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Updated Mineral Reserve Statement

for

Azerbaijan International Mining Company – Gedabek Mineral Deposit

Prepared by CAE Mining

PROPRIETARY NOTICE: The information contained herein is confidential and/or proprietary to CAE Mining Canada Inc., and shall not be reproduced or disclosed in whole or in part, or used for any purpose whatsoever unless authorized in writing by CAE Mining Canada Inc.

27 November 2014

mailto:[email protected]

http://www.cae.com/mining

Gedabek Reserve Statement

1-2 © 2013 CAE Mining

Use or disclosure of this document is subject to the restrictions on the title page of this document.

1 Executive Summary

CAE Mining was requested by Azerbaijan International Mining Company (AIMC) Limited, to estimate the mineral reserves for the existing open pit operation of the Gedabek mineral deposit located in the Republic of Azerbaijan. The estimation is an update of a previous estimate carried out by CAE Mining in May 2012, and was completed in accordance with the Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves (The Joint Ore Reserves Committee, 2012).

The mineral reserve estimate is based on the latest Resource Estimate (May 2014) which, in turn, takes into consideration information from previous (2006, 2007, 2009, 2010 and 2011) and recent (2011, 2012 and 2013) exploration drilling campaigns. The recent campaign was extensive and is based on an additional 99 holes out of 394 total drillholes, giving a total of 61,714.6 m and 27,623 samples.

The main objectives of the drilling campaign were to increase the level of geological knowledge and to increase the confidence in the quantity (tonnage) and quality (gold, copper and silver) estimates.

The Resource Estimate was classified into Measured, Indicated and Inferred mineral resources of the oxide, transition, sulphide and combined mineralisation based on a cut-off grade greater than or equal to 0.3 g/t of Au. This gave a total Resource for Measured and Indicated (M&I) of 61.8 Mt.

It should also be noted that the Resource Statement includes an area to the west of the open pit that is know as Ghadir. This Inferred Resource will be mined by underground methods and has been excluded when estimating the Reserve.

The Reserve estimate assumes a direct correlation between Proven and Probable, and Measured and Indicated, and excludes Inferred Resources. It is also assumed that all material types (oxide, sulphide and transition) can be processed through a combination of process routes that include Heap Leach (HL), Agitation Leach Plant (ALP) and Flotation.

The other key difference between the Resource and Reserve estimate is that the Reserve is not based on a fixed cut-off grade as the material is directed to the most appropriate processing method according to the Net Smelter Return (NSR) for the contained metals gold, silver and copper. For instance, for Flotation, the breakeven cut-off is primarily dependent on the copper grade and silver and gold are bi-products in the concentrate.

It is also assumed that with heap leaching or agitation leaching, the copper and silver is recovered by the Sulphidization - Acidification-Recycle -Thickening (SART) process. In the past this has been limited in capacity but for this exercise this constraint has been lifted, and consequently capital will be required.


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The assumed parameters for the various process routes are shown in Table 1.1.

Table 1.1 Process recoveries

Based on the optimised pit limit, a pit design was prepared using an overall pit slope angle of 37.3° and slope parameters recommended by the CQA Geotechnical Engineer. The resulting Reserve estimate is shown in Table 1.2.

Table 1.2 Ore reserve estimate as at 1 September 2014

Note: The reporting of in situ grade and contained metal includes the modifying factors for tonnage and grade that are based on historical reconciliation factors. Recovered metal also includes the relevant recovery factors for the selected process route.

In addition to the ore Reserve of 20.5 Mt, there is also oxide/transition and sulphide material within the selected pit limit that is classified as Inferred in the geological model. This represents an additional Resource that should be investigated in the future with further drilling.

It should also be noted that the pit limit design was based on the optimised pit shell with an 86% Price Factor (Pit 37). This was an obvious point at which to limit the mine expansion due to the incremental waste between Pit 37 and Pit 43 (32.5 Mt), which only delivers an additional 4.1 Mt of ore, giving a waste stripping ratio of greater than 8:1.

Au Cut-off

Cu Cut-off

Process Au Ag Cu Au Ag Cu Au Ag Cu

(%) (%) (%) (%) (%) (%) (%) (%) (%)

Heap Leach 70% 7% 30%

Flotation 70% 80% 90%

ALP 80% 7% 30%

Processing Route

Heap Leach ALP Flotation

< 2 ppm >= 2 ppm < 2 ppm

< 0.35 % >= 0.35%

Recovery Recovery Recovery

In Situ

(tonnes) (Au g/t) (% Cu) (Ag g/t) (Au Oz) (t Cu) (Ag Oz)

Proven 16,733,000 1.12 0.52 7.63 600,000 87,000 4,105,000

Probable 3,761,000 0.68 0.40 6.12 82,000 15,000 740,000

Total 20,494,000 1.03 0.50 7.35 682,000 102,000 4,845,000

Reserve

Category

In Situ Grades Contained Metal

Ore Reserve




Although the expansion capital for an additional pit expansion would be minimal if contract mining were used it would result in a significant amount of advance stripping and the discounted value of the expansion makes it uneconomic under the current assumptions. Should the metal prices improve, then the geometry of the pit expansion would lend itself to a final pushback as there would be sufficient access width on all benches.

Using the selected pit design, a number of pushbacks have been created which, when scheduled to meet the constraint that the Flotation plant will only be available in mid-2016, mean that it is necessary to limit the mining of material that will be allocated to Flotation so as to maximise the ore exposure of material that is suitable for HL or agitation leach.

Material that has a high copper content (> 0.35% Cu) and lower gold content (Au < 2 g/t) would normally go to Flotation. However, prior to commissioning of the Flotation plant this material will have to be stockpiled. This appears to be possible provided there is a stockpile capacity available of up to 1.2 Mt.

The Life of Mine (LOM) Schedule demonstrated that a practical blend of materials can be achieved that will meet the constraints on maximum plant capacity plus minimise flotation material prior to completion of the new plant. A consequence of this schedule is that the material allocated to ALP will be exhausted early, which will mean that the operation can be run with just HL and Flotation for the last two years as it would be uneconomic to continue with ALP.

Besides evaluating the economics as part of the pit optimisation, the LOM schedule has been evaluated using AIMC’s own financial model. This confirms that the selected pit is economic and is broadly in line with the valuation (before tax) produced by the pit optimiser.

It should be noted that this report refers to an Environmental audit that was conducted in August 2014. This concluded that;

The mine needs to drastically reduce the cyanide concentration in the TSF and the rate of seepage from the toe.

The mine closure plan needs to be updated in the near future to account for the changes since 2012 and measures noted above should be implemented.

The above measures need to be implemented in order to be in compliance and consequently are required for the declaration of a Reserve. On this basis it is concluded that, as at 1 September 2014, the remaining Ore Reserve for the Gedabek open pit is 20.5 Mt, with a contained metal content of 682,000 oz of Au, 102,000 t of Cu and 4,845,000 oz of Ag.


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Table of Contents

1 Executive Summary .........................................................................................................1-2

2 Credentials of the Author ................................................................................................2-7

3 Introduction ......................................................................................................................3-8

4 Resource Model ............................................................................................................. 4-11

5 Modifying Factors .......................................................................................................... 5-12

5.1 Mining Recovery and Ore Dilution ........................................................................................ 5-13 5.2 Geotechnical Parameters ..................................................................................................... 5-14 5.3 Metallurgical Factors ............................................................................................................ 5-17 5.4 Financial Parameters ............................................................................................................ 5-18 5.5 Legal Tenure......................................................................................................................... 5-19 5.6 Environmental Audit ............................................................................................................. 5-20

6 Pit Optimisation .............................................................................................................. 6-21

6.1 Results .................................................................................................................................. 6-21

7 Mine Design .................................................................................................................... 7-25

7.1 Pit Limit Design ..................................................................................................................... 7-25 7.2 Pushback Design .................................................................................................................. 7-29 7.3 Reserves ............................................................................................................................... 7-32

8 Scheduling ...................................................................................................................... 8-34

8.1 Plant Production ................................................................................................................... 8-34 8.2 Mine Production .................................................................................................................... 8-36

9 Conclusions and Recommendations ............................................................................ 9-39

10 References .................................................................................................................... 10-41

11 Compliance Statement ................................................................................................. 11-42




List of Abbreviations/Acronyms

Abbreviation/Acronym Meaning

Ag Silver

AIMC Azerbaijan International Mining Company

ALP Agitation Leach Plant

ARD Acid Rock Drainage

Au Gold

Cu Copper

g/t gram per tonne

HL Heap Leach

Kt kilo

L litre

L/s litre per second

lb pound

LOM Life of Mine

mg/L milligram per litre

Mt million tonnes

NPVS NPV Scheduler

NSR Net Smelter Return

OSA Overall Slope Angle

oz ounce

SART Sulphidization - Acidification-Recycle -Thickening

t tonne

tpa tonne per annum

TSF Tailings Storage Facility

WAD Weak Acid Dissociable Cyanide


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2 Credentials of the Author

Dr Matthew Randall

Mining Engineer – Associate of CAE Mining Africa

Matthew is a Mining Engineer who graduated from the Camborne School of Mines, UK in 1978 with a BSc (Hons) degree in Mining. Following a period of time in industry, he studied for a doctorate in Rock Mechanics at the Camborne School of Mines, graduating in 1989.

In 1978, Matthew started his mining career as a mining engineer with JCI, gaining experience in deep underground mining at a number of their properties in South Africa. During this period, he worked his way up through the Learner Official Training Course (LOTC), with time as a contract miner. In 1980, Matthew moved to Palabora Mining Company (PMC) in the then Northern Transvaal, working in the largest open pit mine in the world at that time.

After leaving PMC in 1983, Matthew followed a programme of part-time study at the Camborne School of Mines towards a doctorate. During this time he was employed as a senior lecturer and also worked on the Hot Dry Rock Geothermal Project.

In 1990, Matthew joined Rio Tinto Technical Services (RTTS) in Bristol, UK as a principal consultant and continued to work for them until 2008. During this time he was seconded to operations in Papua New Guinea, Australia, Spain and the USA. The majority of this rime was spent as Chief Mining Engineer, or Technical Services Manger, looking after the mine engineering services.

In 2008 Matthew returned to the UK and spent three years working for CAE Mining as Group Mining Engineer. Subsequently, in July 2011, he set up his own mining consultancy business.

Matthew is the joint director of Axe Valley Mining Consultants and has worked on a number of major projects across the globe over the past three years, including commodities such as iron ore, rutile, copper and gold.

Matthew regularly acts as the Competent Person (CP) for a number of clients (base metal and industrial minerals), and is responsible for declaring the Reserve statement for these properties.




3 Introduction

AIMC requested CAE Mining to prepare an Ore Reserve Statement for the open pit at Gedabek Mine. The general layout for the mine is shown in Figure 3.2 and consists of:

Existing open pit

Crusher

Processing plants (HL and ALP)

Stockpiles

Waste dumps

Core shed

Camp

Main access roads.

Although the mine is currently primarily focused on the extraction of gold, with bi-products of copper and silver, the intention is to build a new Flotation plant to the north east of the crusher to take advantage of the high copper grades that are seen at Gedabek. However, this plant will only be available in mid-2016 and this must be taken into account when designing the mining sequence and evaluating the pit.

The agreed process configuration is shown in Figure 3.1 in terms of the cut-off grades.

Figure 3.1 Simplified process configuration for Gedabek

Yes

No

No Yes

No Yes Yes No

Mine Production

Heap Leach Flotation Agitation Leach Waste Dump

Au ≥ 2.0 ppm

Cu ≥ 0.35 %

NSR ≥ 6.71

Cu ≥ 1.0 %


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A full description of the location of the property and the geological setting is given in CAE Mining’s Resource Statement of May 2014 and is not repeated here.

For the purposes of the pit optimisation, the minimum cut-off grade for HL was calculated on the basis of the NSR such that the cut-off between ore and waste is where the NSR is greater than or equal to processing plus fixed costs.

An upper limit of 1% copper was also put on the ALP and this material was re-directed to Flotation. This is only a relatively small amount of material (800,000 t) and is therefore not significant.

A technical audit was carried out by the author during the week of 28 August 2014 where the mine operations and processing plants were reviewed. Discussions were also held with the Geotechnical consultant from CQA regarding the slope parameters that were to be used for pit design. This resulted in an updated report being issued with a recommended Overall Slope Angle (OSA) of 37.3°.

During the audit, the Competent Person also confirmed with AIMC the modifying factors to be used for:

Tonnage and grade estimates

Process recoveries

Metal prices

Mining and processing costs

Fixed costs.

The main objective of this report is to document the procedure used to determine the Ore Reserve and ensure that this estimate follows the guidelines set down by JORC (2012).




Figure 3.1 General location plan for Gedabek mine

Note: The name Gedabey is used interchangeably with Gedabek and they refer to the same deposit.


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4 Resource Model

The Resource model used for this exercise was aimc_mod_aug14.dm, which was issued by CAE Mining in August 2014. This model is an update to the model issued in May 2014 (aimc_mod_aug14.dm) and includes modifications requested by AIMC that address changes to the Resource Estimate for Ghadir only.

As confirmed by CAE Mıning, the August 2014 model update does not affect the estimates within the area of the open pit and therefore the May or August 2014 models are equally valid for this work.

The Resource Statement for the Gedabek deposit (including Ghadir) is shown in Figure 4.1.

Figure 4.1 Resource statement for the Gedabek deposit (May 2014)

Oxide MaterialTonnes Au (g/t) Cu (%) Ag (g/t) Au (Oz) Cu (tonnes) Ag (Oz)

Measured 12,020,215 0.750 0.140 6.060 289,817 16,790 2,342,022

Indicated 6,169,672 0.602 0.154 5.139 119,508 9,520 1,019,331

Measured & Indicated 18,189,888 0.700 0.145 5.748 409,326 26,310 3,361,353

Inferred 4,436,283 1.201 0.101 4.790 171,262 4,489 683,227

Transitional MaterialTonnes Au (g/t) Cu (%) Ag (g/t) Au (Oz) Cu (tonnes) Ag (Oz)

Measured 4,714,927 0.691 0.135 4.705 104,804 6,385 713,166

Indicated 2,717,196 0.602 0.143 4.044 52,599 3,892 353,244


Inferred 617,313 1.243 0.106 4.405 24,679 654 87,424

Sulphide MaterialTonnes Au (g/t) Cu (%) Ag (g/t) Au (Oz) Cu (tonnes) Ag (Oz)

Measured 20,454,539 0.894 0.334 6.088 587,676 68,226 4,003,615

Indicated 15,719,225 0.584 0.249 4.012 295,131 39,083 2,027,436


Inferred 4,391,322 0.692 0.173 4.733 97,736 7,586 668,273

Total MineralisationTonnes Au (g/t) Cu (%) Ag (g/t) Au (Oz) Cu (tonnes) Ag (Oz)

Measured 37,189,682 0.822 0.246 5.904 982,298 91,401 7,058,803

Indicated 24,606,093 0.591 0.213 4.298 467,239 52,495 3,400,011

Measured & Indicated 61,795,775 0.730 0.233 5.264 1,449,537 143,896 10,458,814

Inferred 9,444,918 0.967 0.135 4.739 293,678 12,729 1,438,924




5 Modifying Factors

The modifying factors to be used in the Reserve estimate are summarised in Table 5.1 and Table 5.2 . These follow the guidelines set out in the JORC code.

Table 5.1 Modifying factors used to determine the ore reserve

Factor Comments

Cut-off Grade The minimum cut-off grade was determined from a combination of the NSR and copper grade. Lower grade material is either directed to HL or Flotation. High grade material (Au ≥ 2 g/t) is sent to ALP, or Flotation if Cu ≥ 1%.

Mining Factors The Resource was converted to a Reserve by optimising the pit limit with NPV Scheduler (NPVS); the engineered pit design was based on the optimised pit limit, which includes provision for ramp access. The mining method is based on the existing truck and shovel operation and assumes a fixed contract mining rate. Similar equipment will be used in the future and a maximum mining rate of 9 Mtpa of rock has been assumed in the schedule.

Mining Recovery & Ore Dilution

The mining recovery and waste dilution factors were modelled using historical reconciliation factors for specific grade ranges (see Section 5.1 for more detail).

Geotechnical Parameters

The OSA and other geotechnical parameters are as per the recommendations of the Geotechnical consultant (see Section 5.2).

Metallurgical Factors The metallurgical recovery is specified for each element (gold, silver and copper) and each process route (HL, ALP and Flotation). These recovery relationships are based on historical performance of existing plant, or published reports (see Section 5.3).


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Table 5.2 Modifying factors used to determine the ore reserve (continued)

Factor Comments

Costs, Revenue and Smelter Contracts

The cost and revenue factors used in the pit optimisation are shown in Section 5.4. The NSR is calculated by taking into account the selling cost for copper concentrate (sold at gate) and the selling costs and transport costs for Dore.

Market Assessment The pit optimisation is based on a gold price of 1,200 US$/oz, 18 US$/oz for silver and 3 US$/lb for copper. No limit has been put on either Dore or copper concentrate production.

Legal & Other Legal tenure is held on the relevant mining licenses. See Section 5.5.

Classification The Reserves are inclusive of the Resources reported in May 2014 and 100% of the Measured and Indicated Resources in the pit design have been converted into Proven and Probable Reserves respectively.

Audits An environmental audit was conducted and is referenced in this Reserve Statement (see Section 5.6).

Relative Accuracy and Confidence

The modifying factors used in this report are considered to be known to an acceptable level of accuracy. The Reserve Statement is based on a mine schedule that takes into account the pushback designs and process configuration constraints and has, therefore, been shown to be practical as well as economic.

The details of the key modifying factors are discussed in more detail in the following sections of the report.

5.1 Mining Recovery and Ore Dilution

The low grade nature of the deposit, in conjunction with the complex geological setting, makes it very difficult to apply global factors for mining recovery and ore dilution and it has been observed that the correlation between the geological model and actual production can vary considerably on a bench-by-bench basis.

A report prepared by SRK (2014) considered the comparison between the geological model and blast-hole samples. They concluded:




“Currently, the mine applies a 1.5 reconciliation factor to account for the differences between the production model and the CAE resource model. This quite high factor is based on differences noted between the production and the resource model throughout the last two years of production.

As presented in Table 5.3, those differences are especially acute for higher (1.0 - 1.5 g/t) cut-offs with the overall metal content 50% higher from production. This is in line with the current reconciliation factor applied at the mine. From Table 5.3 it can be further concluded that this factor would have to be much lower if the actual cut-off applied at the mine was 0.3 g/t, with only 16% difference in metal content. In short, the factor is directly related to the currently used cut-off grade.”

Table 5.3 Comparison of block grades compared to 2013 production model

The approach taken has been to use the historical reconciliation factors and apply the relevant factor for tonnes and grades to each block. Any discrepancy between predicted ore tonnes and actual ore tonnes is accounted for by a commensurate adjustment to the waste tonnes so that the mass balance is maintained. For now this is done externally in the schedule on the basis of the number of tonnes mined in each grade range.

5.2 Geotechnical Parameters

The original Pre-Feasibility Study (PFS) analysis of the pit slopes was conducted by Knight Piesold (2007), who concluded:

“Geological information indicates the ore will primarily be present in lower levels of the pit, which will have two main rock types; volcanics overlying porphyry. Rock core observations and rock strength tests indicate reduced strength in ore-bearing rock in lower pit walls, so buttressing capabilities of the wall rock located near the bottom of the pit is expected to diminish. This is addressed in stability calculations carried out for the low


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and high pit walls, and overall angles of 45 ̊ to 56 ̊ were examined in the porphyry and volcanic rocks. Controlled blasting of final pit walls is advised.

it is currently understood that groundwater will not play a significant role in pit slope stability due to the elevation of the pits above the surrounding terrain, in conjunction with the low rainfall and consequently low amount of groundwater recharge expected. However, rainfall contacting exposed ore may become contaminated and may require containment within the process ponds or storm water pond. This water may be used as raw water makeup or will require processing by the detox plant prior for disposal to a local watercourse.”

Although Knight Piesold noted that the area has a high probability of seismicity, this does not seem to have been an issue when they recommended the slope angles. Their conclusion was as follows:

“Pit slopes will be in relatively strong volcanic rock overlying weaker porphyry rock. Slope stability analyses showed overall angles for a 210 m high wall of 45° to 50° will be stable. Drilling results suggest groundwater will not be met until near pit bottom. Slope angles can be adjusted during mining as further geotechnical data becomes available, however, steeper porphyry slopes may not be possible.”

Following localised wall failures in the open pit, the Geotechnical engineers used by AIMC (CQA Consultants Ltd) conducted a full site evaluation and prepared a report in 2013 for the southern pit slopes.

This concluded that much shallower slopes would be required:

“In general, there appear to be stability issues with the lower benches from approximately 1720 m to 1650 m elevation, in certain locations AAM have reduced slope angles to increase stability. To further enhance this and improve dynamic stability (earthquake loading) CQA proposes that the southern mine benches are re-constructed to 11 benches, with a 10 m reduction in level between each bench.” “The bench width in the lower benches should be increased to 15 m. This will allow for some failure of the weathered material. In the higher benches where the material is more competent bench width maybe reduced to 7.5 m. This will allow for some localised failure but will reduce the impact of this upon the works. The overall slope angle should be 25◦ this will increase the stability of the overall slope, preventing large-scale failure during earthquake loading”




The AIMC site engineers have typically designed the slopes to 40° without issues. Therefore, CQA’s 2013 findings were queried by the author during the site visit. This resulted in a further evaluation of the slopes by CQA in 2014 and the following recommendation:

“We are pleased to present here our opinion of CAE’s proposal for operational pit slopes, based on our interpretation of the rockmass characteristics.

Our opinion is based on an assessment of static conditions (i.e. without major earthquakes). An overall slope angle of 37.3 degrees (crest to toe) would be appropriate for the site, with full slope heights up to 150m.

This will need to be reviewed if continued excavation exposes large pockets of poor quality rock (for example extending more than 30m horizontally and vertically), especially if the very poor quality rock is found towards the base of the pit slopes. Very poor quality rock means highly weathered material that appears more like a soil, albeit with many rock fragments, rather than a rockmass of interlocking pieces.”

Based on a 10 m bench height, the slope parameters tabled in Table 5.4 have been accepted by CQA and have been used to design the pit limit and estimate the ore reserve.

Table 5.4 Recommended Geotechnical Parameters

Parameter Units Value Comments

Batter Angle degrees 55.0 Toe to crest

Safety Berm metres 6.0 Berm on each bench

Geotech Berm metres 10.0 Catch bench every 50 m

Stack Height metres 50.0 Max wall height between geotech berms

These parameters have been applied to all walls irrespective of orientation.


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5.3 Metallurgical Factors

The process recoveries for the various process routes are specified for gold, silver and copper. For HL and ALP it is assumed that the SART process is used.

The SART process, developed jointly by SGS Lakefield and Teck Corporation, can remove the metallurgical interference of leachable copper and silver, and regenerate cyanide so that it can be recycled to the gold operation. The claimed benefits (SGS, 2007) are:

The revenues received from the sale of the copper sulphide precipitate will likely exceed the operating costs of the SART process. Therefore, the process can add value to a project to the extent that copper leaches naturally from the ore during gold leaching.

The cyanide associated with the copper complex is recycled as free cyanide, available for further leaching. The cost of this cyanide is part of the overall cost of the SART process, which is more than covered by revenues from copper sales, i.e. this represents a source of zero cost cyanide.

The alternative to SART would be to process the HL liquor either periodically during the operating life of the mine or at the end of the project, with a cyanide detoxification process such as the SO2/air process. The significant costs associated with this process will be avoided by incorporation of the SART process.

If the SART process is installed ahead of the gold recovery process, the removal of copper and most of the silver in SART significantly simplifies the gold recovery operation, whether it is by adsorption of gold on carbon or resin, or cementation on zinc powder.

It is interesting to note that SGS (2007) reported that the recovery of gold from heap leaching is relatively insensitive to grade in the range 0.76 g/t to 3.73 g/t Au. A recovery of 70% was achieved with half-inch particle size, and this increased to 90% with a particle size of 75 microns.

SGS (2007) also reported:

“Cyanide consumption is high, owing to the presence of cyanide soluble copper minerals in the Gedabek ore body. Consumption ranged from 3 to 6 kg/t for different samples from the Gedabek deposit that have been tested in this study. A mineralogical investigation revealed that the dominant copper mineral is chalcopyrite, which is insoluble in cyanide solution. However, significant quantities of chalcocite, cuprite, bornite and enargite were identified, all of which are at least partially soluble in cyanide solution.




As a means of mitigating the high costs associated with excessive cyanide consumption, the SART process was recommended. The process breaks down the copper cyanide complex that is produced in the leach liquor, converting it to an insoluble precipitate of copper sulphide (Cu2S).”

The process costs and overall recoveries used (including the contribution from SART) are shown in Table 5.5.

Table 5.5 Metallurgical recovery factors

5.4 Financial Parameters

The financial parameters used to determine NSR and block values are shown in Table 5.6.

Table 5.6 Financial parameters

The selling price is deduced from the market price to determine the NSR. The values used are specified by process route and product shown in Table 5.7.

It should be noted that the cyanide used in HL is regenerated cyanide from the SART process and that, therefore, there is no cyanide cost in HL processing cost.

Au Cut-off

Cu Cut-off

Process Au Ag Cu Au Ag Cu Au Ag Cu

(%) (%) (%) (%) (%) (%) (%) (%) (%)

Heap Leach 70% 7% 30%

Flotation 70% 80% 90%

ALP 80% 7% 30%

Processing Route

Heap Leach ALP Flotation

< 2 ppm >= 2 ppm < 2 ppm

< 0.35 % >= 0.35%

Recovery Recovery Recovery

Parameter Units Value

Total G&A US$/t ore 5.56

ALP Cost US$/t processed 32.84

Heap Leach cost US$/t processed 1.15

SART cost US$/t processed 2.88

Flotation cost US$/t processed 17.36

Mining cost US$/t mined 1.92


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Table 5.7 Selling costs

It can be shown that the theoretical cut-off grade for HL is 0.14 g/t Au and 1.0 g/t Au for ALP. The copper and silver contribution in this example is based on the empirical relationship between copper and silver, and gold where:

Au = Au (g) produced, Cu (t) = (1045/840000)/2 and

Ag (g) = (Au/5) + (Cu x 1000 / 31.1034) x 1.4

Based on the assumed grades of 8 g/t for silver and 0.15 g/t for gold the calculated cut-off for Flotation is 0.35% Cu. This confirms that the fixed cut-off grade of 0.35% Cu used to define Flotation feed is reasonable and will ensure that material sent to Flotation will be economic at 0.35% Cu.

5.5 Legal Tenure

Legal tenure of the Gedabek deposit is held by AIMC under a scheme known as PSA. This is explained on the AIMC website and the relevant sections are shown below;

“The PSA is a revenue sharing contract modelled on Azerbaijan’s oil contract system, which was established to ensure the country benefited from its oil wealth in the presence of international oil company investment and resource exploitation.

The PSA grants the Company a number of periods to exploit defined license areas, known as Contract Areas, agreed on the initial signing with the Azerbaijan Ministry of Ecology Natural Resources (MENR). The exploitation period, provided for in the PSA, for early exploration of the Contract Areas to assess prospective results, can be extended.

A development production period commences on the date that the company issues a notice of discovery and runs for 15 years with two extensions of five years each at the option of the company. Full management control of mining in the Contract Areas rests with AIMC.

Au Ag Cu

Agitation Plant (Dore) 99.5% market price - -

Heap Leach (Dore) 99.5% market price - -

Agitation Plant (SART) 81% market price (SART) 81% market price (SART)

Heap Leach (SART) 81% market price (SART) 81% market price (SART)

Flotation 84% market price 84% market price 84% market price

Selling price - Net of refining and transportationProcesses




Under the PSA, AIMC is not subject to currency exchange restrictions and all imports and exports are free of tax or other restriction. In addition, MENR is to use its best endeavours to make available all necessary land, its own facilities and equipment and to assist with infrastructure.”

AIMC provided a copy of the five-year PSA extension (Permit No 073738) dated 25 December 2013.

5.6 Environmental Audit

An environmental audit was carried out by Daniel Limpitlaw (Datamine, 2014) when he visited site between 27 and 29 August 2014. The conclusions from this visit were:

Cyanide - Discharge from the Tailings Storage Facility (TSF) is a risk as there is seepage at the toe of the impoundment. Discharge was planned to be less than 50 mg/L Weak Acid Dissociable Cyanide (WAD) but current discharges are in excess of 1,000 mg/L WAD. Total cyanide has not been measured but is suspected to be close to 2,000 mg/L.

The discharge flow rate is higher than planned and is currently 15 L/s instead of less than 1 L/s.

Mine Closure - AMEC (2012) recommended the creation of artificial subsoil by crushing waste rock. This recommendation has not been implemented to date. Waste rock resources are not currently placed in accordance with a plan for pit closure. Should the mine close in the short term, rehabilitation to good international industry practice will not be possible due to the topsoil deficit.

Acid Rock Drainage (ARD) - ARD is generated at the open pit and in the stockpiles. The impacts of this drainage are detectable up to 5 km downstream from the mine. As the pits expand and future underground operations commence, it will be increasingly difficult to apportion responsibility and the mine is likely to have to assume responsibility for all ARD.

The required actions that address these issues are:

The mine needs to drastically reduce the cyanide concentration in the TSF and the rate of seepage from the toe. The TSF consultant (CQA) has written a report on how this might be achieved.

The mine closure plan needs to be updated in the near future to account for the changes since 2012 and measures noted above should be implemented.

The above measures need to be implemented in order to be in compliance and consequently are required for the declaration of a Reserve.


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6 Pit Optimisation

The pit optimisation was run in NPVS, which uses the standard Lerchs-Grossman algorithm to determine the pit limit and incremental pit shell. The latter are used as a guide for selecting the pit limit and as the basis for the creation of a sequence of pushbacks within the pit limit.

The main input parameters to NPVS are:

Product prices (gold, silver and copper)

Selling prices

Mining cost

Process cost (by process route)

Process recovery (by process route

Slope parameters.

NPVS was setup so that the rock types are either oxide, transition or sulphide and these are further subdivided into Measured, Indicated and Inferred for reporting purposes. When determining the pit limit and Reserves, the grades for the Inferred material are given a value of zero as they cannot be included in the valuation. However, it is useful to report these values as they represent a potential ore source should it be possible to reclassify them in the future.

The parameters used on the optimisation are discussed in Section 5 and a summary of the results from the pit optimisation are discussed below.

6.1 Results

The pit optimisation was run with an increment of 1% for Price Factor so as to determine if there was a logical breakpoint at which to select the pit limit. Note that at a Price Factor of 100% the metal prices will be equal to the assumed prices presented in Section 5 of this report.

It can be seen in Figure 6.1 that at Pit 37 (86% Price Factor) there is a significant jump in total rock movement and that the total rock increases from 54 Mt to over 90 Mt. The increase in ore tonnage (and NPV) over this increment is relatively small as more than 95% of the final value has already been achieved.

On inspection, it was evident that the pit expansion beyond Pit 37 was more or less limited to an area to the southwest of the main pit, as shown in Figure 6.2 and could, therefore, be treated as a potential expansion for the future if prices rise.




It can also be seen from Figure 6.2, that there are areas around the fringes of the deposit that are marginal and these will be excluded from the mine design. The exception is the satellite pit to the south of the main pit, which forms a sizeable pit in its own right and can be mined independently as required.

Overall the pit optimisation is “well behaved” and provides a good framework from which a detailed mine design can be produced. This mine design will take into account the detailed geotechnical parameters of batters and berms as well as ensuring that there is access space to develop the mine.

Pit 37 has been selected as a suitable point from which the mine design can commence. This does not preclude the opportunity to further expand the pit to Pit Shell 43 (or further) whilst ensuring that the project value has been maximised within the practical constraints such as fleet capacity.

It should be noted that an expansion beyond Pit 37 will require a significant increase in mining fleet capacity and the lead time to expose ore may exceed two years. When this is taken into account, it is estimated that at current prices that this ore will not be economic on a discounted basis. The ore reserve should, therefore, be based on Pit 37 and not one of the larger shells.


© 2013 CAE Mining

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Figure 6.1 Plot of cumulative ore tonnage and NPV vs pit shell number




Figure 6.2 Optimised pit limit shown in blue with potential expansion in grey


© 2013 CAE Mining

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7 Mine Design

Based on the selected pit limit described in Section 6, mine designs were prepared for:

Final pit limit

Interim pit stages/pushbacks.

The pit limit and pushbacks are designed according to the geotechnical parameters discussed in Section 5. It should be noted that the total tonnage within the pit limit will vary slightly from that shown in the optimisation due to the batter angle and smoothing of the wall to avoid potential geotechnical issues with “noses” etc.

The designs are discussed below along with the resulting reserves.

7.1 Pit Limit Design

The final pit wall has been designed to include a 10 m-wide catch bench every five benches (50 m). This bench acts as a haul road for the 30 t trucks so that they can exit either side of the slope and link to the roads to the waste dumps or crusher, as shown in Figure 7.1.

Due to the limited width of each pushback, the pit wall has to be mined bench by bench from the top down and therefore ramp access is not an issue as the material can exit on the active bench. There is limited capability to mine multiple benches as long as temporary ramps are maintained to the nearest catch bench.

Some internal ramps will be required where the pit bottom is below the lowest pit exit at the valley floor. These ramps are relatively short and can be positioned as part of the short term planning work.

It can be seen from Figure 7.1 that the mine is split into three distinct areas and that the main area (existing mine) can be further subdivided into pits that can be developed more or less independently. This follows the current design for Pits 1, 4 and 6 and is discussed further in Section 7.2.

Besides determining the optimal extent of the open pit, an important aspect of the mine design is the distribution of material types within this pit. This is shown in Figure 7.2 and Figure 7.3 in terms of the assigned process route (HL, ALP or Flotation). The mine sequence is constrained by the need to build the Flotation plant and to match the plant throughput constraints for each process.

This is discussed in more detail in Section 8.




Figure 7.1 Pit limit design based on Pit 37


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Figure 7.2 Distribution of material types with the pit limit




Figure 7.3 Distribution of material types, classified by processing route


© 2013 CAE Mining

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7.2 Pushback Design

The main constraints on the design of the pushbacks are:

Slope design parameters (Table 5.4 )

Bench access to pit exits at all times

Minimum bench width for equipment (30 m)

Maximum bench sinking rate (12 benches per year)

Blending to plant feed requirements

Maximum stockpile capacity.

As the Flotation plant will only be commissioned in mid-2016 and the maximum capacity of the feed stockpile for Flotation has been set at 1.2 Mt, the mining sequence needs to minimise the exposure of Flotation material (i.e. high copper content material) so that the mining cost (including stockpiling) is minimised.

This means that the initial mining stages need to focus on the areas that have lower copper grades (< 0.35% Cu), which can be fed to HL or ALP. Ideally this should match the planned throughput of each plant so that stockpile rehandle is minimised due to sustained peaks over long periods (months) of one or other material type.

The limited reserves of ALP will necessarily mean that the supply of ALP material will tail off and it will probably not be economic to keep this plant running at a very low feed rate. In this case, the remaining material allocated to ALP will need to be fed to HL or Flotation.

Besides the constraints on Flotation and ALP, there is also a constraint on the maximum tonnage that can be fed to HL due to the limited space available. Initially, this was initially set at 5 Mt but, as a result of running the pit optimisation, it became apparent that this capacity needs to be increased to at least 8 Mt. The maximum capacity for HL has not been set as a hard constraint as it is possible to reuse the leach pad area if necessary, and/or an on/off leach pad may be used for higher grade material that is crushed to a finer particle size for maximum gold recovery. This should be investigated once detailed planning is undertaken.

The sequence of pushbacks is shown in in Figure 7.4 and the reserves by pushback are shown in Figure 7.5 and Figure 7.6.




Figure 7.4 Pushback sequence


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Figure 7.5 Tonnage by pushback (incremental)

Figure 7.6 Tonnage by pushback (accumulated)

It is evident that the distribution of material types by pushback is far from even and that this will be a major issue when scheduling. It can also be seen that the waste stripping ratio varies considerably between pushbacks and again this will need careful sequencing in order to smooth out the fleet requirements and produce a practical schedule.




7.3 Reserves

A comparison of the reserves for the selected pit limit (Pit Shell 37) and the detailed pit design in Table 7.1 shows that there is less than a 2% variance in the total ore within the pit. The pit design has, therefore, followed the guidelines provided by the pit optimisation with minimal loss of ore as a result of imposing practical mining constraints on the design.

The Reserves are summarised in Table 7.1 in terms of the Reserve categories of Proven and Probable, where Proven and Probable relate directly to Measured and Indicated.

Table 7.1 Reserve summary (as at 1 September 2014)

Note: tonnes (dry) and grades shown are after applying reconciliation factors.

The total waste (including mineralised material that is uneconomic) within the pit is 34.2 Mt, giving a total rock tonnage of 54.7 Mt and an average waste stripping ratio of 1.7. This is relatively low for an open pit and highlights the low grade nature of the deposit and the need to mine to a very low cut-off grade.

The potential for expanding the reserves lies with:

Upgrading the Inferred Resource (additional 0.7 Mt)

Expanding the pit beyond Pit Shell 37 (additional 4.1 Mt)

It should be noted that the method of applying reconciliation factors to the in-situ quantities and qualities assumes that the factors are applied equally to all grades (gold, silver and copper). This approach has some merit in that it attempts to correct for “over smoothing” of the Kriged Resource Estimate, but the factors are not necessarily representative of the whole deposit. Further investigation is required to improve the resource estimates through more drilling and perhaps tighter control on grade domains and/or other geostatistical controls.

In Situ

(tonnes) (Au g/t) (% Cu) (Ag g/t) (Au Oz) (t Cu) (Ag Oz)

Proven 16,733,000 1.12 0.52 7.63 600,000 87,000 4,105,000

Probable 3,761,000 0.68 0.40 6.12 82,000 15,000 740,000

Total 20,494,000 1.03 0.50 7.35 682,000 102,000 4,845,000

Reserve

Category

Ore Reserve

In Situ Grades Contained Metal


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Table 7.2 Comparison of reserves for the pit design and the optimised pit limit (Pit Shell 37)

Notes: 1. Process routes are ALP, HL and Flotation 2. Tonnes and grades are after reconciliation factors have been applied to the model 3. Pit optimisation only uses the OSA and is based on the parent cell size 4. Pit design uses the slope design parameters, including batter angle, berms and max stack height

ALP ALP ALP ALP

Tonnes 2,590,913 Tonnes 2,590,913 Tonnes 2,651,323 Tonnes 2,651,323

In situ Gold (g) 8,514,322 In situ Gold (g/t) 3.29 In situ Gold (g) 8,713,164 In situ Gold (g/t) 3.29

In situ Copper (t) 10,707 In situ Copper (%) 0.41% In situ Copper (t) 10,900 In situ Copper (%) 0.41%

In situ Silver (g) 46,753,138 In situ Silver (g/t) 18.05 In situ Silver (g) 47,705,772 In situ Silver (g/t) 17.99

Flotation Flotation Flotation Flotation





Heap Heap Heap Heap





Total Ore Total Ore Total Ore Total Ore





Pit Design - Ore: in situ Pit Optimization - Ore: in situ




8 Scheduling

Using the pit design and pushback sequence described in Section 7 a Life of Mine (LOM) schedule was created in order to demonstrate that an acceptable mining sequence can be achieved, whilst honouring the various constraints.

The main constraints imposed on the schedule are:

HL throughput (1.1 Mtpa)

ALP throughput (840,000 tpa)

Flotation throughput (1.5 Mtpa)

Maximum stockpile size for Flotation (1.2 Mt)

Flotation plant commissioned in 2016.

The capacity of the Flotation plant is assumed to build up from a pilot stage testing in 2016 (300,000 t), to 800,000 t in 2017 and 1.5 Mtpa from 2018 onwards.

The destination for each ore block depends on a set of constraints for gold and copper grade as described in Figure 3.1. This assumes there is no discernible lithological differentiation between ore types (oxide, transition and sulphide).

The resulting schedule is shown in Section 8.1, which was used by AIMC as the input to their financial model.

8.1 Plant Production

The schedule shown in Table 8.1 demonstrates that the ore production rate can be ramped up from 1.6 Mtpa to 2.87 Mtpa over the period 2015 to 2019. This requires a significant amount of stockpiling of Flotation material in the first two years as the Flotation plant only starts to ramp up from late 2016 onwards.

The other issue is the fact that it is impossible to access sufficient ALP to meet the plant capacity in the first two years and it has been decided to run it at a lower rate over a longer period (2015 to 2020). Even then, there is some residual ALP mined in 2021 and 2022 which will need to be fed to Flotation or HL, depending on the copper grade.


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Table 8.1 Mine Schedule

Schedule Plant feed Stockpile levels

Plant Heap Plant ALP Plant Flotation Schedule Total Ore Stock Heap Stock ALP Stock Flotation

2014 330000 0 0 330000 373696 279097 332304

2015 950 000 350 000 300 000 1 600 000 24 661 194 008 942 443

2016 1 000 000 350 000 800 000 2 150 000 1 094 23 320 1 206 583

2017 1 070 000 400 000 1 100 000 2 570 000 221 844 466 626 708 530

2018 1 070 000 400 000 1 200 000 2 670 000 158 859 469 972 67 988

2019 1 070 000 400 000 1 400 000 2 870 000 504 432 153 205 11 799

2020 1 070 000 400 000 1 400 000 2 870 000 564 707 36 064 38 464

2021 1 120 370 - 1 450 000 2 570 370 337 113 166 474 407 314

2022 1 120 370 - 1 450 000 2 570 370 - 290 912 2 234




8.2 Mine Production

The total material movement as shown in Table 8.2 indicates that the required fleet capacity peaks at 7.8 Mt of rock in 2016 and then gradually declines. This is well within the maximum fleet capacity of 9 Mtpa that was set for the contractor and should not pose a problem.

To attain the smoothed feed profile to the plant movement as shown in Figure 8.1, it is necessary to have a number of stockpiles, the largest of which is the stockpile for Flotation. It may be possible to reduce these amounts with further detailed scheduling but at this level of study it is considered preferable to manage the fluctuations in the mix of materials by stockpiling so as to simplify the mining sequence.

As noted previously, a decision was also made to smooth out the production for ALP at 400,000 tpa rather than peak at 840,000 tpa in 2017 and 2018. This would have meant either running ALP at a very low rate after 2018 or feeding the ALP material to HL or Flotation. This decision will be influenced by the possible introduction of a separate underground operation at Ghadir, which will change the strategy for ALP. Therefore, it is better to wait before making a firm decision on how to run the ALP plant.

Stockpiling of HL material is less of an issue provided additional space can be found for the amounts that have to be stockpiled. These areas can then be processed as required.

The variation in mined grade (in-situ) by material type is shown in Figure 8.1. This shows that the grade (gold or copper) going to HL is relatively well behaved and does not pose many issues for grade control. The mined grades to Flotation, on the other hand, vary considerably by period. As these fluctuations are relatively short term (less than 12 months) this can be accommodated by blending from the stockpile provided it is segregated by grade.

The variation in grade (gold or copper) for ALP is of less concern, particularly for gold, and can easily be managed by the stockpiles.

Overall, the issues are more to do with trying to expose the right mix of materials (HL, ALP and Float) at all times, depending on the configuration of the process plants. This results in substantial stockpiles being built as there is limited potential for selectively mining material types directly from the face. The pushbacks design has been relatively successful in meeting the constraints of blending whilst avoiding excessive advance stripping.


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Table 8.2 Total material movement from the pit

Year

Material 2014 2015 2016 2017 2018 2019 2020 2021 2022 Grand Total

Waste

Tonnage 2,701,391 5,778,871 4,667,225 4,744,938 4,054,767 4,318,624 3,630,510 2,781,227 1,560,147 34,237,700

ALP

Tonnage 279,097 264,911 179,312 843,306 403,346 83,233 282,859 130,410 124,438 2,590,913

In Situ Au 854,133 899,239 510,345 3,154,323 1,322,934 250,462 769,325 399,669 353,893 8,514,322

In situ CU 1,200 1,424 468 3,197 2,178 211 905 587 538 10,707

In Situ Ag 4,202,119 3,820,921 2,264,575 20,648,806 4,315,066 1,324,493 6,074,408 2,760,419 1,342,331 46,753,138

Rec. Au 683,307 719,391 408,276 2,523,458 1,058,347 200,369 615,460 319,735 283,114 6,811,457

Rec. Cu 360 427 140 959 653 63 271 176 161 3,212

Rec. Ag 294,148 267,464 158,520 1,445,416 302,055 92,714 425,209 193,229 93,963 3,272,720

Flotation

Tonnage 332,304 910,139 1,064,140 601,947 559,458 1,343,811 1,426,665 1,818,850 1,044,920 9,102,234

In Situ Au 555,010 1,143,743 406,001 1,030,459 1,537,549 554,338 597,641 720,056 489,326 7,034,122

In situ CU 2,581 8,456 9,117 4,835 6,427 9,288 11,957 14,413 8,647 75,722

In Situ Ag 4,248,837 8,008,423 3,518,308 7,877,242 7,236,384 6,443,396 5,845,013 7,035,643 4,139,465 54,352,711

Rec. Au 388,507 800,620 284,200 721,322 1,076,284 388,037 418,348 504,039 342,528 4,923,886

Rec. Cu 2,323 7,610 8,205 4,351 5,784 8,360 10,761 12,972 7,782 68,149

Rec. Ag 3,399,069 6,406,738 2,814,647 6,301,793 5,789,107 5,154,717 4,676,010 5,628,514 3,311,572 43,482,169

Heap

Tonnage 703,696 600,965 976,433 1,290,750 1,007,015 1,415,573 1,130,275 892,776 783,257 8,800,740

In Situ Au 533,079 355,654 631,739 951,738 720,671 759,328 679,210 479,451 545,119 5,655,989

In situ CU 1,234 1,318 1,610 2,284 1,564 2,705 2,154 1,837 1,052 15,758

In Situ Ag 5,287,507 2,682,146 4,682,441 8,929,227 4,089,720 9,886,845 6,646,006 4,481,971 2,902,368 49,588,232

Rec. Au 373,155 248,958 442,218 666,217 504,470 531,529 475,447 335,616 381,583 3,959,192

Rec. Cu 370 396 483 685 469 811 646 551 315 4,727

Rec. Ag 370,126 187,750 327,771 625,046 286,280 692,079 465,220 313,738 203,166 3,471,176

Total Rock Tonnage 4,016,487 7,554,886 6,887,111 7,480,941 6,024,587 7,161,242 6,470,309 5,623,263 3,512,763 54,731,588

Total Ore Tonnage 1,315,096 1,776,015 2,219,885 2,736,003 1,969,820 2,842,618 2,839,799 2,842,036 1,952,616 20,493,888

Total In Situ Au 1,942,222 2,398,636 1,548,085 5,136,520 3,581,153 1,564,127 2,046,175 1,599,177 1,388,338 21,204,433

Total In situ CU 5,015 11,198 11,196 10,315 10,170 12,204 15,015 16,837 10,237 102,187

Total In Situ Ag 13,738,463 14,511,490 10,465,324 37,455,275 15,641,170 17,654,734 18,565,427 14,278,033 8,384,164 150,694,080

Total Rec. Au 1,444,969 1,768,969 1,134,694 3,910,996 2,639,101 1,119,935 1,509,255 1,159,391 1,007,226 15,694,535

Total Rec. Cu 2,795 7,587 7,917 5,512 6,264 8,305 10,483 12,258 7,395 76,089

Total Rec. Ag 4,488,227 7,662,795 3,652,769 9,159,980 7,101,081 6,583,850 6,150,940 6,839,046 4,022,648 50,226,064




Figure 8.1 Comparison of in-situ grades by material type from the pit

-0.40%

-0.20%

0.00%

0.20%

0.40%

0.60%

0.80%

1.00%

1.20%

-

1.00

2.00

3.00

4.00

5.00

6.00

7.00

8.00

2014 2015 2016 2017 2018 2019 2020 2021 2022

In S

itu

Co

pp

er

Gra

de

(%

Cu

)

In S

itu

Go

ld G

rad

e (

g/t

Au

)

Au-ALP AU-HL Au-Float

Cu-ALP Cu-HL Cu-Float


© 2013 CAE Mining

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9 Conclusions and Recommendations

It is concluded that the total Reserve for the Gedabek open pit contains 20.5 Mt of ore with 682,000 oz of gold, 4,885,000 oz) of silver and 102,000 t of copper. The average in-situ (contained) grades are 1.03 g/t Au, 7.35 g/t Ag and 0.50% Cu.

This reserve estimate, when compared to the previous estimate of May 2012 (20.31 Mt @ 1.14 g/t Au, 9.46 g/t Ag and 0.29% Cu), indicates an increase of approximately 3.9 Mt, when allowing for the depletion between 31 December 2011 and 1 September 2104.

This variance can be accounted for by the changes in:

Metal prices for gold, silver and copper

Mining and processing costs

Process recovery

Fixed costs (e.g. G&A)

Cut-off grade (0.3 g/t Au used in 2012)

Process allocation by lithology (oxide/transition or sulphide)

OSA for the pit.

It impossible to apply a direct correlation between the 2012 and 2014 ore Reserve estimates, and this has not been attempted other than to indicate an overall increase in mine life with the revised assumptions. It is however evident that the average copper grade has significantly increased from 0.29% Cu to 0.50% Cu, which suggests that the optimisation of the pit limit, in conjunction with the optimisation of the cut-off grades, has captured high grade copper ore that was previously missed.

The selected pit was defined at a Price Factor of 86%, which was selected on the basis of maximising NPV. There is potential to expand the pit beyond the selected pit limit (Pit Shell 37) but this would involve a significant increase in the waste stripping ratio. It has been determined that this would be uneconomic at today’s prices due to the time taken to expose the additional ore, plus it would put severe pressure on the operation due to the need to double the fleet capacity in the near term.

The increased fleet capacity would probably require a change to the selected equipment as the number of required units of the current size would be prohibitive for the space available. Finding a contractor with larger equipment could also be a problem.

By careful design of the pushbacks it has been possible to schedule the mine so as to accommodate the transition from primarily a gold producer to a gold, silver and copper producer. The key factors here are the date for commissioning of the Flotation plant and the dwindling reserves of material suitable for leaching.




Whilst the reserve estimate presented here is limited to the open pit, it is recognised that there is the potential for an underground mine at Ghadir that could supplement the feed to the plants. This has not been taken into account in the schedule and may change the strategy should this resource be proven.

As regards the open pit, CAE Mining recommends that:

Reconciliation studies are undertaken to improve the model

Infill drilling over several benches is used to improve grade control

Slopes are monitored to give advance warning of potential failure

Recommendations from the environmental audit are expedited

Detailed scheduling is undertaken to:

o Refine the mining sequence

o Avoid grade excursions where possible

o Optimise the usage of the plants

o Establish cycle times and haul truck requirements

o Optimise the waste dumping strategy.

The tactical scheduling proposed in CAE Mining’s proposal offers the opportunity to provide a detailed schedule that can be implemented so as to maximise project value and reduce the risks associated with ore body knowledge.

There will also be opportunities to improve the schedule as more information is gathered with respect to the process configuration.


© 2013 CAE Mining

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10 References

CAE Mining, May 2014, Updated Mineral Resources, Gedabek Mineral Deposit, Republic of Azerbaijan, Azerbaijan International Mining Company Limited.

The Joint Ore Reserves Committee, 2012, The JORC Code, Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves, The Australasian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia, 2012 Edition, 44 p.

Comparisons of production data with the Resource Model and adjustments to improve the reconciliation process at Gedabek Mine. Technical Memo issued by SRK, April 4 2014, 14pp.

Pit slope Stability, Technical report issued by CQA International Limited, Ref 20280, 2 August 2014, 2pp.

Appendix 8, Pre-Feasibility Study, Open Pit Slope Designs, Heap Leaching Facilities, Mine Infrastructure. Knight Piesold Pty Limited, July 2007, 134 pp.

Gadebey Mine – Southern Pit Slope – Geotechnical Appraisal, CQA Consultants Limited, August 2013, 59pp.

Flotation testing of Gedabek Copper-Gold ore, Optimet Laboratories, South Australia, Optimet Report P0184, 8 February 2007, 19pp.

An investigation into the recovery of gold from the Gedabek deposit, SGS Lakefield Research Limited, Project 11367-002 Report 1, 31 May 2007, 152pp.

Scouting Leachbox and Flotation test work on Gedabek sulphide ore samples, Maelgwyn Mineral Services Africa, Report No 11/10, 2 December 2011, 27pp.

Closure and Rehabilitation Management Plan, AMEC Earth and Environmental UK Ltd, December 2012, 62pp.

Environmental Materiality Assessment, Daniel Limpitlaw on behalf of Datamine, 7 September 2014, 23pp.


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11 Compliance Statement

The information in this report that relates to the Mineral Reserve estimate is based on information compiled by Matthew Randall, a Competent Person who is a Professional Member of The Institute of Materials, Minerals and Mining (MIMMM), which is a recognised by JORC as a professional body.

Matthew Randall is employed by Axe Valley Mining Consultants Ltd and is an associate of CAE Mining.

Matthew Randall has sufficient experience that is relevant to the style of mineralisation and type of deposit under consideration and to the activity being undertaking to qualify as a Competent Person as defined in the 2012 Edition of the Australasian Code for Reporting of Ore Reserves.

Matthew Randall consents to the inclusion in a report of the matters based on this information in the form and context in which it appears.

Matthew Randall

Competent Person

Director – Axe Valley Mining Consultants Ltd

Registered Chartered Engineer 345134

Axe Valley Mining Consultants Ltd

47 West St

Axbridege

Somerset, BS26 2AA

United Kingdom


© 2013 CAE Mining

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9 Fricker Road, Illovo

Sandton, South Africa, 2128 Tel: +27/0 11 772 7960 Fax: +27/0 11 268 6229

www.cae.com/caemining

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