Sec.1 ExecutiveSummary EN 180129 AppdFinal...January 2018 Executive Summary Page 1-4 Figure 1-4: Typical Geological Cross-Section, Line PR2B The products of the hydrothermal alteration

CJSC Bazovye Metally

Kekura Gold Project

Feasibility Study

Section 1

EXECUTIVE SUMMARY

January 2018

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Project No. A9KU

CJSC Bazovye MetallyKekura Gold Project

Feasibility Study

JANUARY 2018 EXECUTIVE SUMMARY

CONTENTS

SECTION 1 EXECUTIVE SUMMARY ................................................................................................. 1-1 1.1 Project Description ........................................................................................................... 1-1 1.2 Geology and Mineral Resources ...................................................................................... 1-2

1.2.1 Summary ............................................................................................................. 1-2 1.2.2 Geology ............................................................................................................... 1-3 1.2.3 Database ............................................................................................................. 1-5 1.2.4 Domain Modelling and Estimation ...................................................................... 1-5 1.2.5 Mineral Resource Classification .......................................................................... 1-5 1.2.6 Geotechnics ........................................................................................................ 1-7

1.3 Mining ............................................................................................................................. 1-13 1.3.1 Open Pit Mining ................................................................................................. 1-13 1.3.2 Underground Mining ......................................................................................... 1-17 1.3.3 Mining Schedule................................................................................................ 1-26 1.3.4 JORC Code, 2012 Edition – Table 1 Report ..................................................... 1-30

1.4 Process Plant ................................................................................................................. 1-57 1.5 Dry Stacked Tailing Storage Facility .............................................................................. 1-58 1.6 Power Supply ................................................................................................................. 1-59 1.7 Onsite Infrastructure....................................................................................................... 1-60 1.8 Offsite Infrastructure....................................................................................................... 1-62 1.9 Project Execution ........................................................................................................... 1-63 1.10 Capital Cost Estimate..................................................................................................... 1-64 1.11 Operating Cost Estimate ................................................................................................ 1-66 1.12 Mine Closure .................................................................................................................. 1-66 1.13 Financial Analysis .......................................................................................................... 1-67 1.14 Risks and Opportunities ................................................................................................. 1-68

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TABLES Table 1-1 Mineral Resources at Kekura Deposit as of December 2017 - Open Pit and

Underground Mining Potential1,2,3,4,5,6 ................................................................................................ 1-6 Table 1-2: Mechanical Properties of Key Geotechnical Domains.............................................................. 1-7 Table 1-3: Hoek-Brown Strength Parameters for Geotechnical Domains ................................................. 1-8 Table 1-4: Average Rock Mass Rating Parameters by Lithological Units ................................................. 1-9 Table 1-5: Limit Equilibrium Slope Stability Estimation Results for Kekura Open Pit .............................. 1-12 Table 1-6: Finite Elements Slope Stability Estimation Results for Kekura Open Pit ............................... 1-12 Table 1-7: Cut-Off Grade Estimation ....................................................................................................... 1-13 Table 1-8: Pushback Volumes ................................................................................................................. 1-15 Table 1-9: Open Pit Mining Fleet ............................................................................................................. 1-16 Table 1-10: Open-Pit Mining Capital Costs .............................................................................................. 1-17 Table 1-11: Equipment Operating Unit Costs .......................................................................................... 1-17 Table 1-12: Indicative Unsupported Spans at Kekura by Q Class.......................................................... 1-19 Table 1-13: Tributary Area Pillar Stress versus Depth at Varying Extraction Ratios .............................. 1-19 Table 1-14: Mine Plan ............................................................................................................................. 1-27 Table 1-15: Ore Reserves for the Kekura Deposit as at 1st January 2018 1, 2, 3 .................................... 1-29 Table 1-16: Sampling Techniques and Data ........................................................................................... 1-30 Table 1-17: Reporting of Exploration Results ......................................................................................... 1-35 Table 1-18: Estimation and Reporting of Mineral Resources ................................................................. 1-38 Table 1-19: Estimation and Reporting of Ore Reserves ......................................................................... 1-47 Table 1-20: Major Process Equipment ..................................................................................................... 1-57 Table 1-21: Project Schedule Major Milestones ...................................................................................... 1-64 Table 1-22: Capital Cost Summary .......................................................................................................... 1-65 Table 1-23: Estimated Operating Costs by Category .............................................................................. 1-66 Table 1-24: Financial Summary ............................................................................................................... 1-67 Table 1-25: NPV Sensitivity to Gold Price and Discount Rate ................................................................. 1-68

FIGURES Figure 1-1: Project Location ....................................................................................................................... 1-1 Figure 1-2: Kekura Project Annual Plant Feed and Ore Grade ................................................................. 1-2 Figure 1-3: General View of Kekura Deposit with Open Pit and Underground Workings ....................... 1-3 Figure 1-4: Typical Geological Cross-Section, Line PR2B ...................................................................... 1-4 Figure 1-5: View of Final Structural Model of Kekura Deposit with Cross-Section .................................... 1-9 Figure 1-6: Pit Domaining Based on Stable Bench Parameters .............................................................. 1-11 Figure 1-7: Example of Finite Element Model for Kekura Open Pit, Section S01 Maximum Shear

Strain ................................................................................................................................................ 1-12 Figure 1-8: Pit Shell and Metal Price Sensitivity Diagram ................................................................ 1-14 Figure 1-9 Final Pit and Dump Design and Layout .................................................................................. 1-15 Figure 1-10: Underground Access Scheme for Kekura Deposit .............................................................. 1-21 Figure 1-11: Capital Development Schedule ........................................................................................... 1-21 Figure 1-12: Stope Development Schedule ............................................................................................. 1-22 Figure 1-13: Losses and Dilution During Development of Ore Drifts ....................................................... 1-23 Figure 1-14: Underground Mining Schedule ............................................................................................ 1-24 Figure 1-15: Operating Costs by Year ..................................................................................................... 1-25 Figure 1-16: Operating Costs by Category .............................................................................................. 1-25 Figure 1-17: Operating Costs by Year ..................................................................................................... 1-26 Figure 1-18: Simplified Overall Process Diagram .................................................................................... 1-57 Figure 1-19: Process Plant General Arrangement ................................................................................... 1-58 Figure 1-20: Dry Stack Tailings Storage Location ................................................................................... 1-59

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Figure 1-21: General Layout of the Chaun-Bilibino Power System Projected for 2021-2025 ................. 1-60 Figure 1-22: Project Site Layout .............................................................................................................. 1-62 Figure 1-23: Location of Port and Road Map ........................................................................................... 1-62 Figure 1-24: Overall CJSC Bazovye Metally Functional Structure .......................................................... 1-63

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January 2018 Executive Summary Page 1-1

SECTION 1 EXECUTIVE SUMMARY

1.1 Project Description

The Kekura deposit is located in northeastern Siberia, Russia, in the Bilibino administration district of the Chukotka region. The project site is 195 km southeast from the district centre of Bilibino, and 550 km west from the regional capital of Anadyr.

Figure 1-1: Project Location

The Kekura deposit is covered by the Licence Agreement No. 14974, which is wholly held by CJSC “Bazovye Metally”. In April 2013, limited liability company (LLC) Highland Gold

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acquired 100% of CJSC “Bazovye Metally” and thus acquired the mining and exploration rights to the Kekura deposits and licence area.

Kekura Gold project includes the initial development of an open pit mine using a conventional shovel and haul truck operation to mine 5.748 Mt of ore over the ten (10) year life of the open pit mine. In Year 5, the underground mine is developed to provide additional 3.126 Mt of ore for twelve (12) years of underground mining.

The Kekura process plant is designed to treat 800,000 tonnes of a gold ore annually, or 2,192 t/d, containing 3.7 to 13.5 g/t of gold with an average head grade of 7.49 g/t for the first nine (9) years of plant operation. For the final seven (7) years of operations, the plant will process 300,000 tonnes of ore per year with an average head grade of 5.58 g/t.

Figure 1-2: Kekura Project Annual Plant Feed and Ore Grade

1.2 Geology and Mineral Resources

1.2.1 Summary

The resource estimate was prepared using data collected from geological exploration in the period from 2006 to 2017, including exploration drilling and pilot mining (Figure 1-3). The figure below shows the location of drill holes and trenches, the mineralisation and the open pit layout and underground workings. The mineralization is shown in red for the main zones and in green for minor lenses.

13.5

8.7 7.8

9.2

7.0 7.3 6.1 5.9

3.7

6.3 6.2 6.3 6.8 5.9

4.1 3.3

‐ ‐

2.0

4.0

6.0

8.0

10.0

12.0

14.0

16.0

‐

100

200

300

400

500

600

700

800

900

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17

g/t Au

Thousand tonnes per an

num

Open Pit U/G Grade g/t

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Figure 1-3: General View of Kekura Deposit with Open Pit and Underground Workings

1.2.2 Geology

The Kekura deposit is hosted by a granodiorite intrusion. Mineralization is interpreted to have a paragenetic relationship to the intrusive rocks: residual fluids became progressively enriched in gold as other components crystallised.

Gold mineralization is concentrated around a series of faults, dipping 15-40°, that control the development of zones of hydrothermally altered rocks (beresites) and quartz veins. The faults appear to be dominantly of a thrust nature, with a cumulative displacement across the mineralised package in the order of tens of metres, as indicated by offset of lamprophyre dykes.

The main package of mineralised structures, known as the Pologay zone, has a strike length of over 1,000m, a down-dip extent of at least 400 m, and a thickness of 80 to 120 m.

The Pologay zone outcrops in the southwest. Steeper faults (50 to 80°, east to southeast dipping) control the extent of mineralisation to the west (Vintovoy Fault) and north, and also provide a structure (Daykovy Fault) that approximately corresponds to the likely boundary between the open pit and underground components of the deposit. The constraining structure to the north is itself mineralised (Krutaya Zone). Many of the steeper faults are intruded by diorite porphyry dykes.

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Figure 1-4: Typical Geological Cross-Section, Line PR2B

The products of the hydrothermal alteration associated with gold mineralization are dominantly quartz, sericite, feldspar, carbonate, chlorite and sulphides. Arsenopyrite is the dominant sulphide mineral (up to 2-3%); molybdenite and chalcopyrite occur as minor components. There is a coarse component to the gold distribution, with grains commonly in the range 0.5 - 1 mm, and occasionally reaching 3 – 5 mm.

An oxidation zone, to a depth of 20 to 30 m from surface, is noted from the occurrence of iron hydroxide mineralised on fracture surfaces, but overall this zone appears to be weakly developed, and does not have a significant effect on the distribution or gold, nor the processing properties of the mineralised rocks.

1.2.2.1 Exploration History

The initial phase of geological study over the territory including Kekura occurred from 1957 to 1962. It was not until the 1990s that surveys at 1:50,000 scale and 1:200,000 scale, including surface mapping, and geochemical and geophysical studies, identified Kekura as one of several highly prospective targets for gold mineralization.

In 2004-2008, “Crystal” LLC completed prospecting and evaluation works within the license area of Koralveemsky ore cluster (150 km2), including drilling and trenching of the Kekura deposit.

The results of the exploration were submitted to the GKZ to define preliminary resources. Following GKZ recommendations, the drilling pattern for resource definition was infilled to a line spacing from 25 m to 40 m, a hole spacing of 20 m along drilling lines. Furthermore,

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the GKZ recommended pilot mining and processing of 150,000 tonnes to quantify likely mining losses, dilution and metallurgical parameters.

In 2017, HGM undertook a Reverse Circulation (RC) drilling program, comprising 459 vertical holes on a 10 m by 10 m pattern, with a cumulative length of 14,615 m, targeting the central part of the deposit. HGM also undertook a program of confirmation drilling to check the continuity of the higher grade mineralization in potential underground mining areas. This diamond core drilling included 22 holes with total length of 4,795 m and maximum depth of 379 m.

1.2.3 Database

Input data for the database was provided by HGM (Client). The database contains information on 1,386 trenches and drill holes, and 98,571 core and channel samples.

SRK reviewed the database for the Kekura deposit, and concluded that the quality and quantity of information is sufficient to support the estimation and classification of Mineral Resources and Ore Reserves according to the definitions and standards of the JORC Code (2012).

1.2.4 Domain Modelling and Estimation

A wireframe model to constrain grade estimation was constructed in Leapfrog GEO software. The boundaries of individual mineralised bodies were modelled using a nominal 1 g/t Au threshold. In total, 64 mineralized domains were interpreted. In addition to these domains, over 200 smaller lenses were modelling, mostly based on single intersections.

A grade-based interpretation of mineralisation limits was necessary, because the visual indicators of mineralisation are not clear enough for a coherent interpretation of mineralisation to be obtained from the geological logging.

The wireframes of the mineralisation interpretation were exported to GEOVIA Surpac software for block model grade estimation. A parent block size of 10m x 10m x 5m was used for the model, with sub-blocking to 0.625m x 0.625m x 0.3125m in order to obtain a high resolution fit to the thin mineralization wireframes.

Samples within the wireframes were composited to 1 m, and block grades were estimated by Ordinary Kriging. Each mineralized domain acted as a hard boundary, ie. only composites from within a domain would influence the blocks coded by that domain. Capping grades were determined separately for each domain, and the influence of capped composites was further limited by distance thresholds: beyond a certain distance (typically 50 m) the capped composite would be set to be unavailable for the block estimate search neighbourhood.

A dry bulk density factor of 2.65 t/m3 was applied to all rock types in the model for the purpose of converting volumes to tonnages.

1.2.5 Mineral Resource Classification

SRK classified Mineral Resources at Kekura deposit in accordance with the JORC Code. Mineral Resource classification in general is based on the quality of the information supporting the estimation, and the spacing of sample points in relation to the interpreted continuity of grade and geological features.In the case of Kekura, SRK’s opinion is that the sample spacing is the limiting factor on confidence in the estimates, rather than concerns about data quality. SRK’s essentially assigned classification as follows:

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“Measured” Mineral Resources for areas covered by RC infill drilling (10 m × 10 m spacing);

“Indicated” Mineral Resources for mineralized zones which have been correlated across multiple section lines, and are covered by a drilling pattern no wider than 40 m x 25 m;

“Inferred” Mineral Resources for inconsistent zones, usually interpreted on a single section line, and areas with wider grid spacing (typically 50 m × 50 m or more).

The Mineral Resource statements as of December 2017 for the Kekura deposit, split into resources considered suitable for open pit and underground mining, are presented in Table 1-1.

Table 1-1 Mineral Resources at Kekura Deposit as of December 2017 - Open Pit and Underground Mining Potential1,2,3,4,5,6

Resource Category Tonnage, MtAverage gold

grade, g/t Gold, t Gold, koz

Open Pit3

Measured 0.55 11.2 6.2 198

Indicated 4.25 10.0 42.4 1362

Inferred 0.08 3.1 0.2 8

Total – Open Pit

M+Ind 4.80 10.1 48.5 1561

Inf 0.08 3.1 0.2 8

Undeground4

Horizontal Orebodies COG = 2.8 g/t

Measured 0.03 8.3 0.3 8

Indicated 3.71 6.5 24.3 780

Inferred 0.03 5.8 0.2 5

Vertical Orebodies COG = 1.7 g/t

Measured - - - -

Indicated 0.76 3.7 2.8 89

Inferred 0.06 2.5 0.1 4

Total - Underground

M+Ind 4.49 6.1 27.3 878

Inf 0.08 3.5 0.3 9

Total Open Pit and Underground Resources

M+Ind 9.3 8.2 75.8 2439

Inf 0.16 3.1 0.5 17

1. Mineral Resources are stated assuming 100% ownership of assets as of December

2017.

2. Mineral Resources are reported in accordance with provisions and guidelines of JORC Code 2012.

3. Zones with underground mining potential are additional to the zones with open pit

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potential

4. Mineral Resources were estimated at cut-off grade of 1.2 g/t and a gold price of 1,500 USD/oz.

5. Mineral Resources were estimated at cut-off grade of 1.7 g/t for Vertical zones and 2.8 g/t for Horizontal zones and at a gold price of 1,500 USD/oz.

6. The Competent Person for the estimation of Mineral Resources is Liubov Egorova, MAusIMM, who has more than 10 years relevant experience.

1.2.6 Geotechnics

The following field geotechnical work was carried out at Kekura deposit under SRK supervision in 2017:

Review of existing geotechnical data

Geotechnical drilling with oriented core

Full geotechnical and structural field logging of geotechnical drill core

Sampling for geotechnical laboratory testing

Geotechnical logging of geological core by photographs and geological logs.

SRK processed and reviewed all the resulting information and used it as the basis for open pit slope stability analysis. The slope stability analysis included the following work:

Review of geotechnical properties with account of all available data (including the data collected in 2017)

Structural geological modelling of the deposit

Analysis of joint orientation

Rock mass rating estimation

Open pit bench and berm design

Slope stability analysis.

1.2.6.1 Review of Geotechnical Properties with Account of All Available Data

The tables below show the key physical and mechanical properties for the key geotechnical domains identified by SRK.

Table 1-2: Mechanical Properties of Key Geotechnical Domains

Code Domain

Compressive Strength, MPa

Tensile Strength, MPa

Young's modulus, GPa

Poisson’s Ratio

Co

un

t

Mea

n

Var

iati

on

, %

Co

un

t

Mea

n

Var

iati

on

, %

Co

un

t

Mea

n

Var

iati

on

Co

un

t

Mea

n

Var

iati

on

, %

MSMT_ABP

Argillic-beresite metasomatite

22 29.5 21 22 3.6 24 - 53.0 - - 0.14 -

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Code Domain

Compressive Strength, MPa

Tensile Strength, MPa

Young's modulus, GPa

Poisson’s Ratio

Co

un

t

Mea

n

Var

iati

on

, %

Co

un

t

Mea

n

Var

iati

on

, %

Co

un

t

Mea

n

Var

iati

on

Co

un

t

Mea

n

Var

iati

on

, %

Qdior Quartz diorite 48 75.9 11 48 7.7 11 - 58.0 - - 0.19 -

LMP Lamprophyre (dykes)

5187.

630 5 18.4 18 5 70.0 25% 5 0.17 22

DrPrph Dioritic porphyrite (dykes)

3157.

547 3 17.9 18 3 63.4 8% 3 0.14 18

Brz_70-90-BRZ_90

Beresite 62 84.1 12 62 7.3 14 55.0

0.15

Brz30

Granodiorite with beresiticalteration (lessthan 30%)

191 97.0 15 191 8.5 24 9 62.6 13% 9 0.16 25

GrDior Granodiorite 108160.

118 108 10.5 67 38 66.0 13% 38 0.15 10

The summary Table 1-3 of Hoek-Brown strength parameters was developed from the collected physical and mechanical properties.

Table 1-3: Hoek-Brown Strength Parameters for Geotechnical Domains

Code Domain Density, g/cm3

Young's Modulus Ei, GPa

SCi, MPa mi

MSMT_ABP Argillic-beresite metasomatite 2.62 53.0 30.1 8

Qdior Quartz diorite 2.69 58.0 76.3 10

LMP Lamprophyre (dykes) 2.74 70.0 181.8 9

DrPrph Dioritic porphyrite (dykes) 2.58 63.4 181.6 7

Brz_70-90 Beresite 2.65 55.0 83.7 6

Brz30 Granodiorite with beresitization (less than 30%)

2.66 62.6 98.1 10

GrDior Granodiorite 2.68 66.0 164.6 12

1.2.6.2 Structural Geological Modelling of the Deposit

SRK developed a structural model based on maps and cross-sections prepared in 2014 as part of the TEO report dated January 1, 2015:

36 geological cross-sections of 1:1000 scale

Geological surface plan of 1:2000 scale.

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SRK also used a geological core logging database of exploration drillholes.

The resulting structural geological model includes all main faults, host granodiorite material, argillitic-beresite metasomatite confined to the fault structures, three units of ore hosting beresite with a difference in quartz content, dykes of diorite porphyrite, and lamprophyre.

Figure 1-5: View of Final Structural Model of Kekura Deposit with Cross-Section

1.2.6.3 Rock Mass Rating Estimation

The rock mass rating was estimated for each identified geotechnical interval. The input parameters were taken from the geotechnical logging database prepared by SRK. SRK determined average geotechnical parameters of all lithological units, which are summarized in the Table 1-4 below.

Table 1-4: Average Rock Mass Rating Parameters by Lithological Units

Geo

tech

nic

al

Do

ma

in

Do

mai

n C

od

e

Inte

rval

Len

gth

, m

Inte

rval

Len

gth

, %

RQ

D

FF

/m

UC

S

RM

RL

J01

RM

RB

89

RM

RL

90(R

QD

+Js

)

RM

RL

90(F

F)

GS

I(20

13)

Q

Quaternary sediments Q4 36.8 0.51% 0 0.0 1 0 0 0 0 0 1.6

Lamprophyre dykes

Lmp 142.4 1.98% 84 5.5 188 61 70 57 59 67 11.7

Dioritic porphyrite

DrPrph 77.3 1.08% 86 4.2 157 59 68 61 57 71 11.5

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Geo

tech

nic

al

Do

mai

n

Do

mai

n C

od

e

Inte

rval

Len

gth

, m

Inte

rval

Len

gth

, %

RQ

D

FF

/m

UC

S

RM

RL

J01

RM

RB

89

RM

RL

90(R

QD

+Js

)

RM

RL

90(F

F)

GS

I(20

13)

Q

Quartz diorite QDior 147.0 2.04% 75 4.2 76 49 58 39 40 62 9.2

Beresite (over 90%)

Brz90 204.7 2.85% 88 3.6 84 58 66 60 54 73 14.0

Granodiorite with beresitic alteration of 70-90%

Brz70 398.4 5.54% 91 3.7 84 56 66 59 52 74 13.0

Granodiorite with beresitic alteration of 30-70%

Brz30 313.7 4.36% 91 3.9 97 50 64 57 51 73 8.5

Granodiorite GrDior 5787.7 80.49% 87 5.1 160 60 70 55 54 70 12.3

Crushed zones

CZ 82.5 1.15% 14 23.1

149 50 52 38 39 32 2.0

Total 7190.4 100%

1.2.6.4 Open Pit Bench and Berm Design

SRK conducted kinematic analysis to understand the most probable bench failure mechanisms based on the identified joint sets in various pit sectors. SRK also used probability estimation methods to determine the optimized parameters of benches and berms by pit sectors.

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D02 – geotechnical domain; H – bench height, m; α – bench slope angle, degrees; В – berm width, m; IRA – inter-ramp slope angle, degrees.

Figure 1-6: Pit Domaining Based on Stable Bench Parameters

1.2.6.4.1 Slope Stability Analysis

Slope stability analysis for overall and inter-ramp scale was conducted by geotechnical modelling using limit equilibrium and finite elements methods. The modelling was conducted without account of groundwater, because the deposit is located in the permafrost zone and the lower boundary of permafrost is not exposed by mining operations. The analysis considered seismic conditions even though the site is not located in a seismic region.

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Figure 1-7: Example of Finite Element Model for Kekura Open Pit, Section S01 Maximum Shear Strain

Table 1-5: Limit Equilibrium Slope Stability Estimation Results for Kekura Open Pit

Limit Equilibrium Slope Stability Estimation Results for Kekura Open Pit (Smallest of GLE/Morgerstern-Price/Bishop Simplified/Spenser)

Cross-section Slope height

(H), m Slope angle

(α), ° Factor of

Safety (FOS)

FoS with Seismicity

S01 119 23 6.7 6.6

S01L 90 41 4.7 4.4

S02 126 30 4.5 4.3

S03 88 51 3.7 3.6

S04 72 46 3.7 3.6

S05 128 40 4.7 4.5

S06 133 44 5.0 4.9

S07 103 36 3.9 3.8

Table 1-6: Finite Elements Slope Stability Estimation Results for Kekura Open Pit

Finite Elements Slope Stability Estimation Results for Kekura Open Pit

Cross-section Slope height

(H), m Slope angle

(α), ° Factor of

Safety (FOS)

FoS with seismicity

S01 119 23 3.2 2.9

S01L 90 41 2.1 2.0

S06 133 44 2.1 2.0

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SRK notes that the analysis indicates a high factor safety which is largely due to the need to incorporate access ramps into the walls. Therefore, SRK considers that the designed open pit will be stable and safe for open pit mining provided that all recommendations presented in this feasibility study report are implemented.

1.3 Mining

1.3.1 Open Pit Mining

1.3.1.1 Introduction

SRK defined the Ore Reserve according to international JORC Code (2012). The Ore Reserve presented in this report forms the basis of the mining schedule for the entire life of the open pit.

1.3.1.2 Modifying Factors

Modifying factors for determination of mineable reserve include ore losses and dilution, and economic factors. No other modifying factors were identified as being relevant.

SRK has assumed that a programme of grade control drilling using RC rigs is used to better define the mineralisation, with accurate blast-hole drilling, blast movement monitoring and GPS guided digging is used to minimise mining losses and dilution. SRK estimated the dilution using the “skin” method. Specifically, the procedure for determining dilution included a number of steps:

The main digging equipment is Komatsu PC 1250 with a bucket width of ~2.5m. The anticipated digging accuracy was defined as 0.4m

Therefore, in order to determine dilution, orebody wireframes were expanded 0,4m in all directions

Expanded wireframes were then split into blocks sized 10m x 10m x 5m (assumed mining block size)

Data from the geological block model was then interpolated into mining blocks

Obtained results were then reconciled with initial data to estimate dilution

Dilution amounted to 22%, mining losses – 3%.

1.3.1.3 Open Pit Optimization

Pit optimisation software was used to define the pit limits and mining sequence. The key parameters are summarised in the following Table 1-7.

Table 1-7: Cut-Off Grade Estimation Parameters Units Value

Production Rate tpa 800 000

Geotechnical Parameters

Hanging Wall Degree 50 Foot Wall Degree 50

Mining Factors Dilution Percent (%) 18.0

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Parameters Units Value

Losses Percent (%) 5.0

Processing Recovery Percent (%) 85.8

Operating Cost Base Mining Cost US$/t 1.10 Transportation Cost US$/t 0.70 Additional Cost of Ore Mining US$/tore 0.70 Cost adjustment Factor for Increased Depth US$/bench 0.10 Reference Elevation for Base Mining Cost Z Elevation 1160 Rehabilitation costs US$/tore 0.35 Process costs US$/tore 32.00 Administration cost US$/tore 6.0

Royalty Percent (%) 6.0

Refining Percent (%) 1

Gold Price (US$/oz) 1 150 (US$/g) 36.97

Discount Rate (%) 10

Cut-Off Grade (g/t Au) 1.60

Analysis of the pit shells indicates that there is a significant step in the stripping after $1000/oz where the material mined would need to increase significantly for little additional value. Consequently SRK used the $1000/oz shell to guide the design of the final pit.

Figure 1-8: Pit Shell and Metal Price Sensitivity Diagram

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The mining sequence was developed as a series of four pushbacks in order to optimize the NPV whilst still providing adequate working area to achieve good operating efficiencies and space to mine the orebodies from the hanging wall to the footwall.

SRK used the software NPV Scheduler was used to guide the production sequence and a maximum vertical advance rate of 70m / year was applied as a constraint. The quantities in each pushback are summarized in the following Table 1-8.

Table 1-8: Pushback Volumes Pushbacks Unit 1 2 3 4 Total

Material Mined Mt 10.8 21.7 33.7 49.3 115.6

Ore Mt 0.7 1.1 1.6 1.7 5.1

Waste Mt 10.1 20.7 32.1 47.7 110.5

Strip Ratio t/t 13.9 19.6 19.9 28.5 21.8

Grade g/t 10.5 8.8 8.1 10.4 9.4

Metal Contained t 7.6 9.2 13.0 17.4 47.3

koz 244 296 418 559 1,518

Metal Recovered t 6.2 7.5 10.6 14.2 38.6

koz 199 241 341 457 1,238

1.3.1.4 Access and Mining Sequence

The Kekura gold deposit is located in hilly terrain. Consequently the plan has been developed to minimise unnecessary uphill or downhill haulage through constructing dumps at a similar elevation and using temporary ramps.

Figure 1-9 Final Pit and Dump Design and Layout

The processing plant is located downhill from the pit crest and therefore once the ore has

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been hauled out of the pit, it will be hauled downhill to the primary crusher. The preceding Figure 1-9 shows the final stages of the open pit and waste dumps, and the haul routes.

SRK expects all material will need to be drilled and blasted. At this stage SRK has not allowed for production dozing of waste into the north-eastern waste dump though notes that there is a limited opportunity for this on some of the upper benches.

Waste stripping will be primarily by diesel hydraulic shovels (with 11.0 m3 bucket) in 7.5 m lifts and loaded into 90 t dump trucks. Ore will be mainly mined by diesel hydraulic backhoes with 6 m3 buckets. These excavators will also mine waste and are planned to be utilised at a rate well below full capacity to support the initiative to minimise mining losses and dilution. Consequently, most of the ore will need to be rehandled via a stockpile using wheel loaders.

Given the high processing costs and also the high gold grades, SRK has assumed that selective mining methods will be used. This will include grade control using RC rigs, accurate blasting methods, blast movement monitoring and GPS guided mining.

The key units in the mining fleet can be summarised as follows:

Table 1-9: Open Pit Mining Fleet Major Equipment Role Units

Komatsu PC 2000 shovel Waste mining; 11.0 m3 bucket 2

Komatsu PC1250 backhoe Ore mining 6.0 m3 bucket 2

Komatsu WA500 front end loader Ore stockpile rehandled, 5.0 m3 bucket 1

Atlas Copco DM45 drill Waste stripping 2

Atlas Copco Smart ROC D55 drill Ore 3

Atlas Copco FlexiROC D65 RC drill RC-drilling 2

Komatsu Bulldozer D375 Waste rock dump 3

Komatsu Bulldozer D275 In-pit operations and ore stockpile 2

Komatsu Bulldozer D155 To support ore mining 1

Dump Truck HD785 (90t) Pit dump truck - waste 8

Dump Truck HD465 (55t) Pit dump truck - Ore 9

The mining schedule assumes gradual ramp-up of the pit production rate starting in 2019 with bench preparation and prestripping using an existing fleet of mining equipment, specifically: two Komatsu articulated 40t. trucks, four 55t CAT773 and three MOaZ 25t rigid-body trucks, four Cat D9R dozers, one РС-300 excavator with 1,6 m3 bucket, one CAT390 excavator with 5 m3 bucket and one CAT 365BL excavator with 3,6 m3 bucket. The waste material will be used to construct the tailing storage facility whilst any ore mined will be stockpiled ahead of the plant commissioning..

The main mining fleet is planned to be commissioned in 2019 and 2020 to provide a total mining capacity of 20 Mtpa. The processing plant is planned to be commissioned in stages with 0.5 Mt processed in 2021. The mining schedule is designed to stockpile low grade material to increase the mill feed grade in the early years improving the NPV.

The pit reaches its design production rate in 2022 and this rate is maintained until 2025, with stripping decreasing. Open pit mining operations cease in 2029.

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1.3.1.5 Capital and Operating Costs

SRK estimated that total open-pit mining capital costs for Kekura deposit will amount to US$53 million, of which US$38.6 million will be required in the first two years to ensure ramp-up to the design production rate.

A general list of equipment was compiled from first principles to estimate mobile mining equipment capital costs. The list also includes required replacements during the open pit mine life. Total capital costs are shown in the following Table 1-12.

Table 1-10: Open-Pit Mining Capital Costs

Parameter Units Total 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029

Capital Costs US$m 54.0 13.3 25.3 0.0 0.0 3.8 8.4

1.5

1.4

0.4 - -

Equipment Capital US$m 38.1 10.6 25.3 0.0 0.0 2.2 -

-

-

- - -

Equipment Replacements US$m 13.3 - 0.0 0.0 0.0 1.6 8.4

1.5

1.4

0.4 - -

Miscellaneous US$m 2.6 2.6 0.0 0.0 0.0 0.0 -

-

-

- - -

SRK estimated operating and capital costs from first principles assuming an owner operated strategy. Detailed monthly and LoM production schedules were developed in order to optimize open-pit mining operations and determine equipment operating requirements. The mining operating costs are presented in the following Table 1-11.

Table 1-11: Equipment Operating Unit Costs Component Units Unit Cost

Operating Costs US$/t 1.45

Labour US$/t 0.27

Maintenance US$/t 0.40

Fuel US$/t 0.53

Lubricants US$/t 0.04

Tires US$/t 0.05

Wear Parts US$/t 0.03

Explosives US$/t 0.12

Sampling US$/t 0.01

Miscellaneous US$/t 0.00

1.3.2 Underground Mining

The underground mining study has been developed at a lower level of study than the open pit as there is more variability which needs to be defined by further infill drilling. This said, SRK considers the study sufficient to declare Ore Reserves.

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1.3.2.1 Mining Method

The main mining methods recommended by SRK are step room and pillar for shallow dipping ore bodies and sub-level caving for steeply dipping ore bodies. A method of stoping with backfill was assessed as an alternative mining method for high grade zones, but insufficient material was identified to make this economically justified.

1.3.2.1.1 Step Room and Pillar Method

The vertical distance between the main haulage levels averages 50 m. The length of a mining area along the strike has been adopted in line with the ore body parameters.

The mining panel comprises:

Haulage drives in waste on the lower and upper levels

Ore sub-level drives

The access decline

Ventilation raises on panel flanks.

Stopes are extracted top down along strike. Barrier pillars are left to support the stopes. The pillar dimensions have been developed based on geotechnical analysis and vary with the dip, thickness and depth of the orebody.

Each stope is mined in the following order:

Connections are driven from a decline to a sub-level from which dead-end rock drifts are developed. Air is supplied by VME-6 auxiliary fans via a 0.6 m ventilation duct.

Once the connection between a rock drift and a flank ventilation raise is completed, stopes are blasted from a ventilation raise to a decline. Ore is broken into the ore drift.

1.3.2.1.2 Sub-Level Caving with Frontal Ore Drawing

A method of sub-level caving with frontal ore drawing is used to mine the north-west area of the vertical ore bodies. A block is mined by retreat breast stoping from one flank to another using sub-levels that are spaced vertically at 20 m and are the length of the ore body on strike. The block width is the horizontal width of the ore body.

A block is developed by drives developed from a spiral decline. Ventilation drifts are developed from the drives and connected with ventilation-manway raises. The block is then prepared for extraction by drill and haulage drifts that are driven from accesses to sub-levels. Drifts are aligned to follow the outline of the ore body in order to minimize the cutting of waste rocks during the development and extraction works.

Ore is broken by rings of upward boreholes drilled from the drill and haulage drift and then mucked by an LHD to the truck loading point.

1.3.2.2 Room Spans

The selection of room/drift spans has been assessed using the Q system and a mining excavation support ratio (ESR) as below:

ESR = 1.6 Travelling ways (strike drives/drifts and ramps)

ESR = 3.0 For short term mine openings: limited man access such as dip stopes and general open stopes

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The Q of the immediate hanging-wall is used as the roof will be formed at the hanging-wall contact. The Q values for the GT logged boreholes at the immediate hanging-wall ranges from Q=3 –15 (Fair) with locally values of Q=1-3 (Poor, possibly at faults/shears). All domains appear similar. The Table 1-12 below, indicates the maximum allowable spans.

Table 1-12: Indicative Unsupported Spans at Kekura by Q Class

Q Class Travelling Ways / Haulages

Strike Drives

Stopes Indicative Q Class %

Poor 3m 6m ~10%

Fair 5.5m 10m 70%

Good 8m 15m ~ 20%

1.3.2.3 Roof Support

General roof support in areas exceeding the maximum unsupported spans away from fault/shear zones will be achieved using untensioned, grouted rock-bolts at a spacing of 1 bolt per 2m2.

Consideration should be given to bolting long term main haulages/ramps with bolting on a 2m spacing for budgetary purposes.

For areas with Q<1 i.e. very poor or worse (e.g. faulted and sheared ground) will require bolting at 1m spacing with meshing or at least 50mm of fibre reinforced shotcrete.

1.3.2.4 Pillar Design

The pillar design is a function of Pillar Strength and Pillar Stress. SRK used Laubscher’s formulae which utilises the Mining Rock Mass Rating and his definition of Design Rock Mass Strength (DRMS).

The intact rock strength is reduced by a series of adjustments to empirically determine the design rock mass strength according to Laubscher to modify for rock mass rating and a mix of stronger and weaker rock. SRK therefore estimated the average design rock mass strength of the pillars using the above approach to be: DRMS = 47.5Mpa

1.3.2.4.1 Pillar Stress

The pillar stress used here is derived using the standard calculation of tributary area theory. For the Kekura underground mine, the ore depths range from 150 to 400m. With an overburden rock density of 2.6 t/m3 and assuming the overburden pressure equals the vertical stress i.e. no higher regional residual vertical stress) SRK estimated the pillar stress ranges as per the Table 1-13 below.

Table 1-13: Tributary Area Pillar Stress versus Depth at Varying Extraction Ratios

Depth (m) Pillar Stress at Extraction ratio (MPa)

65% 75% 80% 85%

150 10.9 15.4 19.1 26.5

200 14.6 20.4 25.5 34.0

250 18.3 25.5 31.9 42.5

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Depth (m) Pillar Stress at Extraction ratio (MPa)

65% 75% 80% 85%

300 21.9 30.6 38.3 51.0

400 29.7 41.6 52.0 69.3

1.3.2.4.2 Pillar Strength

A spreadsheet was developed to determine pillar strength, pillar stress and Pillar Factor of Safety using varying pillar widths, pillar lengths, pillar heights (i.e. ore thicknesses) for a pillar DRMS of 47.5 using the Laubscher method.

It should be noted that the Tributary Area method assumes that all of the vertical overburden stress passes directly to the pillars and there is no load shedding into the abutments at the margins of the excavations. This is appropriate for continuous mine extractions over a large footprint (such as a room and pillar coal mine). At Kekura, however, the areas for economic underground mining are not continuous. The largest of the contiguous underground mining areas is that of Zone 18 West, which has a plan footprint of ~300m maximum on strike and 150m – 220m in the dip direction. In mining terms this is not large and hence a significant amount of overburden stress will be shed into the abutments, in effect reducing the height of the overburden load to less than the depth below the surface.

In addition, the use of backfill (where and if used) effectively further reduces the mine footprint as well as providing support to the pillars within and adjacent to the backfilled voids.

For this study, these reductions in the pillar stresses have been partially taken into account on a conservative basis when determining design pillar dimensions. In all cases, SRK notes that the pillar factors of safety still remain above 1.2 in panel (on the tributary area basis) for stopes where pillar stability is required and over 1.3 in areas prior to backfilling. In areas where subsidence mining on the retreat is planned the tributary area factor of safety is a minimum of 1.0.

1.3.2.5 Access Scheme

Underground reserves are accessed by declines, an adit, cross cuts and rock drifts.

To provide ventilation and emergency exits, ventilation and manway raises (VMR) are driven on the flanks of ore bodies together with block raises that connect main levels.

Reserves in the north-west area are accessed by decline developed from both inside and outside the open pit.

Intake air is supplied from the main ventilation unit via declines, cross-cuts, rock drifts, and VMR to the mining levels and to the extraction blocks. The ventilation scheme is of a flank type and the forced ventilation method is used.

Return air is exhausted via the ventilation level and VMR system, and on the western flank - via decline into the mined-out area of the open pit.

ROM material is transported to surface via decline and adit, with separate declines and adits for mine escape ways. The ore is tipped near the portal and then rehandled into haul trucks for transporting to the process plant.

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Figure 1-10: Underground Access Scheme for Kekura Deposit

The annual capital and stope development schedule is presented in the following ures 1-11 and 1-12 below.

Figure 1-11: Capital Development Schedule

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Figure 1-12: Stope Development Schedule

1.3.2.6 Losses and Dilution Calculation

The following factors have been considered for calculation of ore losses and dilution during mining:

Ore body dip angle;

Ore body thickness;

Mining method;

Equipment used.

Losses and dilution were estimated for each mining unit (extraction block).

Extraction block is a mining area which is developed and prepared for extraction using a single system of development mine workings that are required for implementing one mining method.

1.3.2.6.1 Step Room and Pillar Mining Method

The following types of losses are formed when this mining method is used:

General mine losses - losses in the pillars under bottom and sides of the open pit

Unblasted ore (at contact with host rocks, in triangles under the ore drift – Figure 1-13)

Blasted ore (on cleaned floor after shipment of ore from the face)

During pillar extraction.

There will also be losses arising from errors in the geological model where the actual mineralisation lies outside of the stope limits.

Based on the geotechnical section of the report, given the stable span of the exposed stope surface, SRK expects to leave 25% of reserves in pillars and extract 75% of the mineable resource. Given that the boundaries of the economic zones are irregular, it may be possible to modify the panels and pillars to place the pillars in waste material, to increase the proportion of the mineable resource mined.

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Dilution will be generated:

At contacts with host rocks during development of ore drifts (Figure 1-13)

At contacts with host rocks due to cutting of rocks during stope extraction

During pillar extraction

Losses and dilution were estimated for each mining unit (extraction block).

SRK estimated the mining losses and dilution for each mining block based on its specific geometries. The average losses and dilution for the step room and pillar method were therefore estimated to be:

Losses – 10.5%

Dilution – 19.1%

Losses and dilution for the sub-level caving method are:

Losses – 11,3%

Dilution – 26,5%

SRK assumed that the dilution has a grade of 0.34 g/t.

Figure 1-13: Losses and Dilution During Development of Ore Drifts

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1.3.2.7 Mining Schedule

The underground reserves are mainly located away from the open pit allowing concurrent mining. SRK estimates the underground production rate during the period of concurrent mining of the open pit to be 200,000 t of ore. Later, as open pit production decreases, SRK considers that it is possible to increase the underground production to 300,000 t per annum. The underground production rate is mainly a function of the number of faces that can be accessed which in turn is a function of the development rate and resources allocated.

1.3.2.8 Mine Mechanization

SRK propose that men and materials will be transported by a mobile mine auxiliary machine such as the Normet Multimec SF060 with moveable body accessories (cassettes).

Ore is loaded from the stope by a diesel load-haul machine such as the LH-307 Sandvik and then rehandled into mine trucks such as the TH-320 Sandvik for transportation to the surface.

Drilling in development ends is undertaken drill rigs such as the DD-210 drill rig, whilst a drill

such as the DL-310 drill rig will be used for production drilling in stopes.

Figure 1-14: Underground Mining Schedule

1.3.2.9 Capital Costs

Capital costs for the underground mine are based on the required amount of development works, supplier quotations and historical data. The estimates take into account the cost of equipment purchasing, as required, as well as equipment replacement costs for the life of mine.

SRK estimates the average capital cost to be 14 $/t.

1.3.2.10 Operating Costs

All costs are shown in US dollars. The exchange rate of RUB 60 / USD and costs of

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electrical power, diesel fuel, explosives and delivery of goods to the site were provided by the Client. The operating costs were calculated from first principles based on the mining schedule, equipment operating conditions, manpower and consumption of materials and power. Estimated underground unit cost is equal to 35,4 USD per t of ore,

Figure 1-15: Operating Costs by Year

Figure 1-16 below shows the split of underground mining operating costs split by category whilst the following chart shows the trend of unit underground mining operating costs over the project life.

Figure 1-16: Operating Costs by Category

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Figure 1-17: Operating Costs by Year

1.3.3 Mining Schedule

The mining schedule assumes a gradual ramp-up of open pit production. Mining is scheduled to start in 2019 using the existing equipment to mine 5Mt. The waste material will be used to construct the tailing storage facility and the ore will be stockpiled.

New mining equipment is due to be commissioned during 2019 and 2020 to provide an ultimate mining capacity of 20 Mtpa. The full ore mining rate is reached in 2022.

As the underground reserves are located to the side of the open pit, underground mining can take place at the same time as open pit mining.. The production schedule is therefore structured to start underground production five years after the open pit starts operating (~2025) at a rate of 200,000 tpa of ore. Once the open pit ceases operation, the reserves adjacent to the open pit can be developed which, when combined with the fact that more mining areas have been opened up, allows the production to increase to 300,000 tpa.

When the open pit operation stops, the mill will be fed using material from the stockpiles and from underground. When the annual production rate drops below 800,000 tpa, the underground ore will be stockpiled and the mill operated for part of the year only unless another source of ore is found. This will increase the unit processing and G&A costs, and therefore may require the cut-off grade to be increased from 2,9 g/t to 3,1g/t for horizontal ore bodies and 2.1 to 2.6 g/t for vertical ore bodies. As the Company considers that it is likely that additional ore can be proved up, SRK has not adjusted the underground reserves to reflect this scenario.

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Table 1-14: Mine Plan

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1.3.3.1 Ore Reserve

SRK estimated the ore reserves in accordance with the guidelines of the JORC Code (2012) using SRK’s geological model and is based on measured and indicated resource categories only. The open pit and underground reserve estimates is provided in the following Table 1-15.

Table 1-15: Ore Reserves for the Kekura Deposit as at 1st January 2018 1, 2, 3

Reserves Category Tonnage Grade Metal

(Mt) (g/t) (t) (koz)

Open Pit Mining 4

Proved Reserve 0.65 9.24 6.0 193

Probable Reserve 5.10 7.82 39.9 1,282

Underground Mining 5

Proved Reserve - - - -

Probable Reserve 3.13 5.31 16.6 534

Total Ore Reserve

Proved Reserve 0.65 9.24 6.0 193

Probable Reserve 8.23 6.86 56.5 1,815

1. Ore Reserves are reported in accordance with provisions and guidelines of JORC Code 2012.

2. The Competent Person for the open pit ore reserves is David Pearce, FAusIMM who has 30 years’ experience in open pit mining.

3. The Competent Person for the underground ore reserves is Michael Beare, MIMMM who has more than 20 years’ experience in underground mining

4. Ore reserves estimated for COG 1.6 g/t and a gold price of 1,150 USD/oz

5. Ore reserves estimated for COG 2 g/t for vertical zones and 3 g/t for horizontal zones, and a gold price of 1,150 USD/oz

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1.3.4 JORC Code, 2012 Edition – Table 1 Report

1.3.4.1 JORC Section 1 Sampling Techniques and Data

Criteria in this section apply to all succeeding sections.

Table 1-16: Sampling Techniques and Data

Criteria JORC Code explanation Commentary

Sampling techniques

Nature and quality of sampling (eg cut channels,

random chips, or specific specialised industry

standard measurement tools appropriate to the

minerals under investigation, such as down hole

gamma sondes, or handheld XRF instruments, etc).

These examples should not be taken as limiting the

broad meaning of sampling.

Include reference to measures taken to ensure

sample representivity and the appropriate

calibration of any measurement tools or systems

used.

Aspects of the determination of mineralisation that

are Material to the Public Report.

In cases where ‘industry standard’ work has been

done this would be relatively simple (eg ‘reverse

circulation drilling was used to obtain 1 m samples

from which 3 kg was pulverised to produce a 30 g

charge for fire assay’). In other cases more

explanation may be required, such as where there

is coarse gold that has inherent sampling problems.

Core, channel, slurry and geochemical sampling

was conducted on site. The database used to

estimate the mineral resource contains information

on 1,386 trenches and drill holes, and 98,571 core

and channel samples

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Unusual commodities or mineralisation types (eg

submarine nodules) may warrant disclosure of

detailed information.

Drilling techniques

Drill type (eg core, reverse circulation, open-hole

hammer, rotary air blast, auger, Bangka, sonic, etc)

and details (eg core diameter, triple or standard

tube, depth of diamond tails, face-sampling bit or

other type, whether core is oriented and if so, by

what method, etc).

Core drilling with 48mm diameter core and double

tube. Vertical (for Pologaya zone) and inclined (for

Krutaya zone) drilling.

RC drilling used for close-spaced verification drilling

of a small proportion of the deposit.

Drill sample recovery

Method of recording and assessing core and chip

sample recoveries and results assessed.

Measures taken to maximise sample recovery and

ensure representative nature of the samples.

Whether a relationship exists between sample

recovery and grade and whether sample bias may

have occurred due to preferential loss/gain of

fine/coarse material.

Core recovery was controlled by linear and weight

measurements. Double core barrel, shorter drilling

runs, reduced impact drilling modes.

No correlation was found between core recovery

and gold grades.

Logging Whether core and chip samples have been

geologically and geotechnically logged to a level of

detail to support appropriate Mineral Resource

estimation, mining studies and metallurgical studies.

Whether logging is qualitative or quantitative in

nature. Core (or costean, channel, etc)

photography.

The total length and percentage of the relevant

intersections logged.

All drillholes and trenches were logged from collar

to end, total length of geologically logged intervals

is 139,273.6 m.

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Sub-sampling techniques and sample preparation

If core, whether cut or sawn and whether quarter,

half or all core taken.

If non-core, whether riffled, tube sampled, rotary

split, etc and whether sampled wet or dry.

For all sample types, the nature, quality and

appropriateness of the sample preparation

technique.

Quality control procedures adopted for all sub-

sampling stages to maximise representivity of

samples.

Measures taken to ensure that the sampling is

representative of the in situ material collected,

including for instance results for field

duplicate/second-half sampling.

Whether sample sizes are appropriate to the grain

size of the material being sampled.

Core was cut into two halves. One half was used a

sample, and the other was put back into the core

box for storage as a duplicate, or in some cases

was used for metallurgical or other sampling.

The sample preparation flowsheet was developed

using the Richards-Chechett equation, and the

accepted gold distribution factor was equal to 1,

which corresponds to extremely irregular

distribution of gold.

Quality control was regular by screen size analysis

of prepared samples and by cleaning the crushing

and grinding equipment with compressed air.

Quality of assay data and laboratory tests

The nature, quality and appropriateness of the

assaying and laboratory procedures used and

whether the technique is considered partial or total.

For geophysical tools, spectrometers, handheld

XRF instruments, etc, the parameters used in

determining the analysis including instrument make

and model, reading times, calibrations factors

applied and their derivation, etc.

Nature of quality control procedures adopted (eg

standards, blanks, duplicates, external laboratory

All samples went through screening followed by X-

Ray spectral and assay testing.

Duplicates (internal and external control), standard

and blank samples were used for the QA/QC.

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checks) and whether acceptable levels of accuracy

(ie lack of bias) and precision have been

established.

Verification of sampling and assaying

The verification of significant intersections by either

independent or alternative company personnel.

The use of twinned holes.

Documentation of primary data, data entry

procedures, data verification, data storage (physical

and electronic) protocols.

Discuss any adjustment to assay data.

SRK has seen the core remaining from sampling

during the site visit.

SRK did not re-sample the mineralized intervals.

Drillholes and trench sampling were verified by a

number of verification drillholes and trenches.

Initial logging was conducted in standard field

logging forms, including deviation survey and collar

survey results; all records were them transferred to

a digital database.

No adjustments were made to the assay data.

Location of data points

Accuracy and quality of surveys used to locate drill

holes (collar and down-hole surveys), trenches,

mine workings and other locations used in Mineral

Resource estimation.

Specification of the grid system used.

Quality and adequacy of topographic control.

Collars of drillholes and trenches were located

using Leica electronic total station.

Downhole deviation survey was conducted every

20 m using Mir 36 tool.

The survey was conducted in SK-1942 coordinate

system in the Baltic height system in Gauss–Krüger

projection.

Quality of topographic surveys was ensured by

locating the drillholes and trenches against the state

triangulation network.

Data spacing and distribution

Data spacing for reporting of Exploration Results.

Whether the data spacing and distribution is

sufficient to establish the degree of geological and

Core drilling spacing is typically 25m (line) by 20m

(spacing) for the western part of the deposit, up to

50m line spacing in the central part, and 50 to 100m

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grade continuity appropriate for the Mineral

Resource and Ore Reserve estimation procedure(s)

and classifications applied.

Whether sample compositing has been applied.

lines, with 50m spaced holes, for eastern part of the

deposit.

SRK considers that the data spacing is sufficient to

establish geological and grade continuity to the

extent required for the Mineral Resources reported.

Samples were not composited before preparation

and analysis.

Orientation of data in relation to geological structure

Whether the orientation of sampling achieves

unbiased sampling of possible structures and the

extent to which this is known, considering the

deposit type.

If the relationship between the drilling orientation

and the orientation of key mineralised structures is

considered to have introduced a sampling bias, this

should be assessed and reported if material.

The orientation of drilling is generally at a

sufficiently high angle to the orientation of

mineralization for there to be no material risk of

introducing a sampling bias.

Sample security The measures taken to ensure sample security. Sample duplicates are stored in purpose-built

facilities in the accommodation camp of Kekura

Project.

Audits or reviews The results of any audits or reviews of sampling

techniques and data.

SRK’s review of all sampling types has shown no

significant errors in the sampling results;

methodologies of sampling, sample preparation and

testwork are in line with industry standards and are

followed on site; in addition, quality control of

laboratory testwork is conducted continually

throughout exploration campaign.

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1.3.4.2 Section 2 Reporting of Exploration Results

Criteria listed in the preceding section also apply to this section.

Table 1-17: Reporting of Exploration Results Criteria JORC Code explanation Commentary Mineral tenement and land tenure status

Type, reference name/number, location and

ownership including agreements or material issues

with third parties such as joint ventures,

partnerships, overriding royalties, native title

interests, historical sites, wilderness or national park

and environmental settings.

The security of the tenure held at the time of

reporting along with any known impediments to

obtaining a licence to operate in the area.

CJSC Bazoviye Metally (part of Highland Gold

Mining (UK)), with management company

Rusdragmet LLC, is the current holder of the subsoil

license AND 14974 BE.

At the time of reporting the Competent Person was

not aware of any limitations associated with the

license conditions.

Exploration done by other parties

Acknowledgment and appraisal of exploration by

other parties.

The area was subject to geological prospecting

work by various state exploration teams in the

second half of the 20th century. Material information

for this Mineral Resource estimation was collected

on site from 2006.

Geology Deposit type, geological setting and style of

mineralisation.

The deposit is confined to the central part of Kekura

intrusive massif. The mineralized zones represent

series of echelon-shaped beresite zones with

mainly north-east dipping at 12-15 to 25-35° and

north-western strike. The mineralization is

represented by carbonate-cericite-chlorite-quartz

formations. Gold is the main economic component,

which is mainly found in free form. Mineralization

has low sulphide content. No oxidation zone was

found, oxidation of ore minerals does not exceed

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Criteria JORC Code explanation Commentary 13% in the shallow part of the deposit.

Drill hole Information

A summary of all information material to the

understanding of the exploration results including a

tabulation of the following information for all Material

drill holes:

o easting and northing of the drill hole collar

o elevation or RL (Reduced Level – elevation

above sea level in metres) of the drill hole collar

o dip and azimuth of the hole

o down hole length and interception depth

o hole length.

If the exclusion of this information is justified on the

basis that the information is not Material and this

exclusion does not detract from the understanding

of the report, the Competent Person should clearly

explain why this is the case.

A full database of all drillholes is not included in this

report. The Competent Person’s opinion is that

results for individual intersections are not material,

and characteristics of the Kekura deposit are

adequately represented by the Mineral Resource

estimate.

Data aggregation methods

In reporting Exploration Results, weighting

averaging techniques, maximum and/or minimum

grade truncations (eg cutting of high grades) and

cut-off grades are usually Material and should be

stated.

Where aggregate intercepts incorporate short

lengths of high grade results and longer lengths of

low grade results, the procedure used for such

aggregation should be stated and some typical

examples of such aggregations should be shown in

detail.

The assumptions used for any reporting of metal

Individual intersections are not presented in this

report, therefore not applicable.

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Criteria JORC Code explanation Commentary equivalent values should be clearly stated.

Relationship between mineralisation widths and intercept lengths

These relationships are particularly important in the

reporting of Exploration Results.

If the geometry of the mineralisation with respect to

the drill hole angle is known, its nature should be

reported.

If it is not known and only the down hole lengths are

reported, there should be a clear statement to this

effect (eg ‘down hole length, true width not known’).



Diagrams Appropriate maps and sections (with scales) and

tabulations of intercepts should be included for any

significant discovery being reported These should

include, but not be limited to a plan view of drill hole

collar locations and appropriate sectional views.

Diagrams are presented in the Appendices to this

Report.

Balanced reporting

Where comprehensive reporting of all Exploration

Results is not practicable, representative reporting

of both low and high grades and/or widths should be

practiced to avoid misleading reporting of

Exploration Results.



Other substantive exploration data

Other exploration data, if meaningful and material,

should be reported including (but not limited to):

geological observations; geophysical survey results;

geochemical survey results; bulk samples – size

and method of treatment; metallurgical test results;

bulk density, groundwater, geotechnical and rock

characteristics; potential deleterious or

contaminating substances.

Geological mapping of the Kekura deposit and

surrounding area forms the conceptual framework

for modelling the estimation domains. Otherwise,

there are no other exploration data material to the

Mineral Resource estimation.

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Criteria JORC Code explanation Commentary Further work The nature and scale of planned further work (eg

tests for lateral extensions or depth extensions or

large-scale step-out drilling).

Diagrams clearly highlighting the areas of possible

extensions, including the main geological

interpretations and future drilling areas, provided

this information is not commercially sensitive.

Currently, no further exploration work is planned to

occur before the construction of the mine.

1.3.4.3 JORC Section 3 Estimation and Reporting of Mineral Resources

Criteria listed in section 1, and where relevant in section 2, also apply to this section.

Table 1-18: Estimation and Reporting of Mineral Resources Criteria JORC Code Explanation Commentary Database integrity Measures taken to ensure that data has not been

corrupted by, for example, transcription or keying

errors, between its initial collection and its use for

Mineral Resource estimation purposes.

Data validation procedures used.

SRK has reviewed primary data from several drill

holes, and compared this against the records in

the electronic database. No significant

discrepancies were found.

The drill hole database was validated by

automated checks for internal consistency,

carried out during uploading of the database into

SRK’s modelling software. Extensive visual

checks followed during the course of SRK’s

geological modelling work.

Site visits Comment on any site visits undertaken by the

Competent Person and the outcome of those

visits.

If no site visits have been undertaken indicate

why this is the case.

The Competent Person has not visited site.

SRK and Highland Gold agreed that the benefits

to be obtained from a site visit (in terms of

information that could be gleaned by visual

inspection of core and outcrops) would not justify

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Criteria JORC Code Explanation Commentary the considerable logistical effort required to get to

site. Colleagues of the Competent Person

(geotechnical specialists from SRK) have visited

the site, and inspected the core from the 2004-

2008 and 2012-2013 drilling campaigns.

Geological interpretation

Confidence in (or conversely, the uncertainty of )

the geological interpretation of the mineral

deposit.

Nature of the data used and of any assumptions

made.

The effect, if any, of alternative interpretations on


The use of geology in guiding and controlling


The factors affecting continuity both of grade and

geology.

SRK’s confidence in the geological interpretation

of the deposit is high enough to meet the

Indicated and Measured classifications applied to

most the Mineral Resource estimate. This

confidence has is supported by the results

obtained in 2017 from the program of close-

spaced (10m x 10m) RC drilling, and diamond

core verification holes on the deeper part of the

deposit.

In detail, the mineralized domains are modelled

based on a 1 g/t Au grade threshold. The domain

model was prepared using Leapfrog GEO

software.

64 separate mineralized domains were modelled

by SRK. In some places the correlation of

particular domains between holes is ambiguous.

Furthermore, the precise grade threshold could

be adjusted by up to several increments of 0.1 g/t

Au. Overall, however, SRK expects that most

alternative interpretations, using the same

database, would result in a mineralization model

of similar extent and capturing a similar subset of

the sample grades.

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Criteria JORC Code Explanation Commentary The geological interpretation of the deposit by

Highland Gold, building on earlier studies, forms

the conceptual framework for modelling the

mineralization domain, in particular the shallow

NE-dipping orientation of maximum interpreted

grade continuity is used to guide correlation

between intersections for the major (Pologay)

mineralized domain.

Gold mineralization is concentrated around a

series of faults, dipping 15-40°, that control the

development of zones of hydrothermally altered

rocks (beresites) and quartz veins.

Dimensions The extent and variability of the Mineral

Resource expressed as length (along strike or

otherwise), plan width, and depth below surface

to the upper and lower limits of the Mineral

Resource.

The main package of mineralized structures

(Pologay Zone) has a strike length of 1,000m, a

down dip extent of at least 400m and a thickness

of 80 to 120m. This zone outcrops to the west,

and to the east is intersected by core holes

approximately 350m below the surface. Within

the Pologay Zone, individual mineralized

domains are approximately parallel, 2 to 3m thick

on average, and typically several metres apart.

Estimation and modelling techniques

The nature and appropriateness of the estimation

technique(s) applied and key assumptions,

including treatment of extreme grade values,

domaining, interpolation parameters and

maximum distance of extrapolation from data

points. If a computer assisted estimation method

was chosen include a description of computer

Block grades were estimated by Ordinary Kriging

from 1m composites. Estimation was done using

Surpac Geovia software.

Three passes of kriging were used to inform

blocks with grade estimates. The first pass used

a search ellipsoid with a major axis of 80m;

second pass 160m major axis; third pass 320m

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Criteria JORC Code Explanation Commentary software and parameters used.

The availability of check estimates, previous

estimates and/or mine production records and

whether the Mineral Resource estimate takes

appropriate account of such data.

The assumptions made regarding recovery of by-

products.

Estimation of deleterious elements or other non-

grade variables of economic significance (eg

sulphur for acid mine drainage characterisation).

In the case of block model interpolation, the block

size in relation to the average sample spacing

and the search employed.

Any assumptions behind modelling of selective

mining units.

Any assumptions about correlation between

variables.

Description of how the geological interpretation

was used to control the resource estimates.

Discussion of basis for using or not using grade

cutting or capping.

The process of validation, the checking process

used, the comparison of model data to drill hole

data, and use of reconciliation data if available.

major axis. The length ratios between axes were

3.25 (major:intermediate) and 6 (major:minor).

The minimum number of composites required for

an informed block grade was 6; the maximum

number of composites chosen within the search

neighbourhood was 20.

The anisotropy of the search neighbourhood, and

the anisotropy of the variogram model used for

kriging, both varied locally (“Dynamic

Anisotropy”), according to orientation attributes

stored in the block model, obtained from the

orientation of the wireframed domain model.

Previous estimates were used as comparative

checks on the results of the estimation, in

particular a preliminary estimate made by SRK

earlier in 2017 (before the program of RC drilling

and deep core holes), and the Mineral Resource

estimation prepared for Highland Gold by Wardell

Armstrong International Consultants, effective

date July 16, 2015.

Gold is the only element considered for the

Mineral Resource estimation, no assumptions

have been made regarding recovery of by-

products.

No deleterious elements or other non-grade

variable of economic significance were

estimated.

A parent block size of 10m x 10m x 5m was used

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Criteria JORC Code Explanation Commentary for the model, with sub-blocking to 0.625m x

0.625m x 0.3125m in order to obtain a high

resolution fit to the thin mineralization

wireframes. Grade estimates were made in the

parent block.

The 10m x 10m lateral block size is just under

half the 20m (hole) x 25m (line) spacing that has

been achieved by drilling over much of the

western part of the deposit.

The Ordinary Kriging estimate was not further

processed by applying “recoverable reserves”

geostatistical techniques (such as Uniform

Conditioning), therefore no assumptions were

made at the resource estimation stage

concerning dimensions of selective mining units,

smaller than the parent block size.

Gold is the only element considered for the

Mineral Resource estimation, no assumptions

have been made about correlation between

variables.

The 64 mineralised domains, and the minor

mineralized lense zones, each acted as a hard

boundary domain for estimation, ie. Composites

grades from one domain would not influence

block estimates from another domain.

Outlier high grades were identified based on

examination of the tails of histograms and

cumulative frequency plots, and the highlighted

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Criteria JORC Code Explanation Commentary samples were further examined in 3D. The radius

used to restrict the influence of the highest

grades was verified based on the size of clusters

of the highest grades intersected by the close-

spaced RC infill drilling.

The block model estimates were validated

statistically (principally by swath plots, against

the composite grades), and visually (in Leapfrog

GEO and Geovia Surpac, by comparisons

against the composites, domain wireframes, and

drilling database).

Moisture Whether the tonnages are estimated on a dry

basis or with natural moisture, and the method of

determination of the moisture content.

Tonnages are estimated on a dry basis.

Cut-off parameters The basis of the adopted cut-off grade(s) or

quality parameters applied.

Mineral Resources were reported at the following

cut-off grades:

o 1.2 g/t for the open pit component;

o 1.7 g/t for steeply-dipping

mineralized zones in the

underground component

o 2.8 g/t for shallow-dipping

mineralized zones in the

underground component.

These cut-off grades were calculated by

SRK’s mining engineers, based on assumed

mining and processing parameters and

costs, and a gold price of 1500 USD/oz.

Mining factors or Assumptions made regarding possible mining SRK assumed that the deposit will be suitable for

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Criteria JORC Code Explanation Commentary assumptions methods, minimum mining dimensions and

internal (or, if applicable, external) mining

dilution. It is always necessary as part of the

process of determining reasonable prospects for

eventual economic extraction to consider

potential mining methods, but the assumptions

made regarding mining methods and parameters

when estimating Mineral Resources may not

always be rigorous. Where this is the case, this

should be reported with an explanation of the

basis of the mining assumptions made.

a combination of selective open-pit mining

(bench height 5m), and underground mining

(step room and pillar methods for shallow-dipping

mineralized zones, sublevel caving for steeper

zones).

Mining factors and assumptions are discussed

more fully in Section 4 of this table.

Metallurgical factors or assumptions

The basis for assumptions or predictions

regarding metallurgical amenability. It is always

necessary as part of the process of determining

reasonable prospects for eventual economic

extraction to consider potential metallurgical

methods, but the assumptions regarding

metallurgical treatment processes and

parameters made when reporting Mineral

Resources may not always be rigorous. Where

this is the case, this should be reported with an

explanation of the basis of the metallurgical

assumptions made.

An average recovery, from gravity and leaching

processes, of 85.8%, was assumed for the

purpose of determining the cut-off grades stated

above.

Mineral processing factors and assumptions are

discussed more fully in Section 4 of this table.

Environmental factors or assumptions

Assumptions made regarding possible waste and

process residue disposal options. It is always

necessary as part of the process of determining

reasonable prospects for eventual economic

extraction to consider the potential environmental

For the purposes of preparing the Mineral

Resource estimation, SRK has assumed that

there are no environmental issues that pose a

material risk to “eventual economic extraction”

from the Kekura deposit.

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Criteria JORC Code Explanation Commentary impacts of the mining and processing operation.

While at this stage the determination of potential

environmental impacts, particularly for a

greenfield project, may not always be well

advanced, the status of early consideration of

these potential environmental impacts should be

reported. Where these aspects have not been

considered this should be reported with an

explanation of the environmental assumptions

made.

Environmental factors and assumptions are

discussed more fully in Section 4 of this table.

Bulk density Whether assumed or determined. If assumed,

the basis for the assumptions. If determined, the

method used, whether wet or dry, the frequency

of the measurements, the nature, size and

representativeness of the samples.

The bulk density for bulk material must have

been measured by methods that adequately

account for void spaces (vugs, porosity, etc),

moisture and differences between rock and

alteration zones within the deposit.

Discuss assumptions for bulk density estimates

used in the evaluation process of the different

materials.

A dry bulk density of 2.65 t/m3 was applied to all

rock types in the deposit.

This factor has been shown to be consistent

between density samples taken from two phases

of study by Highland Gold (2004-2008, and 2012-

2013, and also between core samples (521

density measurements) and two small (1.5 m3)

test pits.

The 2.65 t/m3 factor is also a consistent average

density between mineralized beresite and host

granodiorite samples.

Classification The basis for the classification of the Mineral

Resources into varying confidence categories.

Whether appropriate account has been taken of

all relevant factors (ie relative confidence in

tonnage/grade estimations, reliability of input

SRK’s opinion is that sample spacing is the

limiting factor on confidence in the

estimates, rather than concerns about data

quality. SRK assigned classification as

follows:

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Criteria JORC Code Explanation Commentary data, confidence in continuity of geology and

metal values, quality, quantity and distribution of

the data).

Whether the result appropriately reflects the

Competent Person’s view of the deposit.

o Measured for areas covered by the

RC infill drilling (10 m × 10 m

spacing);

o Indicated for mineralized zones

which have been correlated across

multiple section lines, and are

covered by a drilling pattern no

wider than 40 m x 25 m;

o Inferred for inconsistent zones,

usually interpreted on a single

section line, and areas with wider

grid spacing (typically 50 m × 50 m

or more).

SRK considers that the classifications

applied take account of all relevant factors,

and the result appropriately reflects the

Competent Person’s view of the deposit.

Audits or reviews The results of any audits or reviews of Mineral

Resource estimates.

SRK’s Mineral Resource estimate has been

reviewed and accepted by technical specialists

on Highland Gold’s staff. No other audits or

reviews of the 2017 estimation have been

undertaken.

Discussion of relative accuracy/ confidence

Where appropriate a statement of the relative

accuracy and confidence level in the Mineral

Resource estimate using an approach or

procedure deemed appropriate by the Competent

Person. For example, the application of statistical

The Competent Person’s view of relative

accuracy and confidence level in the Mineral

Resource estimate is qualitatively represented by

the Measured, Indicated and Inferred

classifications applied.

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Criteria JORC Code Explanation Commentary or geostatistical procedures to quantify the

relative accuracy of the resource within stated

confidence limits, or, if such an approach is not

deemed appropriate, a qualitative discussion of

the factors that could affect the relative accuracy

and confidence of the estimate.

The statement should specify whether it relates

to global or local estimates, and, if local, state the

relevant tonnages, which should be relevant to

technical and economic evaluation.

Documentation should include assumptions

made and the procedures used.

These statements of relative accuracy and

confidence of the estimate should be compared

with production data, where available.

Classifications have been applied locally, to each

block in the block model.

At the time of preparing this Mineral Resource

estimation, mining had not yet commenced, so

no production data were available for

comparison.

1.3.4.4 JORC Section 4 Estimation and Reporting of Ore Reserves

Criteria listed in section 1, and where relevant in sections 2 and 3, also apply to this section.

Table 1-19: Estimation and Reporting of Ore Reserves Criteria JORC Code Explanation Commentary

Mineral Resource estimate for conversion to Ore Reserves

Description of the Mineral Resource estimate used as a

basis for the conversion to an Ore Reserve.

Clear statement as to whether the Mineral Resources

are reported additional to, or inclusive of, the Ore

Reserves.

The Measured and Indicated Resources were used as

the basis of the Ore reserve. The Inferred Resources

were treated as being waste.

Mineral resources are inclusive of Ore Reserves

Site visits Comment on any site visits undertaken by the The Competent Person did not visit the site as it is a

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Criteria JORC Code Explanation Commentary

Competent Person and the outcome of those visits.

If no site visits have been undertaken indicate why this

is the case.

greenfield site. Staff under the CP’s direct

management visited site to undertake geotechnical and

geological studies.

Study status The type and level of study undertaken to enable

Mineral Resources to be converted to Ore Reserves.

The Code requires that a study to at least Pre-

Feasibility Study level has been undertaken to convert

Mineral Resources to Ore Reserves. Such studies will

have been carried out and will have determined a mine

plan that is technically achievable and economically

viable, and that material Modifying Factors have been

considered.

SRK undertook the mining aspects of the Feasibility

Study and Fluor Engineering the processing and

infrastructure aspects.

Cut-off parameters

The basis of the cut-off grade(s) or quality parameters

applied.

The cut-off grades were determined using a gold price

of USD 1,150/oz. The cut-off grade for open pit mining

was 1.6 g/t, for underground mining using step room

and pillar methods 3 g/t and using sub-level caving

methods 2 g/t.

Mining factors or assumptions

The method and assumptions used as reported in the

Pre-Feasibility or Feasibility Study to convert the

Mineral Resource to an Ore Reserve (i.e. either by

application of appropriate factors by optimisation or by

preliminary or detailed design).

The choice, nature and appropriateness of the selected

mining method(s) and other mining parameters

including associated design issues such as pre-strip,

access, etc.

The assumptions made regarding geotechnical

The mining loss and dilution were determined for the

open pit reserves by applying a 0.4m skin to the lenses

and then excluding blocks with block grades below the

cut-off grade. This results in a mining loss of 3% and

dilution of 22%. The dilution grade was assumed to be

0 g/t. The factors assume that good grade control

practices are applied, namely RC drilling ahead of

mining, minimising blast movement, measuring blast

movement using blast movement monitors, and GPDS

guided digging.

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parameters (eg pit slopes, stope sizes, etc), grade

control and pre-production drilling.

The major assumptions made and Mineral Resource

model used for pit and stope optimisation (if

appropriate).

The mining dilution factors used.

The mining recovery factors used.

Any minimum mining widths used.

The manner in which Inferred Mineral Resources are

utilised in mining studies and the sensitivity of the

outcome to their inclusion.

The infrastructure requirements of the selected mining

methods.

The open pit economic limits were defined using pit

optimisation software using a gold price of $1,150/oz, a

fixed gold recovery rate of 85.8%, a base mining cost

of $1.80/t with incremental ore mining cost of $0.70/t

and a depth adjustment factor of $0.10/t/10m,

processing costs of $32/t and a G&A of $6/t for a

production rate of 800,000 tpa. The overall slope

angles of 41-44 degrees were checked using finite

element stability software and found to have a Factor of

Safety of 2.1. This is due to the presence of haul

ramps in the walls.

The mining loss and dilution were determined for the

underground reserves as follows:

Room & Pillar.

o The mining loss includes pillars,

wedges and material left on the floor.

The losses have been determined to

be 10.5% estimated for each mining unit (extraction block)

o The dilution includes material mined in

the drive floors and at contacts. The

dilution was determined to be 19.1% at

a grade of 0.34 g/t.

o The minimum mining width was

assumed to be 1.5m

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o Potentially economic blocks were

identified using the Mine Stope

Optimiser software and isolated stopes

were excluded if they were not located

near a main mining area.

Sub-Level Caving

o The mining loss was estimated to be

11.3% based on estimate for each mining unit (extraction block)The

dilution was estimated to be 26.5%

based on estimated for each mining unit (extraction block)

Metallurgical factors or assumptions

The metallurgical process proposed and the

appropriateness of that process to the style of

mineralisation.

Whether the metallurgical process is well-tested

technology or novel in nature.

The nature, amount and representativeness of

metallurgical test work undertaken, the nature of the

metallurgical domaining applied and the corresponding

metallurgical recovery factors applied.

Any assumptions or allowances made for deleterious

elements.

The existence of any bulk sample or pilot scale test

work and the degree to which such samples are

considered representative of the orebody as a whole.

For minerals that are defined by a specification, has the

The process flowsheet comprises a single stage

crushing followed by SAG mill / ball mill grinding circuit

in closed circuit with Knelson concentrators has been

well tested and modelled using well proven software.

The major share of the gold dore production will be

produced by this single stage gravity stage in

conjunction with Intensive Leaching and direct electro-

winning of the resulting eluate. This circuit effectively

recovers the coarse gravity recoverable gold as

inexpensively as possible.

The remaining gold tends to be very fine grained and

associates itself with Arsenopyrite and to a lesser

extent with Pyrite and non-gangue minerals and this

gold is effectively recovered using cyanide leaching in

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ore reserve estimation been based on the appropriate

mineralogy to meet the specifications?

CIP modus operandi, since there has never been any

evidence of “preg robbing”.

In late 2015, 8 samples were produced by Highland

Gold that represented each two years of production

from the open pit (4) and separately from

Underground (4). No significant difference was

observed in the gold recovery between open pit and

underground.

A clear relationship has been developed between g/t

Au in plant feed and gravity recoverable gold and

overall gold recovery. As a consequence, the financial

modelling has been able to utilise this relationship

rather than adopting a single recovery number for life

of mine.

The variability hardness testing has shown very little

significant variability albeit the samples tested were

mainly annual mine sequence composites. The fact

that the individual rock samples also did not

demonstrate significant variation does suggest that

the hardness variability is not significant.

Environmental The status of studies of potential environmental

impacts of the mining and processing operation. Details

of waste rock characterisation and the consideration of

potential sites, status of design options considered and,

where applicable, the status of approvals for process

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residue storage and waste dumps should be reported.

Infrastructure The existence of appropriate infrastructure: availability

of land for plant development, power, water,

transportation (particularly for bulk commodities),

labour, accommodation; or the ease with which the

infrastructure can be provided, or accessed.

The site is remote with no on-site infrastructure. A

winter road exists to the nearest town of Bilibino.

Materials will be transported during summer to the port

of Pevek on the north coast and then during the winter

to site. A power line is due to be developed to site.

Studies have indicated that there is sufficient power in

the local grid to support production.

The main source of fresh water for the plant is from the

Dva Ozera River in the warm months of the year (May

– September). The river will feed a fresh water lake

with a capacity of 504,500m3 and this fresh water lake

will be topped up in summer months. In winter months,

the fresh water lake will freeze and an allowance has

been made for the water that is locked up with ice.

On-site accommodation will be constructed and there

will be a seasonal roster of 6 months on, 6 months off.

The company has existing operations in Far East

Russia which can provide skilled supervision initially.

Costs The derivation of, or assumptions made, regarding

projected capital costs in the study.

The methodology used to estimate operating costs.

Allowances made for the content of deleterious

elements.

The source of exchange rates used in the study.

Equipment, labour and consumables were estimated

from first principles. Unit prices were supplied by the

Company from a combination of quotes and their

experience in the region.

The exchange rate used is RUB 60 / USD.

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Derivation of transportation charges.

The basis for forecasting or source of treatment and

refining charges, penalties for failure to meet

specification, etc.

The allowances made for royalties payable, both

Government and private.

The financial analysis and cut-off grade analysis

considers royalties of 6%.

Revenue factors The derivation of, or assumptions made regarding

revenue factors including head grade, metal or

commodity price(s) exchange rates, transportation and

treatment charges, penalties, net smelter returns, etc.

The derivation of assumptions made of metal or

commodity price(s), for the principal metals, minerals

and co-products.

A constant gold price of USD 1,150/oz was used in the

financial analysis and is in line with consensus

forecasts. The cost of refining was taken as 1% of gold

revenue. Royalties are 6% for gold. There are no co-

products or penalty elements.

Market assessment

The demand, supply and stock situation for the

particular commodity, consumption trends and factors

likely to affect supply and demand into the future.

A customer and competitor analysis along with the

identification of likely market windows for the product.

Price and volume forecasts and the basis for these

forecasts.

For industrial minerals the customer specification,

testing and acceptance requirements prior to a supply

contract.

N/A

Economic The inputs to the economic analysis to produce the net

present value (NPV) in the study, the source and

confidence of these economic inputs including

estimated inflation, discount rate, etc.

SRK estimated operating and capital costs to assess

unit mining costs on an owner operated basis, whilst

Fluor Engineering estimated the processing costs from

first principles for the production rate of 800,000 tpa.

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NPV ranges and sensitivity to variations in the

significant assumptions and inputs.

The unit processing cost was adjusted using the

6/10ths power rule for processing the underground ore

at 300,000 tpa though this approach was also checked

by MPH. The financial model was a real terms model

with no adjustment for inflation.

The NPV was estimated to be USD 331 using a real

terms discount rate of 10% with an IIR of 31%.

Social The status of agreements with key stakeholders and

matters leading to social licence to operate.

There are no nearby settlements.

Other To the extent relevant, the impact of the following on

the project and/or on the estimation and classification

of the Ore Reserves:

Any identified material naturally occurring risks.

The status of material legal agreements and marketing

arrangements.

The status of governmental agreements and approvals

critical to the viability of the project, such as mineral

tenement status, and government and statutory

approvals. There must be reasonable grounds to

expect that all necessary Government approvals will be

received within the timeframes anticipated in the Pre-

Feasibility or Feasibility study. Highlight and discuss

the materiality of any unresolved matter that is

dependent on a third party on which extraction of the

reserve is contingent.

No other material risks were identified.

Classification The basis for the classification of the Ore Reserves into

varying confidence categories.

The Ore reserves were classified as Proved and

Probable reserves based on the Mineral Resource

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Whether the result appropriately reflects the Competent

Person’s view of the deposit.

The proportion of Probable Ore Reserves that have

been derived from Measured Mineral Resources (if

any).

Classification

Audits or reviews The results of any audits or reviews of Ore Reserve

estimates.

No audits or reviews have been undertaken of the

Feasibility Study as yet.

Discussion of relative accuracy/ confidence

Where appropriate a statement of the relative accuracy

and confidence level in the Ore Reserve estimate using

an approach or procedure deemed appropriate by the

Competent Person. For example, the application of

statistical or geostatistical procedures to quantify the

relative accuracy of the reserve within stated

confidence limits, or, if such an approach is not

deemed appropriate, a qualitative discussion of the

factors which could affect the relative accuracy and

confidence of the estimate.

The statement should specify whether it relates to

global or local estimates, and, if local, state the relevant

tonnages, which should be relevant to technical and

economic evaluation. Documentation should include

assumptions made and the procedures used.

Accuracy and confidence discussions should extend to

specific discussions of any applied Modifying Factors

that may have a material impact on Ore Reserve

viability, or for which there are remaining areas of

uncertainty at the current study stage.

The relative confidence level of the open pit reserves is

higher than for the underground reserves as there has

been more exploration in the open pit area and the

open pit mining method is more flexible in terms of

defining the location of the mineralisation. Given that

the underground reserves are not scheduled to be

mined until Year 5 of operations, SRK believes that the

confidence of the underground reserves can be

increased prior to development of the underground

mine.

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It is recognised that this may not be possible or

appropriate in all circumstances. These statements of

relative accuracy and confidence of the estimate

should be compared with production data, where

available.

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1.4 Process Plant

The plant is designed to treat 800,000 tonnes of gold ore annually.

Figure 1-18: Simplified Overall Process Diagram

The metallurgical processes used in the process facilities are all well proven technologies, as indicated in the simplified flow sheet above in Figure 1-128.

Major process equipment is summarized in the following Table 1-20.

Table 1-20: Major Process Equipment Item Unit Quantity Equipment Size

Primary Crusher mm 1 1,000 X 760

SAG Mill D x L, m 1 6 X 2.9

Ball Mill D x L, m 1 4.2 X 6.5

Pre-Leach Thickener m, diameter 1 20

Pre-Oxidation Tank m3 1 1,150

Leach Tanks m3 6 1,150

CIP Tanks m3 6 40

Gravity Concentrator - 2 KC-QS48

Intensive Cyanidation - 1 CS4000

Electrowinning (Intensive Cyanidation) m3 1 2.12

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Item Unit Quantity Equipment Size

ADR Circuit t 1 3

Electrowinning (CIP) m3 2 3.54

Refinery kW 2 50

Effluent Treatment Tanks m3 2 170

Pressure Filtration (maximum) m2 2 (1

standby) 992

The existing process plant building (84 m long by 42 m wide) will be extended by four bays (24 m) to house the grinding, gravity concentrator, leaching, and the adsorption desorption recovery plant (ADR) including gold room. In addition, there will be a 24 m wide by 48 m long extension to the south to house the tailing filters. The floor arrangement is shown in Figure 1-19 below.

Figure 1-19: Process Plant General Arrangement

The process plant will be constructed in a single phase.

1.5 Dry Stacked Tailing Storage Facility

The leach tailings from the cyanide detoxification circuit are thickened in a single tailings thickener before being pumped to the pressure filtration circuit. The tailings are dewatered to a target moisture content of approximately 15% suitable for dry stacking. The filter press filtrate water is pumped to the process fresh water tank to be recycled and used again throughout the processing plant. The dewatered tailings drop onto the floor of the tailing area building from where a front end loader will load haulage trucks that will for transport the tailings to the lined dry stack pad.

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Figure 1-20: Dry Stack Tailings Storage Location

1.6 Power Supply

Power supply for Kekura Gold project will be from the regional electrical grid. The 110 kV line to Peschanka planned for 2020 will also supply power for the Kekura project. In order to provide connection of PS 110 / 6 kV Kekura substation to the Chukotenergo power grid facilities, a set of measures must be implemented in accordance with the technical specifications issued by the Department of Capital Construction of the Chukotka Autonomous Okrug. Commissioning of the floating nuclear power station in Pevek and delivery to the network of approximately 63 MW by 2020 are in planning. The design and estimate documentation are developed. The work has started on construction of 2 circuits of the Pevek-Bilibino 110 kV overhead line. The projected completion date of the 1st circuit is not earlier than the IV quarter of 2020, and the 2nd circuit not earlier than the IV quarter of 2021. The basic technical solutions are developed for construction of an energy source for approximately 24 MW in Bilibino. The term is tentatively not earlier than the IV quarter of 2020. Thus, power from the branch line to the PS 110/6 kV Kekura substation is to be available not earlier than IV quarter of 2020.

East Abutment Area

West Abutment Area

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Figure 1-21: General Layout of the Chaun-Bilibino Power System Projected for 2021-2025

The power supply to the PS110/6 kV substation is outside battery limits of the project and not included in the capital cost. The PS 110/6 kV substation is included as part of the project scope and capital cost.

Emergency electrical power will be generated on site by modularized diesel generators to provide emergency backup for the life-critical services and essential process facilities as defined by consumers of the 1st and 2nd categories according to the Electrical Installation Code.

1.7 Onsite Infrastructure

The onsite infrastructure area comprises non-process facilities within the boundaries of the plant site and connecting road/service corridors required to support a safe and efficient

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mining operation. These will include operations facilities, maintenance facilities, roads, and site utilities.

The necessary facilities to support the day-to-day plant operations will comprise:

Industrial complex building

Operation camp

Other infrastructure.

The industrial complex building, approximately 114 metres long, 33 metres wide by 18 metres high, will serve as the internal hub for operations and maintenance, and a base for the visitors and vendors.

The operation camp, with a capacity of 670 persons (bed), will be located close to the industrial complex building.

Other infrastructure required to support the mine, but not included within the industrial complex are:

Sewage treatment facilities

Waste oil storage

Warehouse (existing) and storage yard

Container storage

Temporary repair shop

Gate house

Diesel storage

Diesel power station with main 6 kV switchgear

Light vehicle fuel station

Boiler house

Mine truck parking area

Ammonium nitrate storage

Explosives storage

Water treatment plant

Raw/fire/potable water pumphouse

Solid waste landfill.

A dedicated boiler house with boilers, all boiler ancillary equipment, heat exchangers, and pumping system, located adjacent to the diesel power station, will produce heat for all of the plantwide facilities

Process water for the Kekura site will be sourced from the Two Lakes water storage reservoir, which has a capacity of 550,000 m3, adequate to support the plant process requirements in winter months .

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Figure 1-22: Project Site Layout

1.8 Offsite Infrastructure

The main item of project offsite infrastructure is the transportation road. Existing all-weather road, when it is in good condition, will be used to transport the supplies from Pevek to as far as Ilirney during the summer period, and then move them from Ilirney to site on the winter road.

Figure 1-23: Location of Port and Road Map

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Any supplies that are not transported on the summer all-weather road will be transported from Pevek to km 28 intersection on the government built winter road, open from January to mid-April, and further to site on the Kekura access ice-road maintained by HGM.

The port of Pevek is navigable for 100 days a year (early July to late October). Vessels with displacement more than 8,000 tonnes will typically be able to use the port no earlier than 20 to 25 July.

1.9 Project Execution

The project consists of five major areas as follows:

1. Gold Process Facilities – This includes all process facilities starting at the primary crusher, and including the conveyors, coarse ore storage, grinding, leach, CIP, gold recovery, reagents, tailing filtration, transport of dry tailings to the tailing storage facility, and all support utilities.

2. Onsite Infrastructure – This includes site preparation; water supply; PS 110/6 kV Kekura substation; emergency diesel generators; boiler house and heating system; warehouse; fuel farm; communication systems; and industrial complex containing mine maintenance complex, administration, laboratory, plant maintenance, and workshops.

3. Mine Area – The Mine area includes the surface and underground mine development, mine pre-strip and capital mining operations, mine dewatering, and in-pit and mine infrastructure electrical distribution .

4. Tailing Storage Facility (TSF) Area – The TSF area includes the dry stack tailing pad, seepage control system, and all civil construction of the TSF.

5. Offsite Infrastructure – This includes winter access road from Bilibino to mine site and other required offsite infrastructure.

Figure 1-24: Overall CJSC Bazovye Metally Functional Structure

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Owner will manage the project execution. Russian Design Institutes (RDIs) will be contracted by Owner to complete the engineering and Design Documentation. Owner will assemble a project management team within a standard functional organization structure to manage the engineering and procurement contracts, construction contracts, and coordination of services to complete all project scope.

A project execution schedule was developed during the feasibility study. Key elements of the schedule are developed to level 2 detail, with some minor sections at level 1.The schedule considers the impact of the seasonal shipping window to Pevek, and the seasonal transportation on the winter road from Pevek to the project site. The project schedule major milestones are listed in Table 1-21 below:

Table 1-21: Project Schedule Major Milestones Activity ID Activity Name Date

M0009 Issue FS Report - Final 24 Jan 2018

M0015 Investment Decision 01 Mar 2018

M0800 Start Basic Engineering Phase - By RDI (Continue From FS Phase) 02 Mar 2018

M0016 Award POs for Critical Major Equipment - Release for Manufacture - Delivery to Site in Q1, 2019 02 Mar 2018

M3010 Complete Erection of Process Building Steel Shell 18 Jul 2018

M3020 Complete Installation of Mill Piers 31 Oct 2018

M0010 PO 6-0001 - Generators - Award & Release for Manufacture (Delivery to Site in Q1, 2020)

3 Dec 2018

M0019 Award POs for Major Equipment - Release for Manufacture (2019) - Delivery to Site in Q1, 2020

15 Jan 2019

M0900 Complete Rough-Set of Mills 30 Jun 2019

M0300 Complete Construction (Excluding Pre-Commissioning) 10 Nov 2020

M0400 Complete Pre-Commissioning - Mechanical Completion 27 Jan 2021

M0001 First Gold 2 Jun 2021

M0000 Complete Ramp-Up to Full Production 4 Aug 2021

Maintaining the location and arrangement of the facilities, the critical path of the schedule includes commitment to expedited vendor data and key equipment delivery, and fast track engineering and construction for the 2018 site works. Purchase orders for critical major equipment will be prepared by HGM procurement in advance so as to award immediately after positive investment decision.

The start-up of the onsite laboratory by SGS needs to be planned to provide necessary services with sample preparation on site for geological grade control and ship “pulps” to SGS Chita for analysis and then build the laboratory later to start up just before the plant start-up.

1.10 Capital Cost Estimate

The class 3 capital cost or capital expenditure (CAPEX) estimate and sustaining capital expenditure estimate were prepared for Highland Gold Mining Company (HGM) during the Kekura Gold project feasibility study.

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The capital cost estimate provides for the supply and installation of material and equipment using contract labour, including provisions for engineering, construction management services, and pre-commissioning costs for the following scope:

Mining operation to be an open pit and underground shaft located at the Kekura site, including mine fleet, mine maintenance facilities, and explosive storage and manufacture

The Kekura ore processing facilities to be a conventional processing plant, including crushing, crushed ore stockpile, grinding, gravity concentration, intensive cyanidation, leaching, CIP, electrowinning and refining, which will produce gold doré and tailings

All required infrastructure including operations camp, administration buildings, warehouses, roads, water supply

Dry stack tailing storage facility and reclaim water system

An access ice-road to be built by the project to connect the processing facilities at Kekura and km 28 intersection.

Key offsite infrastructure required for the project, but not included in the capital cost estimate are:

Facilities at Pevek port to service the mining operation with import of fuel, consumables, and equipment

Remainder of the winter road from km 28 intersection to Pevek port used for transporting supplies and fuel to the Kekura site to support the mining operation will be maintained by the state.

The initial capital estimate is summarized in Table 1-22 below:

Table 1-22: Capital Cost Summary

Mining Infrastructure Process Plant Total

(M USD)

Direct 45.54 55.01 69.57 170.12

Indirect 15.63 22.83 38.46

Sub-total 45.54 70.64 92.40 208.58

Contingency @ P65 6.82 8.92 15.73

Sub-total 45.54 77.46 101.31 224.32Owner’s Cost 4.80

Project Total 229.12

Contingency/Direct+Indirect 9.65% 9.65% 7.5%

Project contingency @ P65 was chosen to be used in the capital cost estimate.

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1.11 Operating Cost Estimate

The purpose of the operating cost estimate is to determine the most likely cost to use as part of the project economic analysis, and to be used in the financial model to determine project viability.

The operating cost estimate is based on the inputs from the following:

Mining OPEX including both open pit and underground is by SRK.

OPEX for process plant and tailings is by Fluor, calculated based on a typical 800,000 tonne throughput. Six tenths power rule is used in the financial model to adjust the OPEX for those years when the throughput is less than 800 ktpa.

OPEX for infrastructure, G&A and logistic is provided by Highland Gold based on in-house data for similar mines.

The estimated total life of mine operating costs and unit rates by category are summarized in the Table 1-23 below.

Table 1-23: Estimated Operating Costs by Category

Category LOM Cost

(US$M) Unit Rate

Open Pit (incl. overburden) 185.14 1.45 /t rock mined Underground 110.49 35.34 /t ore mined

Process Plant 224.21 25.26 /t ore milled

Tailings and Water Management 13.55 1.53 /t ore milled

Infrastructure 93.10 10.49 /t ore milled

G&A incl. logistic 82.59 9.31 /t ore milled Plant Total 709.08 79.90 /t ore milled

1.12 Mine Closure

Conceptual closure activities included in this report represent a typical approach based on available information and industry practice.

Decommissioning and closure activities will take several months. During the final closure phase, the waste storage facilities will be reclaimed and infrastructure not required in the post-closure period will be decommissioned, demolished or removed from site. Once all the applicable standards and requirements have been met, a site reclamation report will be issued and signed by qualified professionals and government agencies, at which point the reclamation will be considered complete.

The mine closure cost estimate at this stage of development is estimated at 4% of direct costs, which equates to US$ 5.14 million.

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1.13 Financial Analysis

Economic analysis was conducted to assess the economic viability of constructing and operating the Kekura Gold project as designed. This analysis was evaluated using a real (non-escalated), after-tax discounted cash flow (DCF) model on a 100% project equity financing (unlevered).

The economic evaluation presents the after tax net present value (NPV), payback period, and the after tax internal rate of return (IRR) for the project based on annual cash flow projections. End-of-year discounting is applied in the analysis. NPV valuation date is January 2018.

A marketing study was not undertaken. The gold price of $1,250 per ounce provided by Highland Gold used in the financial analysis is consistent with median broker consensus.

The results of the Discounted Cash Flow (DCF) analysis show an NPV of US$ 311 million at 10% discount rate, 38.1% IRR and 5.2 years discounted-payback period.

Table 1-24: Financial Summary Rock Mined- Open Pit Kt 127,886

Rock Mined – Underground (U/G) Kt 3,126 Total Mined Kt 131,013 Ore Processed - Open Pit Kt 5,748 Ore Processed – Underground (U/G) Kt 3,126 Total Processed Kt 8,874 Strip Ratio for OP 21.2

Gold in Ore Koz 2,010 Gold Recovered Koz 1,744 Gold Recovery % 86.8% Gold Price US$/oz 1,250 Value of gold in Dore US$M 2,180

Total Operating Cost* US$M 870 Total Cost Before Taxes US$M 1019 Total Cost Including taxes US$M 1236

EBITDA US$M 1,287 Net Cash Flow (undiscounted) US$M 798 Jan 2018 NPV@10% US$M 311 Jan 2018 [email protected]% US$M 334 2018 IRR % 38.1% Payback Period Years 5.2

Initial Capital US$M 229 Sustaining Capital US$M 52 Closure Cost US$M 5

* Including MET and property tax

The sensitivity of NPV to metal price and recovery is shown in the following Table 1-25.

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Table 1-25: NPV Sensitivity to Gold Price and Discount Rate

Financial Metric Gold Price (US$/oz)

1100 1200 1250 1300 1400

NPV@5%Discount Rate 365 452 495 537 623

NPV@10%Discount Rate 221 281 311 340 400 NPV@15%Discount Rate 130 173 195 216 259

1.14 Risks and Opportunities

A risk assessment for the Kekura Gold project was completed during feasibility study. No very high priority risk was identified during the risk assessment analysis. Key risks with high priority, regardless of the risk categories, were identified as follows:

Delay due to missing the summer shipping window to Pevek port, particularly in 2018

Delay due to lack of availability of the winter road

Delay in critical equipment delivery because of damage during transit/storage

Delay due to Russian Design Institute (RDI) late finish of the design documentation

Increase in government approval timeline for project documentation (Rosnedra and Glavgosexpertiza)

Schedule impact due to lack of quality and capabilities of contractors, subcontractors and vendors.

Opportunities that have been identified to improve the NPV of the project are:

De-risk and understand the peroxide detoxification flow (bleed off %) required and effect on cake quality in terms of class V tailings achievement

Work with a selected pressure filtration vendor and perform test work to size the pressure filters. Optimise Diefenbach filter operational settings and most suitable materials of construction with Diefenbach.

Produce filter cake for subsequent geotechnical evaluation. Perform rheology test work on leach tailings to identify key design parameters for dry stacking such as optimal moisture content.

These enhancements to the project will be completed early during the execution stage value engineering period.

Documents

Sec.1 ExecutiveSummary EN 180129 AppdFinal...January 2018 Executive Summary Page 1-4 Figure 1-4: Typical Geological Cross-Section, Line PR2B The products of the hydrothermal alteration