 ML per year assuming average Heathcote climatic conditions.

Total inflows are estimated from water balance data to be approximately 515 ML per year plus 87 ML per year of rainfall. Hence, the disposal of mine and process water presents a significant challenge to the mine as a further 220 ML per year of evaporation capacity is required to dispose of the current water disposal requirements.

It has been identified that additional infrastructure needs to be constructed to ensure sufficient capacity is available to not only to dispose of Augusta Mine water but also additional inflows from the Cuffley access development and mining operations.

Mandalay has recently purchased 30 Ha of land within the Costerfield district and has commenced designing an evaporation facility with a net effective evaporation surface of approximately 20 Ha. It is envisaged that this facility will be constructed and commissioned by the first quarter 2013, pending receipt of all required approvals.

Investigations into water disposal infrastructure are on-going with approximately AUD4.5M being included into the Costerfield Operations financial model for these purposes.

To ensure long term water disposal is achievable with potentially greater water generation from development and mining of the Cuffley lode, a new evaporation facility is to be constructed and commissioned in the fourth quarter of 2013. A reverse osmosis treatment facility is to be trialled at the mill in the second and third quarter of 2013, with trucking of water to the Heathcote Open Pit also occurring in the second and third quarter of 2013. Commissioning of the new evaporation facility will provide sufficient water management capacity to eliminate the need for on-going trucking of water to the Heathcote Open Pit.

18.9 Augusta to Brunswick ROM Pad Transport As the Augusta Mine and Brunswick Processing Plant is divided by the Heathcote-Nagambie Road, all ROM product 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.

The average cost of the trucking has been calculated at AUD3.85 per tonne delivered to the Brunswick ROM pad. This cost has been incorporated into the Costerfield Operations financial model.

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. An underground workshop facility has been included at Cuffley to allow basic maintenance and daily servicing to be performed underground.

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 planned Bombay TSF lift.



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 roadtrain that transports the concentrate by road from the Brunswick Processing Plant to the Port of Melbourne. The average payload of each B-double3 roadtrain 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 scheduled once per month and consist of between 600 to 900 wet metric tonnes of concentrate.

Ships depart Melbourne for Shanghai Port in China where the concentrate is transported through Changsha and onto the customer’s smelter at Chenzhou in the province of Hunan.

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 AUD195 per dry metric tonne (dmt).

19.2 Marketing Costerfield is a combined gold and antimony mine with the business being sensitive to the price of both metals. Antimony is not traded on international metal exchanges, with prices being agreed upon between producer and consumer. Pricing is dependent on the quality and form of antimony product sold.

Globally, world antimony production peaked in 2011 at an estimated 203,500 t due to continued growth in Chinese demand (Roskill, 2012), refer to Figure 19-1. The Augusta Mine produced 1,577 t in 2011, accounting for about 0.8% of the world market (Mandalay, 2012).

Figure 19-1: World Antimony production and consumption, 2000 – 2011 (t Antimony) Source: Roskill, 2012.

3 A B double roadtrain 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.



Antimony is primarily used as a flame retardant and in the production of lead acid batteries, these markets together accounted for nearly 80% of antimony consumption worldwide in 2011 (Roskill, 2012).

China is the world’s largest producer of antimony, accounting for approximately 70% of world production in 2011 (Roskill, 2012). Production in China is unlikely to increase in the next few years 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.

This action by the Chinese has resulted in variable production levels which have had a direct impact on pricing. Many Chinese producers have had to increase the level of concentrate imports in order to sustain domestic refined production.

Antimony consumption is forecasted to continue to grow from 2013 onwards, as shown by Figure 19-2. This growth is likely to lead to a supply deficit. The size of this deficit is primarily dependent on investment in new mine production worldwide and China’s success in halting illegal mining and exports of antimony. These market dynamics are likely to result in the antimony price remaining in the range of USD10,000 to USD13,000/ t (Roskill, 2012). Historically, antimony pricing has been steadily increasing as shown by Figure 19-3.

Figure 19-2: World Antimony demand and supply forecast, 2000 – 2016 (t Antimony) Source: Roskill 2011.



Figure 19-3: China trends in domestic FOB prices for Antimony metal and trioxide

Source: Roskill 2011.

19.3 Contracts At present, there is an agreement in place between Mandalay and the Hunan Zhongan Antimony and Tungsten Trading Co Ltd for the sale of the antimony-gold concentrate produced from Augusta. The agreement is valid until the end of December 2014. Mandalay receives payment based on the concentration of the antimony and gold within the concentrate.

Transportation from Melbourne to the smelter in Chenzhou, shipment documentation, freight administration and assay exchange/ returns are conducted by Minalysis Pty Ltd. The marketing of the concentrate is conducted through Penfold Marketing Pty Ltd who markets the product to the customer and provides market information to Mandalay.



20 Environmental Studies, Permitting, and Social or Community Impact

20.1 Environment and Social Aspects

20.1.1 Increased Processing Rates The Costerfield Operation has increased the feed from the current nameplate plant capacity of 7,200 t/mth to approximately 10,000 t/mth from 2013.

The increased throughput is not expected to result in any significant changes in environmental impact at the Brunswick Processing Plant. Dust, noise and traffic management controls associated with production truck movements and crushing operations will be the same as current operations.

Management of tailings is discussed in Section 18.

20.1.2 Mine Ventilation Development of the Cuffley lode will require the installation of a new ventilation shaft to maintain suitable air quality and volumes within the expanded underground mine. The Cuffley ventilation shaft will be located on private land owned by Mandalay Resources and will act as the primary exhaust for the Cuffley development area. The facility will include 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.

A Work Plan Variation approval and Planning Permit Amendment is required to install and operate the new ventilation shaft for the Cuffley Mine. A draft Work Plan Variation was submitted in May 2013. The Planning Permit application can be lodged when the WPV is approved. Approval is required by August 2013.

The new exhaust shaft will give rise to a visible vapour plume during cool weather conditions and therefore community consultation will need to make adjacent residents and the Country Fire Authority aware of the intermittent visual impact of the shaft.

Noise associated with the construction of the ventilation shaft will need to comply with noise limits prescribed for the operation.

Ventilation fans will be located underground and will therefore not result in additional noise emissions on surface.

The site will be fenced to prevent public access.

Some minor removal of native vegetation may be required to establish an easement for power supply to the site from the adjacent power lines. If native vegetation is to be removed then native vegetation offsets will need to be acquired in accordance with Victoria’s Native Vegetation Management – A Framework for Action. Mandalay Resources has purchased a native vegetation offset site in Peels Lane Costerfield which has surplus native vegetation offset credits which can be used fulfil offset obligations.

20.1.3 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. Higher than average rainfall was recorded in Heathcote in 2010 and 2011 with 979 mm and 759 mm respectively, 554 mm, close to the annual average rainfall, was recorded for the 2012 calendar year.

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.

An evaporator is 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 capacity is calculated to be approximately 150 ML/year, which is insufficient to sustainably manage current water disposal requirements. Therefore, a further 550 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. This facility is designed to meet the site’s water disposal needs for the next 1-2 years. Beyond this additional water disposal capacity will be required; the facility will require a Work Plan to be approved by DPI and a Planning Permit to be issued by the City of Greater Bendigo. The land purchase was finalised on 13 December 2012. Whilst this Technical Report has assumed water disposal requirements will be met through the construction of large evaporation facilities, other options to remove water, including the expanded use of evaporators and water treatment with off-site disposal, will be considered and evaluated as an optimisation phase of the project design.

20.1.4 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 DPI 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.

20.1.5 Tailings Disposal All tailings are currently deposited in the Bombay TSF, on which a three metre lift was completed in the first quarter of 2012. A further two metre lift was completed in the second half of 2013. These two expansions will bring the total storage capacity of the Bombay TSF to 361,000 m3 and will enable tailings to be deposited until around mid-2014. A further lift of the Bombay TSF or a new TSF will need to be brought on line from mid-2014. Various locations are being considered for future tailings storage requirements, and may require the purchase of land or approvals to construct on Crown land. Any new TSF will require a variation to the work plan and an amendment to the Planning Permit.

The old Brunswick Open Pit, currently used for storing groundwater, is also being considered as a potential future TSF. This will require comprehensive assessment before being pursued, but there are precedents for this type of TSF approach being used in the Victorian region. If the Brunswick Open Pit is not deemed suitable, then another location will need to be identified for a new facility.

20.2 Impacts

20.2.1 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 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;

• 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.



20.2.2 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 450 ML/year.

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.

Figure 20-1: Groundwater elevation contour map of the areas surrounding the Augusta Mine

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.

Mandalay has applied to the rural water authority, Goulburn-Murray Water, for an increase in dewatering capacity from the current groundwater extraction license limit of 179 ML per annum to 700 ML per annum. The water authority’s assessment will consider whether the increase in groundwater extraction is likely to have a detrimental impact on beneficial users of groundwater or the surface water system. An independent hydrogeologist has completed a review and compiled Mandalay’s hydrogeological information in order to support the application to increase the volume of groundwater extracted.



20.2.3 Noise The approved Environmental Monitoring Plan for the Costerfield Operation includes a noise monitoring programme which comprises routine attended and unattended noise monitoring at five 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.

20.2.4 Blasting and Vibration DPI 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.

20.2.5 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 (EVC’s) 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 EVC’s plus a number of old trees and consequently Mandalay retains a surplus of native vegetation offset for potential future needs.

The proposed expansion of the Costerfield Operation 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 future needs.

20.2.6 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.



Additional visual impacts will need to be considered and mitigation measured identified if tailings cannot be deposited into the old Brunswick Open Pit or a new TSF is required post 2015.

20.2.7 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.

20.2.8 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 DPI 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.

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. Community relations are considered to be very good.

20.2.9 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 TSF’s post closure.

The rehabilitation bond associated with the Company’s Mining Licence MIN4644 was reviewed and increased in mid-2012. The WPV associated with the proposed expansion of the current Costerfield Operation will require the rehabilitation bond to be reviewed and amended. These costs have been included as part of the financial evaluation of the project and are addressed in more detail in Section 21.

20.3 Regulatory Approvals

20.3.1 Work Plan Variation The establishment of new ventilation shaft for Cuffley and the construction of new groundwater evaporation facilities will require a WPV to be approved. The DPI facilitates this approval process and will engage with relevant referral authorities, as required. The DPI 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 the regulatory authorities, and any conditions or proposed amendments requested to the WPV are generally negotiated to the satisfaction of both parties.

20.3.2 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, TSF and vent shaft.

Goulburn-Murray Water

Licence to take and use water to be increased to 700 ML/ year to allow for expected increase in groundwater dewatering rates.

Groundwater bore construction licences for new bores for monitoring groundwater in the vicinity of the Cuffley lode.

DSE

Compliance with Native Vegetation Management Framework for removal of native vegetation associated with the vent shaft power supply, evaporation facility and new TSF.

Consent to install a pipeline through Crown land from the Augusta mine to the new evaporation facility.

Parks Victoria Consent has been granted to extend the mine under a parcel of Restricted Crown Land on the eastern side of the Heathcote-Nagambie Road.

EPA EPA consent to pump groundwater off the mining licence to a new evaporation facility.

20.4 References The environmental and socio-economic content presented in Section 20 is based on an independent report provided in September 2012, SKM, Technical Assessment of draft environmental, permitting and social impact content for Costerfield Operation PEA, 3 September 2012.



21 Capital and Operating Costs The costs 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 2013 Australian dollars and are to a level of accuracy of ±20 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 estimates using the AUD:USD exchange rate of 0.90.

The capital and operating cost estimates have been consolidated into the file Costerfield PEA Financial Model_20130903.xlsx.

This preliminary economic assessment is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realised.

21.1 Capital Costs Table 21-1 summarises the total capital required for the Costerfield Operation. A detailed breakdown of the individual capital items included in the Financial Model was sourced from the 2013/ 14 budget document Version 4.

Table 21-1: Summary of Capital Costs

Item LoM Total (AUD M)

CY 13 (AUD M)

CY 14 (AUD M)

CY 15 (AUD M)

CY 16 (AUD M)

CY 17 (AUD M)

Sub-total Plant 3.7 0.1 3.15 0.6 0.0 0.0

Sub-total Admin 0.4 0.4 0.0 0.0 0.0 0.0

Sub-total Enviro 4.5 2.5 2.0 0.0 0.0 0.0

Sub-total OH&S 0.0 0.0 0.0 0.0 0.0 0.0

Sub-total Ops Geology 1.3 1.0 0.3 0.0 0.03 0.0

Sub-total Exploration 0.0 0.0 0.0 0.0 0.0 0.0

Sub-total Mining 6.2 2.1 2.2 1.7 0.2 0.0

Total - Plant & Equipment 16.1 6.0 7.6 2.3 0.2 0.0

Capital Development 15.2 3.3 4.5 4.7 2.7 0.0

Total 31.3 9.3 12.1 7.0 2.9 0.0

21.1.1 Capital Lateral Development Decline development quantities have been based on 3D mine designs prepared for the Augusta Mine and proposed Cuffley mine design. 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, miscellaneous storage and CRF mixing bays.



The estimated unit cost for lateral development has been developed from Augusta historical costs for labour, equipment, consumables, services, as well as achieved productivities. An allowance for the haulage of waste rock to surface has also been included.

The lateral development for Augusta and Cuffley will continue to be undertaken as an ‘owner operator’ for the remaining LoM. Approximately 3,700 m of waste development is required at an estimated cost of AUD 4,100/ m.

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 developed from budget estimates compiled by qualified mining contractors. The unit costs include mobilisation and demobilisation of the contractor and the contractor’s indirect fees and profit.

Point Break Mining, a specialist vertical development contractor, is currently contracted to undertake the vertical air leg rising and shrinking of the return airway system at Augusta. It has been assumed that these short return airway extensions between levels will continue to be completed by a contractor.

Additional planned vertical ventilation development comprises an upgrade and expansion of the current Augusta primary ventilation system and establishment of the Cuffley ventilation system. This work will require the services of a specialist shaft contractor to construct the vertical development. A total of 1,207 m has been allowed for in the schedule and incorporated into the operating costs.

21.1.3 Infrastructure The main infrastructure cost item is the construction of additional evaporation facilities. The additional evaporation facilities are required to assist with the projected increase in the underground dewatering volumes in line with the deepening of Augusta. All associated costs have been estimated from previous surface construction activities and have taken into account the need to purchase land to accommodate these water disposal facilities.

All remaining infrastructure costs have been based on Mandalay’s recent experience with similar installations.

21.1.4 Mobile Plant Mobile plant purchases include LHD units, underground haul trucks, production drill and auxiliary equipment required to meet production targets.

Mobile plant cost estimates have been based on recent quotations from appropriate mobile plant suppliers.

21.1.5 Processing Plant Mandalay has identified and estimated the costs associated with the Brunswick Processing Plant expansion and other mill site related initiatives including:

• Bombay TSF lift;

• New TSF construction;

• Flotation shed facility upgrades;

• Refurbishment of existing plant and key components;

• Purchase of critical spares;



• Compressed air upgrade; 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.

The exploration infill diamond drilling is related to further defining the Augusta and Cuffley Mineral Resource used in this Technical Report, with the primary focus to convert Inferred Resources to the Indicated category. The brownfields drilling is related to testing possible extensions of the Cuffley lode that have been identified but are yet to be drill tested. This drilling has the potential to expand the Cuffley Mineral Resource outside that specified in this Technical Report.

21.1.7 Closure Closure costs have been included for the rehabilitation of facilities constructed specifically for the deepening of Augusta. The low net cost is due to an expected refund of current rehabilitation bonds held by the regulatory authorities.

21.2 Operating Costs The operating cost estimates applied in this Technical Report are summarised in Table 21-2, Table 21-3 and described further in the following sections.

Table 21-2: Operating Cost Inputs


Mining

Jumbo Development AUD/m 4,100

Air Leg Development AUD/m 1,742

Stoping AUD/t 113

Resource Definition Drilling AUD M 1.451

Supervision AUD/mth 505,496

Services AUD/mth 680,999

Processing Cost AUD/t milled 60

Site Services AUD/day 6,059

General and Administration AUD/day 14,377

Selling Cost AUD/t con 195

Note: The Resource Definition Drilling occurs during Month 1 to 9 and is for conversion of the Inferred mineral resource to Indicated.



Table 21-3: Costerfield Operation - Operating Cost Summary

Description Operating Cost

AUD M AUD/t USD M USD / t

Mining 122 240 110 216 Processing 31 60 27 54 Site Services 9 17 8 15 General and Administration 21 40 19 36 Total 182 358 1647 322

21.2.1 Lateral Development The estimated unit cost for lateral development has been developed from Augusta 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 (capital and operating) for Augusta and Cuffley will continue to be undertaken as an ‘owner operator’ for the remaining LoM. It has been estimated from 3D mine designs that approximately 17,508 m of lateral operating development 1,669 m of waste and 15,839 m of production drive development will be required for the combined Augusta and Cuffley operation.

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 Production Stoping The direct costs related to production stoping has been developed from Augusta historical costs for labour, consumable material and equipment operating and maintenance, as well as achieved productivities associated with:

• Installation of secondary ground support;

• Drilling, loading, and blasting longholes 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.

21.2.3 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.

21.2.4 Processing Brunswick processing costs have been estimated from Mandalay actual 2013 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 financial model was developed by SRK.

All costs are constant in 2013 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. The financial indicators examined included pre-tax cash flow, net present value (NPV), internal rate of return (IRR) and payback period.

This preliminary economic assessment is preliminary in nature and includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves. There is no certainty that the preliminary economic assessment will be realised.

22.2 Principal Assumptions The principal assumptions in the mining schedule described in Sections 15 and 16 of this Technical Report and the economic parameters in Section 21.

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 509,000

Gold grade (g/t) 10.5

Antimony grade (%) 3.8

Project Life months 47

Average Production Rate t/mth 10,800

Maximum Mining Rate t/mth 13,000

Metallurgical Recovery Gold (%) 89

Antimony (%) 96

Gravity Gold % 34

Concentrate Grade Gold (g/t) 71

Antimony (%) 53

Concentrate Selling Expenses AUD/dmt 195

Payable

Gold in Con (%) 78.5

Gravity Gold (%) 98

Antimony (%) 63

Exchange Rate AUD:USD 0.90

Commodity Prices Gold USD/oz 1,300

Antimony USD/t 9,500

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 Sales The economic model assumes that concentrate shipments are made at the end of each time period, monthly. The payables of the shipments and 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 all gold recovered reports to the concentrate.

22.2.3 Exchange Rate The economic model has assumed an exchange rate of AUD:USD 0.90 for the entire project life.

SRK notes that over the period that this Technical Report was prepared the exchange rate has fallen and is trading at a new level, refer Figure 22-1. As will be discussed in the sensitivity section the project financials are most sensitive to a change in exchange rate.

Figure 22-1: AUD:USD Exchange Rate Chart Source: www.xrates.com, 19 July 2013

22.2.4 Taxes The Australian Government taxes on Mandalay’s 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.

The impact of the recently introduced carbon tax on the base cost calculations has not been factored into the economic model in this Technical Report.

The Australian Government mining resource rent tax doesn’t apply to gold or antimony.

For the purposes of this Technical Report, a pre-company tax cashflow projection has been generated.


http://www.xrates.com/


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 Operation. 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 economic model. 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 economic model.

22.3 Economic Summary A summary of the economic factors associated with the project are presented in Table 22-2. Figure 22-2 shows the monthly and cumulative cashflow.

Table 22-2: Project Economics

Description Units Quantity Units Quantity

Tonnes Milled Tonnes 509,000 Tonnes 509,000

Recovered Gold Ounces 152,000 Ounces 152,000

Recovered Antimony Tonnes 18,000 Tonnes 18,000

Operating cost AUD M 182 USD M 164

Operating Cost per Payable ounce AUD / Oz Eq1 851 USD/oz eq 766

Capital cost AUD M 31 USD M 28

Total Cost (Operating + Capital) AUD M 213 USD M 192

Total Cost per Payable Ounce AUD / Oz Eq 998 USD/oz eq 898

Payable Gold Ounces 129,000 Ounces 129,000

Payable Antimony Tonnes 12,000 Tonnes 12,000

Payable (Saleable) Metal – Au Eq Oz Eq 214,000 Tonnes 214,000

Net Revenue (less selling expenses) AUD M 299 USD M 269

After Tax Profit AUD M 84 USD M 76

After Tax NPV5 AUD M 74 USD M 67

IRR % 3309 % 3309

Max Negative Cashflow AUD M -2 USD M -2

Max Negative Cashflow Mth Mar 2014 Mth Mar 2014

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 22-2: Undiscounted Cashflow (AUD)

22.4 Sensitivities 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 22-3 and Figure 22-3.

The sensitivity analysis demonstrates that the cash flow will be most sensitive in order to exchange rate, operating costs, gold price, gold grade then gold recovery.

A change in the AUD:USD exchange rate of 0.05 results in a revenue change of AUD16M and an NPV change of AUD14M.

Noting that the Capital Cost estimates used foe 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.

Table 22-3: Sensitivity at 5% NPV

-20% (AUD M)

-10% (AUD M)

Base (AUD M)

10% (AUD M)

20% (AUD M)

Gold Price 42 59 74 87 99

Antimony Price 54 65 74 82 90

Mill Feed Tonnes 34 55 74 90 105

Exchange Rate 21 48 74 94 114

Metallurgy Gold Recovery 42 59 74 87 99

Metallurgy Antimony Recovery 56 66 74 82 89

Mill Feed Gold Grade 42 59 74 87 99

Mill Feed Antimony Grade 56 66 74 82 89

Capital Cost 77 75 74 73 72

Operating Cost 100 87 74 58 40

Rounded to nearest million

-30

-20

-10

0

10

20

30

40

50

60

70

80

90

-3

-2

-1

0

1

2

3

4

5

6

7

8

9

Aug-

13

Oct

-13

Dec-

13

Feb-

14

Apr-

14

Jun-

14

Aug-

14

Oct

-14

Dec-

14

Feb-

15

Apr-

15

Jun-

15

Aug-

15

Oct

-15

Dec-

15

Feb-

16

Apr-

16

Jun-

16

Aug-

16

Oct

-16

Dec-

16

Feb-

17

Apr-

17

Jun-

17

Cumulative Cashflow(AUDM)

Monthly Cashflow(AUDM)

Monthly Cumulative



Figure 22-3: Sensitivity Analysis

22.5 After Tax Cashflow The after tax estimate is based on a company tax rate of 30% company tax rate, straight line depreciation, an opening book value of AUD43M and AUD42M of tax losses carried forward.

The depreciation schedule is presented in Table 22-4 was carried forward into the after tax cashflow calculation presented in AUD Table 22-5 and USD in Table 22-6.

Table 22-4: Depreciation Schedule

CY 13

(AUD M)

CY 14

(AUD M)

CY 15

(AUD M)

CY 16

(AUD M)

CY 17

(AUD M)

Total

(AUD M)

30 June Carried forward 43.116 0 0 0 0 43.116

PEA Capital Requirements 9.334 12.071 6.976 2.914 0 31.295

Total Capital 52.451 12.071 6.976 2.914 0 74.412

Depreciation Schedule

Calender Year 2013 10.490 0 0 0 0 10.490

Calender Year 2014 10.490 3.018 0 0 0 13.508

Calender Year 2015 10.490 3.018 2.325 0 0 15.833

Calender Year 2016 10.490 3.018 2.325 1.457 0 17.290

Calender Year 2017 10.490 3.018 2.325 1.457 0 17.290



Table 22-5: After Tax Profit & Loss and Valuation (AUD)

Units Aug-Dec 13 CY 14 CY 15 CY 16 CY 17

Revenue (Nett ie less Royalty) AUD M 27.268 88.616 86.995 61.112 34.617

Operating cost AUD M 18.753 49.535 48.223 43.735 21.604

EBITDA AUD M 8.516 39.081 38.771 17.377 13.013

Depreciation AUD M 10.490 13.508 15.833 17.290 17.290

EBIT AUD M (1.975) 25.573 22.938 0.087 (4.277)

Tax Losses carried forward AUD M (42.029) (44.004) (18.430) 0.000 0.000

Tax Losses added AUD M (1.975) 0.000 0.000 0.000 (4.277)

Tax Losses used AUD M 0.000 (25.573) (18.430) 0.000 0.000

Taxable Income / (Loss) AUD M (1.975) 0.000 4.508 0.087 (4.277)

Corporate Tax Rate % 30 30 30 30 30

Tax Owing- Real AUD M 0.000 0.000 1.352 0.026 0.000

Nett Profit After Tax (NPAT) AUD M (1.975) 0.000 3.155 0.061 (4.277)

Valuation

Revenue (Nett ie less royalty) AUD M 27.268 88.616 86.995 61.112 34.617

Operating Cost AUD M 18.753 49.535 48.223 43.735 21.604

Capital cost AUD M 9.334 12.071 6.976 2.914 0.000

Tax Owing- Real AUD M 0.000 0.000 1.352 0.026 0.000

After Tax Cashflow AUD M (0.819) 27.010 30.443 14.437 13.013

Valuation Cashflow AUD M (0.819) 27.010 30.443 14.437 13.013

Discount Rate % 5

NPV AUD M 74.2

IRR % 3,309



Table 22-6: After tax Profit & Loss and Valuation (USD)

Units Aug-Dec 13 CY 14 CY 15 CY 16 CY 17

Revenue (Nett ie less Royalty) USD M 24.542 79.754 78.295 55.001 31.155

Operating cost USD M 16.878 44.581 43.401 39.362 19.443

EBITDA USD M 7.664 35.173 34.894 15.639 11.712

Depreciation USD M 9.441 12.157 14.250 15.561 15.561

EBIT USD M (1.777) 23.016 20.644 0.078 (3.849)

Tax Losses carried forward USD M (37.826) (39.603) (16.587) 0.000 0.000

Tax Losses added USD M (1.777) 0.000 0.000 0.000 (3.849)

Tax Losses used USD M 0.000 (23.016) (16.587) 0.000 0.000

Taxable Income / (Loss) USD M (1.777) 0.000 4.057 0.078 (3.849)

Corporate Tax Rate % 30 30 30 30 30

Tax Owing- Real USD M 0.000 0.000 1.217 0.023 0.000

Nett Profit After Tax (NPAT) USD M (1.777) 0.000 2.840 0.055 (3.849)

Valuation

Revenue (Nett ie less royalty) USD M 24.542 79.754 78.295 55.001 31.155

Operating Cost USD M 16.878 44.581 43.401 39.362 19.443

Capital cost USD M 8.401 10.864 6.278 2.622 0.000

Tax Owing- Real USD M 0.000 0.000 1.217 0.023 0.000

After Tax Cashflow USD M (0.737) 24.309 27.399 12.993 11.712

Valuation Cashflow USD M (0.737) 24.309 27.399 12.993 11.712

Discount Rate % 5

NPV USD M 66.8

IRR % 3309

22.6 Cash Flow Forecast The proposed production schedule presented in Table 22-7 was used to derive the estimated cash flow forecast is provided in AUD in Table 22-8 and USD in Table 22-9.



Table 22-7: Proposed Production Schedule

Total 2013 Q3 2013 Q4 2014 Q1 2014 Q2 2014 Q3 2014 Q4 2015 Q1 2015 Q2 2015 Q3 2015 Q4 2016 Q1 2016 Q2 2016 Q3 2016 Q4 2017 Q1 2017 Q2

PHYSICALSDevelopment

Capital Development m 3,713 300 510 250 283 283 284 289 299 266 298 234 179 144 95 0 0Operating Development m 15,976 941 1,415 1,324 1,326 1,057 1,063 1,074 1,213 1,289 1,382 1,544 577 733 558 294 187Vertical Development m 770 0 223 0 110 48 48 48 65 31 96 6 48 2 46 0 0

Underground Ore

Total Tonnes t 508,057 21,199 33,558 36,759 39,287 34,932 32,989 34,806 28,995 34,624 34,644 31,643 31,529 29,149 23,212 27,507 33,223g/t Au 10.46 7.35 7.37 7.17 10.63 13.35 13.70 14.82 13.14 12.44 9.50 13.33 8.82 8.85 6.42 10.96 7.42

Ounces 170,863 5,011 7,956 8,479 13,422 14,995 14,531 16,581 12,247 13,851 10,579 13,563 8,945 8,293 4,790 9,691 7,928% Sb 3.75 3.96 4.15 3.55 3.36 4.32 4.21 4.28 3.38 3.70 3.33 3.90 3.10 2.80 3.26 3.87 4.67

Tonnes 19,060 840 1,392 1,305 1,318 1,511 1,388 1,489 980 1,282 1,153 1,235 976 816 758 1,064 1,553

METALLURGY Mill Feed t 508,057 21,199 33,558 36,759 39,287 34,932 32,989 34,806 28,995 34,624 34,644 31,643 31,529 29,149 23,212 27,507 33,223

Feed Grade g/t Au 10.46 7.35 7.37 7.17 10.63 13.35 13.70 14.82 13.14 12.44 9.50 13.33 8.82 8.85 6.42 10.96 7.42% Sb 3.75 3.96 4.15 3.55 3.36 4.32 4.21 4.28 3.38 3.70 3.33 3.90 3.10 2.80 3.26 3.87 4.67

Metallurgical Recovery % Au 89% 89% 89% 89% 89% 89% 89% 89% 89% 89% 89% 89% 89% 89% 89% 89% 89%% Sb 96% 96% 96% 96% 96% 96% 96% 96% 96% 96% 96% 96% 96% 96% 96% 96% 96%

Payable Gold ounces 129,337 3,793 6,022 6,419 10,160 11,351 10,999 12,551 9,271 10,485 8,008 10,266 6,771 6,277 3,626 7,336 6,002Payable Antimony tonnes 11,528 508 842 789 797 914 840 901 593 776 697 747 591 493 458 644 939

eq ounces 84,241 3,712 6,152 5,767 5,826 6,677 6,136 6,581 4,330 5,667 5,097 5,459 4,315 3,605 3,349 4,704 6,864Payable Gold Equivalent ounces 213,578 7,505 12,175 12,185 15,986 18,027 17,135 19,133 13,601 16,152 13,104 15,726 11,087 9,882 6,974 12,041 12,866

REVENUEPayable % Au 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785 0.785

% Sb 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63 0.63Price $/oz Au 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300 1,300

$/t Sb 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500 9,500Exchange Rate AUD:USD 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90

2013 2014 2015 2016 2017



Table 22-8: Estimated Pre-Tax Cash Flow Summary in AUD


CAPITAL COSTS Sub-total Plant AUD M 3.7 0.0 0.0 1.4 1.7 0.0 0.0 0.6 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Admin AUD M 0.4 0.2 0.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Enviro AUD M 4.5 1.9 0.6 1.5 0.5 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total OH&S AUD M 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Ops Geology AUD M 1.3 0.4 0.6 0.3 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Exploration AUD M 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Mining AUD M 6.2 0.2 1.9 1.5 0.3 0.4 0.1 0.7 0.2 0.7 0.1 0.0 0.1 0.1 0.0 0.0 0.0 Total-Plant and Equipment AUD M 16.1 2.7 3.3 4.6 2.5 0.4 0.1 1.3 0.2 0.7 0.1 0.0 0.1 0.1 0.0 0.0 0.0

Mine Capital Development AUD M 15.2 1.2 2.1 1.0 1.2 1.2 1.2 1.2 1.2 1.1 1.2 1.0 0.7 0.6 0.4 0.0 0.0 Total Capital Cost AUD M 31.3 3.9 5.4 5.7 3.6 1.5 1.2 2.5 1.4 1.8 1.3 1.0 0.9 0.7 0.4 0.0 0.0

OPERATING COSTSMining AUD M 122.2 4.8 7.5 8.6 8.4 8.5 7.9 8.2 8.1 8.3 8.1 8.1 7.1 7.4 6.7 7.0 7.3Processing AUD M 30.5 1.3 2.0 2.2 2.4 2.1 2.0 2.1 1.7 2.1 2.1 1.9 1.9 1.7 1.4 1.7 2.0Site Services AUD M 8.7 0.4 0.6 0.5 0.6 0.6 0.6 0.5 0.6 0.6 0.6 0.5 0.6 0.6 0.6 0.5 0.6G&A AUD M 20.5 0.9 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3 1.3Total Operating cost AUD M 181.8 7.4 11.4 12.7 12.6 12.5 11.8 12.2 11.7 12.3 12.1 11.9 10.9 11.0 10.0 10.5 11.2Total Operating cost AUD /payable oz Au Eq 851 980 936 1,041 787 691 689 636 857 761 923 755 980 1,116 1,429 868 867

CAPITAL COST + OPERATING COST AUD M 213.1 11.3 16.8 18.4 16.2 14.0 13.0 14.7 13.1 14.1 13.4 12.9 11.7 11.7 10.4 10.5 11.2AUD /oz Au Eq 998 1,505 1,379 1,507 1,015 775 761 767 961 870 1,023 818 1,057 1,185 1,485 868 867

Gross Revenue AUD M 308.5 10.8 17.6 17.6 23.1 26.0 24.8 27.6 19.6 23.3 18.9 22.7 16.0 14.3 10.1 17.4 18.6Concentrate Charges AUD M 6.7 0.3 0.5 0.5 0.5 0.5 0.5 0.5 0.3 0.5 0.4 0.4 0.3 0.3 0.3 0.4 0.5Royalty 2.75% of Antimony Revenue AUD M 3.2 0.1 0.2 0.2 0.2 0.3 0.2 0.2 0.2 0.2 0.2 0.2 0.2 0.1 0.1 0.2 0.3Net Revenue AUD M 298.6 10.4 16.9 16.9 22.4 25.3 24.0 26.9 19.1 22.7 18.3 22.1 15.5 13.9 9.7 16.8 17.8

PRE-TAX CASHFLOWAnnual AUD M 85.5 -0.9 0.1 -1.4 6.2 11.3 11.0 12.2 6.1 8.6 4.9 9.2 3.8 2.1 -0.7 6.4 6.6Cumulative AUD M -0.9 -0.8 -2.3 3.9 15.2 26.2 38.4 44.5 53.1 58.0 67.2 71.0 73.1 72.4 78.8 85.5

20172013 2014 2015 2016



Table 22-9: Estimated Pre-Tax Cash Flow Summary in USD


CAPITAL COSTS Sub-total Plant USD M 3.3 0.0 0.0 1.2 1.5 0.0 0.0 0.5 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Admin USD M 0.4 0.1 0.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Enviro USD M 4.1 1.7 0.5 1.4 0.5 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total OH&S USD M 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Ops Geology USD M 1.2 0.4 0.5 0.2 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Exploration USD M 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 Sub-total Mining USD M 5.5 0.2 1.7 1.4 0.3 0.3 0.1 0.6 0.2 0.6 0.1 0.0 0.1 0.1 0.0 0.0 0.0 Total-Plant and Equipment USD M 14.5 2.4 3.0 4.2 2.2 0.3 0.1 1.2 0.2 0.6 0.1 0.0 0.1 0.1 0.0 0.0 0.0 Mine Capital Development USD M 13.7 1.1 1.9 0.9 1.0 1.0 1.0 1.1 1.1 1.0 1.1 0.9 0.7 0.5 0.4 0.0 0.0 Total Capital Cost USD M 28.2 3.5 4.9 5.1 3.3 1.4 1.1 2.2 1.3 1.6 1.2 0.9 0.8 0.6 0.4 0.0 0.0 OPERATING COSTS Mining USD M 109.9 4.4 6.8 7.8 7.5 7.6 7.2 7.4 7.3 7.5 7.3 7.3 6.4 6.7 6.0 6.3 6.6 Processing USD M 27.4 1.1 1.8 2.0 2.1 1.9 1.8 1.9 1.6 1.9 1.9 1.7 1.7 1.6 1.3 1.5 1.8 Site Services USD M 7.8 0.3 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 0.5 G&A USD M 18.5 0.8 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 1.2 Total Operating cost USD M 163.7 6.6 10.3 11.4 11.3 11.2 10.6 11.0 10.5 11.1 10.9 10.7 9.8 9.9 9.0 9.4 10.0

USD / payable Oz Eq 766 882 843 937 708 622 620 573 771 685 831 679 882 1,005 1,286 781 780 CAPITAL COST + OPERATING COST USD M 191.8 10.2 15.1 16.5 14.6 12.6 11.7 13.2 11.8 12.7 12.1 11.6 10.5 10.5 9.3 9.4 10.0

USD / payable Oz Eq 898 1,354 1,241 1,356 914 698 685 690 865 783 921 736 951 1,067 1,337 781 780

Gross Revenue USD M 277.7 9.8 15.8 15.8 20.8 23.4 22.3 24.9 17.7 21.0 17.0 20.4 14.4 12.8 9.1 15.7 16.7Concentrate Charges USD M 6.1 0.3 0.4 0.4 0.4 0.5 0.4 0.5 0.3 0.4 0.4 0.4 0.3 0.3 0.2 0.3 0.5Royalty USD M 2.8 0.1 0.2 0.2 0.2 0.2 0.2 0.2 0.1 0.2 0.2 0.2 0.1 0.1 0.1 0.2 0.2Net Revenue USD M 268.7 9.4 15.2 15.2 20.2 22.7 21.6 24.2 17.2 20.4 16.5 19.9 14.0 12.5 8.7 15.2 16.0

PRE-TAX CASHFLOW USD M 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0Annual USD M 76.9 -0.8 0.1 -1.3 5.6 10.2 9.9 11.0 5.5 7.7 4.4 8.3 3.4 1.9 -0.6 5.7 6.0Cumulative USD M 615.0 -0.8 -0.7 -2.0 3.5 13.7 23.6 34.5 40.0 47.8 52.2 60.5 63.9 65.8 65.2 71.0 76.9

20172013 2014 2015 2016



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 (through AGD). In the immediate area of the Augusta Mine there are no advanced projects, nor are there any other Augusta-style antimony-gold operations in production in the Costerfield district. Exploration on adjacent prospects (EL3316, EL5310 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

The ownership and status of each of the surrounding exploration leases is shown in Table 23-1.

Table 23-1: Augusta Mine Adjacent Properties (DPL, 2012)

Title Owner Status First Granted Expiry

EL336 Nagambie Mining Ltd Granted 29/12/1992 29/12/2012

EL5406 Victoria Mining Exploration Pty Ltd Application N/A N/A

EL5310 Mr Glenn Connor Granted 21/10/2009 20/10/2014



The Costerfield Operation is situated between 35 to 51 km from other significant central Victorian mining operations. Table 23-2 states the distance from the Augusta Mine site to mines within 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

AMC makes the following observations regarding sample collection:

• Historically, core recovery was an issue for all methods of drilling at 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 practice given the location of drill collars and the expected deviation encountered during drilling. As such, potential for significant impact on results is minimised;

• The nature of geological information collected during the logging process is considered to be in line with industry best practice and consistent with other systems employed at deposits that are similar in nature, that is, discrete, structurally- controlled, narrow-vein deposits; and

• Drilling pattern and density are also considered to be appropriate because 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.

AMC makes the following observations regarding sample preparation and assaying:

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

• 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 50% of blank samples exceeded the acceptable threshold is of concern and is being investigated. However, AMC does not believe that the high blank values are material for the Mineral Resource estimate because the highest blank value obtained (0.81 g/t Au) is still <3% of the average lode grade of 32.7 g/t Au;

• The poor reproducibility observed in the gold assays for duplicate pulps assayed by Onsite and ALS suggests that Onsite may be undercalling higher-grade assays. It is worth noting that this is this is also observed in the gold standard reference material results, but not to the same extent; and

• Security of samples being transported to the laboratory is adequate. However, the storage of returned pulps needs to be reviewed and rectified.

AMC makes the following observations regarding the Mineral Resource estimate:

• The E, W, NW and P lodes show good agreement between the composite grade and the block model estimated grade due to the dense sample spacing. These structures were predominantly informed by face sample data;

• The NE and Cuffley lodes with fewer or no face samples show poorer agreement between the composite grade and the block model estimated grade. This is reflected in the Mineral Resource classification;



• The accuracy of the Cuffley lode estimate will benefit substantially from another phase of drilling that focuses on defining the orientation of best grade continuity. This additional information could allow meaningful variography to obtained for the Cuffley lode and allow ordinary kriging to be applied; and

• Reconciliation results show reasonable agreement between the Resource block model and the processing plant during 2012, considering the unquantified errors that influence this result. Such errors include stockpiling, ore-waste misallocation, and unplanned dilution. Over the period, the ounces of gold predicted by the model were 14% lower than that produced by the plant. The tonnes of antimony predicted by the model were 3% lower than that produced by the plant.

In AMC’s opinion, the geological data used to inform the Augusta and Cuffley Mineral Resource 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. The reconciliation results also give confidence to the sample collection procedures, the quality of the assays, and the resource estimation methodology.



26 Comments and Recommendations 26.1 Geology

The Costerfield Property is an advanced property and Mandalay has a history of successful exploration and mining on the Property. AMC has observed that Mandalay geologists are appropriately trained and invest substantial resources into the collection and analysis of structural information from drill core and underground mapping. This structural analysis contributes to the geological understanding of the Property and is used to guide exploration and mining. AMC is aware that Mandalay actively explores the property and devotes considerable resources and energy into maintaining the resource base. Therefore, AMC is not in a position to recommend a work program that would add value to the existing work plans.

26.2 Mining Based on the findings of this Technical Report the following recommendations are provided:

• Review the mineral resource to assess potential to identify additional tonnes closer to Augusta;

• Undertaking more detailed scheduling work to minimise the negative cashflow at the start of the mining schedule;

• Undertake additional scheduling to smooth the development and production profile; and

• Continue to refine and review mining methods and capital and operating cost estimates.

Compiled by

Mr Peter Fairfield

Principal Consultant (Project Evaluations)

Peer Reviewed by

Ms Anne-Marie Ebbels

Principal Consultant (Mining)



27 References Biosis Research, 2005. Flora, Fauna and habitat hectare values of native vegetation at Peels Lane,

Costerfield South, Victoria.

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.

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Documents

Mandalay Resources Corporation Costerfield Operation ......Brett Muller, SRK Associate, Principal Consulting Engineer (Metallurgy), BEng (Minerals Engineering and Extractive Metallurgy),BCom