Blueprint

EX-99.1 2 ego_ex991.htm KISLADAG TECHNICAL REPORT Blueprint

Technical Report

Kişladağ Gold Mine

Turkey

Centered on Latitude 38° 28' 56" N and Longitude 29° 08' 58" E

Effective Date: January 17, 2020

Prepared by:

Eldorado Gold Corporation

1188 Bentall 5 - 550 Burrard Street

Vancouver, BC V6C 2B5

Qualified Persons	Company
Mr. David Sutherland, P.Eng.	Eldorado Gold Corporation
Dr. Stephen Juras, P.Geo.	Eldorado Gold Corporation
Mr. Paul Skayman, FAusIMM	Eldorado Gold Corporation
Mr. Richard Miller, P.Eng.	Eldorado Gold Corporation
Mr. Sean McKinley, P.Geo.	Eldorado Gold Corporation

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Table of Contents

SECTION ● 1 SUMMARY	1-1
1.1 Introduction	1-1
1.2 Property Description and Location	1-1
1.3 Geology, Drilling and Sampling	1-2
1.4 Mineral Processing	1-4
1.5 Mineral Resources Estimates	1-4
1.6 Mineral Reserves and Mining Methods	1-6
1.7 Metallurgical Testwork and Recovery Methods	1-7
1.8 Project Infrastructure	1-8
1.9 Market Studies and Contracts	1-9
1.10 Environmental	1-9
1.11 Capital and Operating Costs	1-9
1.12 Economic Analysis	1-11
1.13 Interpretations and Conclusions	1-12
SECTION ● 2 INTRODUCTION	2-1
SECTION ● 3 RELIANCE ON OTHER EXPERTS	3-1
SECTION ● 4 PROPERTY DESCRIPTION AND LOCATION	4-1
4.1 Introduction	4-1
4.2 Property Location	4-1
4.3 Land Tenure	4-2
4.4 Royalties	4-3
4.5 Environmental Liabilities	4-3
4.6 Permits and Agreements	4-3
SECTION ● 5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	5-1
5.1 Site Topography	5-1
5.2 Accessibility	5-1
5.3 Physiography and Climate	5-2
5.4 Local Resources	5-2
SECTION ● 6 HISTORY	6-1
SECTION ● 7 GEOLOGICAL SETTING AND MINERALIZATION	7-1
7.1 Regional Geology	7-1
7.2 Local Geology	7-1
SECTION ● 8 DEPOSIT TYPES	8-1
8.1 Deposit Geology	8-1
8.2 Deposit Model	8-3
SECTION ● 9 EXPLORATION	9-1
SECTION ● 10 DRILLING	10-1
SECTION ● 11 SAMPLE PREPARATION, ANALYSES AND SECURITY	11-1
11.1 Sample Preparation and Assaying	11-1

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11.2 Quality Assurance/Quality Control (QA/QC)	11-2
11.3 Sample Counts for QA/QC	11-2
11.4 Blank Sample Performance	11-3
11.5 Standards Performance	11-4
11.6 Duplicate Performance	11-6
11.7 Specific Gravity Program	11-9
11.8 Concluding Statement	11-9
SECTION ● 12 DATA VERIFICATION	12-1
SECTION ● 13 MINERAL PROCESSING AND METALLURGICAL TESTWORK	13-1
13.1 Introduction	13-1
13.2 Ore Characterization	13-1
13.3 Drill Program	13-1
13.4 Sample Preparation	13-2
13.5 Column Leach Testing	13-3
13.6 Intermittent Bottle Roll Tests (IBRT)	13-4
13.7 High Pressure Grinding Rolls Investigation	13-5
SECION ● 14 MINERAL RESOURCE ESTIMATES	14-1
14.1 Geologic Models	14-1
14.2 Data Analysis	14-1
14.3 Evaluation of Extreme Grades	14-3
14.4 Variography	14-3
14.5 Model Setup	14-5
14.6 Estimation	14-5
14.7 Modelling of Gold Recovery from 2019 Column Testwork Program	14-6
14.8 Validation	14-7
14.9 Mineral Resource Summary	14-11
SECTION ● 15 MINERAL RESERVE ESTIMATES	15-1
15.1 Mineral Reserve Classification and Summary	15-1
15.2 Open Pit Optimization	15-2
15.3 Pit Design	15-13
15.4 Mineral Reserves	15-16
15.5 Risk Factors	15-16
SECTION ● 16 MINING METHODS	16-1
16.1 Introduction	16-1
16.2 Mine Design	16-3
16.3 Mine Production Schedule	16-8
SECTION ● 17 RECOVERY METHODS	17-1
17.1 General Description	17-1
17.2 Recovery Methods	17-1
17.3 Process Selection	17-2
17.4 Plant Design Basis	17-4
17.5 Process Description	17-4

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SECTION ● 18 PROJECT INFRASTRUCTURE	18-1
18.1 Site Location	18-1
18.2 Site Infrastructure	18-1
18.3 Water Management	18-6
SECTION ● 19 MARKET STUDIES AND CONTRACTS	19-1
19.1 Markets	19-1
19.2 Contracts	19-1
19.3 Taxes	19-1
SECTION ● 20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT	20-1
20.1 Baseline Conditions	20-1
20.2 Environmental Considerations	20-1
20.3 Social Impact	20-2
SECTION ● 21 CAPITAL AND OPERATING COSTS	21-1
21.1 Introduction	21-1
21.2 Growth Capital Costs	21-1
21.3 Sustaining Capital Costs	21-4
21.4 Operating Costs	21-7
SECTION ● 22 ECONOMIC ANALYSIS	22-1
22.1 Summary	22-1
22.2 Methods, Assumptions and Basis	22-1
22.3 Production Schedule	22-1
22.4 Royalties and Other Fees	22-2
22.5 Closure and Salvage Value	22-3
22.6 Taxation	22-3
22.7 Financing Costs	22-3
22.8 Third Party Interests	22-4
22.9 Sensitivity Analysis	22-4
22.10 Cash Flows	22-4
SECTION ● 23 ADJACENT PROPERTIES	23-1
SECTION ● 24 OTHER RELEVANT DATA AND INFORMATION	24-1
24.1 HPGR Project Schedule	24-1
24.2 Manpower Estimate	24-1
SECTION ● 25 INTERPRETATION AND CONCLUSIONS	25-1
25.1 Mineral Resources and Mineral Reserves	25-1
25.2 Mining Methods	25-1
25.3 Metallurgical Testwork	25-2
25.4 Process Design	25-2
25.5 Project Infrastructure	25-2
25.6 Waste Rock Dump	25-3
25.7 Capital and Operating Costs	25-3
25.8 Economic Analysis	25-3
SECTION ● 26 RECOMMENDATIONS	26-1
SECTION ● 27 REFERENCES	27-1
SECTION ● 28 DATE AND SIGNATURE PAGE	28-1

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LIST OF FIGURES

Figure 1 1: Location Map showing Western Turkey	1-2
Figure 1 2: Sensitivity Analysis on Gold Price	1-12
Figure 4 1: Location Map showing Project Location in Western Turkey	4-1
Figure 4 2: Kişladağ Land Position	4-3
Figure 5 1: Project Road Map	5-1
Figure 7 1: Geological Map of Uşak-Güre Basin showing the Location of the major Volcanic centers and Kışladağ Mine (modified after Karaoğlu et al., 2010)	7-3
Figure 7 2: Geological Map of the Kışladağ Deposit and surrounding Area (modified from Baker et al., 2016)	7-4
Figure 8 1: Geological Cross Section of the Kışladağ Deposit (modified from Baker et al., 2016)	8-1
Figure 8 2: Scanning Electron Microscope Images of Au located within Pyrite in Argillic Altered Sample and K-feldspar in Potassic altered Sample	8-4
Figure 10 1: Kışladağ Mine Drillhole Location Map	10-2
Figure 11 1: Kışladağ Blank Data – 2010 to 2011 Standard Blank COB05	11-3
Figure 11 2: Kışladağ Blank Data – 2015 to 2016 Standard Blank COB07	11-4
Figure 11 3: Standard Reference Material Chart, 2010 to 2011, Standard COS053 (KIS-14)	11-5
Figure 11 4: Standard Reference Material Chart, 2010 to 2011, Standard COS055 (KIS-16)	11-5
Figure 11 5: Standard Reference Material Chart, 2015 to 2016, Standard COS058 (KIS-19)	11-6
Figure 11 6: Standard Reference Material Chart, 2015 to 2016, Standard COS081 (SLGR05)	11-6
Figure 11 7: Relative Difference Plot of Kışladağ Coarse Reject Duplicate Data, 2010 to 2011	11-7
Figure 11 8: Percentile Rank Plot, Kışladağ Coarse Reject Duplicate Data, 2010 to 2011	11-8
Figure 11 9: Relative Difference Plot of Kışladağ Pulp Duplicate Data, 2015 to 2016	11-8
Figure 11 10: Percentile Rank Plot, Kışladağ Pulp Duplicate Data, 2015 to 2016	11-9
Figure 13 1: Metallurgical Drill Hole Location relative to the Kışladağ Open Pit	13-2
Figure 13 2: Averaged Column Gold Leach Kinetics	13-4
Figure 13 3: Averaged IBRT vs Column Gold Leach Kinetics	13-5
Figure 13 4: Averaged IBRT Tests Gold Recoveries	13-7
Figure 14 1: Relationship between the PACK or Mineralized Shell and Lithology Units	14-2
Figure 14 2: West – East Cross Section 4261400 N of Kişladağ modeled Gold Grades (g/t). Measured+Indicated Blocks are Full Size; Inferred Cells are the smaller Set	14-7
Figure 14 3: Plan view of Kişladağ modeled gold grades (g/t), 750 m Plan. Measured+Indicated Blocks are Full Size; Inferred Cells are the smaller Set	14-8
Figure 14 4: West – East Cross Section 4261400 N of Kişladağ Modeled Leach Recovery Values (%). Measured+Indicated Blocks are Full Size; Inferred Cells are the smaller Set	14-8
Figure 14 5: Herco Plots for Mineralization Shell	14-10
Figure 15 1: HPGR Ore Heap Leach Recovery Model on Section	15-4
Figure 15 2: Distribution of Au Recovery by Alteration	15-5
Figure 15 3: Primary Slope Sector Locations	15-7
Figure 15 4: Slope Coding in the Block Model	15-7
Figure 15 5: Bench Plan NSR & Lerchs-Grossmann Pit Limits	15-9
Figure 15 6: Cross Section Looking North	15-9
Figure 15 7: Cross Section Looking West	15-10
Figure 15 8: Pit Optimization Shells	15-13

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Figure 15 9: Pit Optimization Shells	15-13
Figure 15 10: Topography Surface Year End 2019	15-14
Figure 15 11: Final Pit Limits	15-15
Figure 15 12: Cross Section Looking Southwest	15-15
Figure 16 1: General Arrangement (photo taken early December 2019)	16-3
Figure 16 2: Design Elements	16-4
Figure 16 3: Footprint of Remaining Open Pit Production	16-5
Figure 16 4: 800 m Bench Pit Phase Contours	16-6
Figure 16 5: 700 m Bench Pit Phase Contours	16-7
Figure 16 6: Ore Qualities by Bench	16-7
Figure 16 7: Mine Material Movement Schedule	16-9
Figure 16 8: Haul Truck Requirement	16-9
Figure 16 9: Mine Development Mid-Year 2021	16-11
Figure 16 10: Mine Development 2023	16-12
Figure 16 11: Mine Development 2027	16-13
Figure 16 12: Mine Development 2034	16-14
Figure 16 13: Current Waste Dump Configuration (as of December 31, 2019)	16-15
Figure 17 1: Kişladağ Revised Flow Sheet	17-3
Figure 18 1: Project Area	18-2
Figure 22 1: Kişladağ Production Schedule and Grade	22-2
Figure 22 2: Kişladağ Gold Production Cash Flow	22-2
Figure 22 3: Sensitivity Analysis on Gold Price	22-4
Figure 24 1: Kişladağ HPGR Project, Implementation Schedule	24-1

LIST OF TABLES

Table 1 1: Kişladağ Mineral Resources, as of January 17, 2020	1-6
Table 1 2: Kişladağ, Mineral Reserves Effective January 17, 2020	1-6
Table 1 3: Growth Capital Cost Summary	1-10
Table 1 4: Sustaining Capital Cost Summary	1-10
Table 1 5: Operating Costs	1-11
Table 2 1: Cross-reference List	2-2
Table 4 1: Key Project Permits	4-4
Table 8 1: Summary of Au Deportment Work at Kışladağ	8-4
Table 10 1: Summary of Kışladağ Mine Drilling	10-1
Table 11 1: Number of Samples used for 2010-2011 and 2015-2016 Drill Campaigns	11-3
Table 12 1: Annual Reconciliation Summary	12-1
Table 14 1: Kişladağ Deposit Statistics for 5 m Composites – Au g/t Data	14-3
Table 14 2: Au Correlogram Parameters for Kişladağ Deposit	14-4
Table 14 3: Azimuth and Dip Angles of Rotated Correlogram Axes, Kişladağ Deposit	14-4

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Table 14 4: Global Model Mean Gold Values by Mineralized Shell Domain	14-10
Table 14 5: Kişladağ Mineral Resources, as of January 17, 2020	14-12
Table 15 1: Kışladağ, Mineral Reserve Estimates Effective January 17, 2020	15-1
Table 15 2: Recovery Summary	15-5
Table 15 3: Block Model Limits 2020	15-5
Table 15 4: Slope Sector Parameters	15-8
Table 15 5: Lerchs-Grossmann in-Pit Resources	15-11
Table 15 6: Kışladağ, Mineral Reserves Effective January 17, 2020	15-16
Table 16 1: Major Mining Equipment	16-1
Table 16 2: Final Pit Dimensions	16-5
Table 16 3: Mine Material Movement Schedule	16-10
Table 18 1: Water Sources Available	18-6
Table 21 1: Exchange Rates	21-1
Table 21 2: Growth Capital Cost Summary	21-1
Table 21 3: Primary Source for Unit Costs	21-3
Table 21 4: Primary Source of Quantities	21-3
Table 21 5: Basis of Indirect Costs	21-4
Table 21 6: Sustaining Capital Cost Summary	21-5
Table 21 7: Basis of Sustaining Capital	21-6
Table 21 8: Operating Costs	21-7
Table 22 1: Gold Royalty	22-3
Table 22 2: Annual Cashflow Summary	22-5

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GLOSSARY

Units of Measure


Annum (year)	a
Billion	B
Centimeter	cm
Cubic centimeter	cm3
Cubic meter	m3
Day	d
Days per year (annum)	d/a
Degree	°
Degrees Celsius	°C
Dollar (American)	US$
Dollar (Canadian)	CAN$
Euro	€
Gram	g
Grams per litre	g/L
Grams per tonne	g/t
Greater than	>
Hectare (10,000 m2)	ha
Hour	h
Hour per Year	h/y
Kilo (thousand)	k
Kilogram	kg
Kilograms per cubic meter	kg/m3
Kilograms per hour	kg/h
Kilograms per square meter	kg/m2
Kilometer	km
Kilometers per hour	km/h

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Kilopascal	kPa
Kilotonne	kt
Kilovolt	kV
Kilowatt hour	kWh
Kilowatt hours per tonne	kWh/t
Kilowatt hours per year	kWh/a
Kilowatt	kW
Less than	<
Litre	L
Megavolt Ampere	MVA
Megawatt	MW
Meter	m
Meter above Sea Level	masl
Metric ton (tonne)	t
Microns	µm
Milligram	mg
Milligrams per litre	mg/L
Millilitre	mL
Millimeter	mm
Million cubic meters	Mm3
Million ounces	Moz
Million tonnes per Annum	Mtpa
Million tonnes	Mt
Million	M
Million Years	Ma
Newton	N
Ounce	oz
Parts per billion	ppb
Parts per million	ppm
Percent	%
Percent by Weight	wt%
Square centimeter	cm2
Square kilometer	km2
Square meter	m2
Thousand tonnes	kt
Three Dimensional	3D
Tonne	t
Tonnes per day	t/d or tpd
Tonnes per hour	tph
Tonnes per year (annum)	tpa
Turkish Lira	₺
Volt	V
Watt	W
Weight/volume	w/v
Weight/weight	w/w

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Abbreviations and Acronyms

Acidity or Alkalinity	pH
Aluminum	Al
Analytical Detection Limit	ADL
Adsorption, Desorption, Regenerating	ADR
Antimony	Sb
Argillic	ARG
Argon	Ar
Arsenic	As
Association for the Advancement of Cost Engineering	AACE
Atomic Adsorption	AA
Barium	Ba
Bond Abrasion Index	Ai
Bond Ball Mill Work Index	BWi
Bond Rod Mill Work Index	RWi
Bottle Roll	BR
Bottle Roll Carbon in Pulp	BCIP
Bed Volumes	BV
Business Opening and Operations Permit	GSM
Cadmium	Cd
Calcium Hydroxide	Ca(OH)2
Carbon-in-leach	CIL
Canadian Institute of Mining, Metallurgy, and Petroleum	CIM
Cobalt	Co
Coefficient of Variance	CV
Construction Management	CM
Copper	Cu
Cyanide	CN
Diamond Drill Hole	DDH
Directorate of State Hydraulic Works	DSI

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Semi pure gold alloy	Doré
East	E
Eldorado Gold Corporation	Eldorado
Engineering, Procurement, Construction Management	EPCM
Environmental Impact Assessment	EIA
Environmental Management Plan	EMP
European Union	EU
Fast Radial Basis Function	FastRBF™
Feasibility Study	FS
Flocculant	FLOC
Flow Moisture Point	FMP
Friable	FRB
General and Administration	G&A
General Directorate of State Hydraulic Works	DSI
Geological Strength Index	GSI
Ground-Engaging Tools	GET
Gold	Au
Gold Equivalent	Au Equiv
HERCO Discrete Gaussian Model aka HERCO (Hermite Coefficient)	Herco
High Density Polyethylene	HDPE
High Grade	HG
High Pressure Grinding Roll	HPGR
Hydrochloric Acid	HCl
Hydrogen Oxide	H2O
Induced Polarization	IP
Inductively Coupled Plasma	ICP
Inner Diameter	ID
Internal Rate of Return	IRR
International Financial Reporting Standards	IFRS
International Organization for Standardization	ISO
Intrusion #3	INT3
Investment Tax Credit	ITC
Iron	Fe

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Kilborn Engineering Pacific Limited	Kilborn
Kişladağ Concentrate Treatment Plant	KCTP
Lead	Pb
Lerchs-Grossman	L-G
Life-of-mine	LOM
London Metal Exchange	LME
Manganese	Mn
Mechanical, Piping, Electrical, Instrumentation	MPEI
Measured & Indicated	M&I
Mercury	Hg
Micon International	Micon
Ministry of Environment and Urban Planning	MEUP
Motor Control Center	MCC
National Instrument 43-101	NI 43-101
Nearest Neighbour	NN
Nearest Neighbour Kriging	NNK
Net Present Value	NPV
Net Smelter Return	NSR
Nickel	Ni
North	N
North East	NE
North Heap Leach Facility	NHLP
North Rock Dump	NRD
North West	NW
Operator Control Station	OCS
Ordinary Kriging	OK
Outer Diameter	OD
Polyvinyl Chloride	PVC
Potassic	POT
Potassium	K
Potential of Hydrogen	pH
Prefeasibility Study	PFS
Probability Assisted Constrained Kriging	PACK
Process Control Systems	PCS
Programmable Logic Controllers	PLCs

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Quarter	Q
Qualified Person(s)	QP(s)
Quality assurance	QA
Quality control	QC
Quartz	Qz
Request for Quotations	RFQ
Reverse Circulation	RC
Rock Quality Designation	RQD
Run of Mine	ROM
Selective Mining Unit	SMU
Selenium	Se
Silicon	Si
Silver	Ag
Sodium Cyanide	NaCN
Sodium Hydroxide	NaOH
Sodium Metabisulphite	Na2S2O5
Sodium Metabisulphite	SMBS
South	S
South East	SE
South Rock Dump	SRD
South West	SW
Specific Gravity	SG
Spherical	SPH
Standard Reference Material	SRM
Strontium	Sn
Sulfur	S
Sulfur Dioxide	SO2
Sulphide	S2-
Sulphuric Acid	H2SO4
Technical Study	TS
Tourmaline	WMT
Transportable Moisture Limit	TML
Tüprag Metal Madencilik Sanayi Ve Ticaret Limited Sirketi	Tüprag
Turkish Electricity Distribution Corporation	TEDAS
Turkish Electricity Transmission Corporation	TEIAS
Uninterrupted Power Supply	UPS
Universal Transverse Mercador	UTM
Uranium	U
Value Added Tax	VAT
West	W
Work Breakdown Structure	WBS
Zinc	Zn

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SECTION ● 1

Summary

1.1

Introduction

Eldorado Gold Corporation (Eldorado), an international gold mining company based in Vancouver, British Columbia, owns and operates the Kişladağ gold mine in Turkey through its wholly owned Turkish subsidiary, Tüprag Metal Madencilik Sanayi Ve Ticaret Limited Sirketi (Tüprag). Eldorado has prepared this technical report of the Kişladağ gold mine in support of a material change in ore processing relative to the existing NI43-101 Technical Report whose Effective Date was March 16, 2018 (Technical Report, Kışladağ Milling Project, Turkey, 2018). After a period of successful test work on extended gold leaching, the Kışladağ mine has recently returned to being a heap leaching operation. This report describes the new metallurgy and field leaching procedures that form the basis of updated economics of the operation and its new mineral reserves.

Information and data for this report originated from the Kışladağ gold mine. The work entailed review of pertinent geological, mining, process and metallurgical data in sufficient detail to support the preparation of this technical report.

The qualified persons responsible for preparing this technical report as defined in National Instrument 43-101 (NI 43-101), Standards of Disclosure for Mineral Projects and in compliance with 43-101F1 (the “Technical Report”) are David Sutherland, P.Eng., Sean McKinley, P.Geo., Stephen Juras, Ph.D., P.Geo., Richard Miller, P.Eng. and Paul Skayman, FAusIMM, whom are all employees of Eldorado. All of the qualified persons have visited the Kışladağ gold mine.

1.2

Property Description and Location

The Kişladağ gold mine has been an operating open pit mine in commercial production since 2006 with surface facilities consisting of a crushing plant, heap leach pads and an adsorption, desorption, regeneration (ADR) plant, along with ancillary buildings.

Kişladağ is located in west-central Turkey lying 180 km to the west of the Aegean coast between Izmir and Ankara. The Project site lies 35 km southwest of the city of Uşak, which has a greater area population of approximately 370,000 inhabitants and near the village of Gümüşkol as shown on Figure 1-1. The mine site sits on the western edge of the Anatolian Plateau at an elevation of approximately 1,000 m, in gentle rolling topography. The climate in this region is arid with warm dry summers and mild wet winters.

There are no permanent water bodies in the area and water supply is limited to ephemeral streams and shallow seasonal stock ponds. Water is supplied to the mine from various well fields with a capacity of approximately 280 m3 per hour. A dam was constructed in partnership with the water authority in 2016 and is connected to the site to serve as an additional reservoir to support operations.

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The Turkish Electricity Distribution Corporation provides power to the site via two transmission lines from the Uşak industrial zone, 154 kV (27.7km) and 34.5 kV (25km).

Figure 1-1: Location Map showing Western Turkey

The Kişladağ Project land position consists of a single operating licence, number 85995, with a total area of 17,192 ha. According to Turkish mining law, Tüprag retains the right to explore and develop any mineral resources contained within the licence area provided fees and taxes are maintained. The licence was issued on April 9, 2003 and renewed on May 10, 2012 and is currently set to expire on May 10, 2032. Duration of mining licence can be extended if the mine production is still going on at the end of licence period.

No environmental liabilities have been assumed with the Project.

The current project Environmental Impact Assessment (EIA) area covers 2,509 ha. The land is classified as forestry (54%), treasury (11%), with the remaining area belonging to private land holders. As of December 31, 2019, Tüprag is the owner of 86.95% of the private land within the concession. The scope of the existing EIA is sufficient to accommodate envisioned heap leach pad Project.

1.3

Geology, Drilling and Sampling

Kışladağ gold mine is a gold-only porphyry deposit located in the eroded Miocene Beydağı stratovolcano in western Turkey. The gold mineralization occurs mainly within monzonite intrusive rocks emplaced within and above pre-Cretaceous Menderes metamorphic rocks. Deformation within the Beydağı volcanic sequence is minor in and around the deposit. Stratigraphic layering dips gently radially outward from the eroded center of the volcanic system, with no evidence of fault-related tilting.

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The Kışladağ deposit is hosted by a suite of nested subvolcanic monzonite porphyry intrusions that are subdivided into Intrusions #1, #2, #2A, and #3. Intrusion #1 is the oldest, and generally best mineralized phase. It forms the core of the system and is cut by the younger porphyritic intrusions. It is an E-W oriented elongate elliptical body (~1,300 m x ~500 m), and in the subsurface has a sill-like form intruding along the contact of the basement and volcanic package. At depth, the main body extends beyond the current limit of drilling (~1,000 m).

Alteration comprises an overlapping zoned system that contains a high temperature potassic core, an outer white mica-tourmaline zone and pervasive argillic alteration. The latter is particularly dominant in the western upper levels and throughout much of the surrounding volcanic sequence. Within the deposit, the largest zone of intense kaolinite alteration is focused in Intrusion #2A and a second smaller zone is focused in the southwest corner of the pit within Intrusion #1. Montmorillonite commonly overprints biotite in the potassic alteration zone. Porphyry-style sheeted to stockwork quartz veins occur with the potassic and white mica-tourmaline alteration zones.

Gold is very fine grained at Kışladağ. Gold in the argillic alteration occurs primarily with pyrite whereas in the white mica tourmaline alteration the gold grains occur with pyrite and muscovite. In the potassic samples, the majority of gold is hosted in K-feldspar.

Several drilling campaigns by both core drilling and RC drilling took place from 1998 through 2016 for a total of 198,000 m of which 38% was drilled in 2007 to 2010 and 26% in 2014 to 2016. It is this later drilling, mostly core holes, that provided information to enable upgrading of the mineral resource.

Diamond drilling in Kışladağ was done with wire line core rigs and mostly of HQ size. Drillers placed the core into wooden core boxes with each box holding about 4 m of HQ core. Geology and geotechnical data are collected from the core and core is photographed (wet) before sampling. SG measurements were done approximately every 5 m. Core recovery in the mineralized units was excellent, usually between 95% and 100%. The entire lengths of the diamond drill holes were sampled (sawn in half by diamond saw). The core library for the Kışladağ deposit is kept in core storage facilities on site.

Samples were prepared at Eldorado’s in-country preparation facility near Çanakkale in north-western Turkey. A Standard Reference Material (SRM), a duplicate and a blank sample were inserted into the sample stream at every 8th sample. From there the sample pulps were shipped to the ALS Chemex Analytical Laboratory in North Vancouver until April 2015 and Bureau Veritas (formerly Acme Labs) in Ankara since then. All samples were assayed for gold by 30 g fire assay with an AA finish and for multi-element determination using fusion digestion and ICP analysis.

Monitoring of the quality control samples showed that all data were in control throughout the preparation and analytical processes. In Eldorado’s opinion, the QA/QC results demonstrate that the Kışladağ deposit assay database is sufficiently accurate and precise for resource estimation.

1.4

Mineral Processing

The Kişladağ Project is an open pit mine and heap leach operation with a three-stage crushing plant. The process plant will continue to operate as a three-stage crushing plant but the third stage will be replaced by a high pressure grinding rolls circuit (HPGR). Ore is conveyed to a leach pad and irrigated with cyanide solution, solution is recovered and processed in an adsorption desorption regeneration (ADR) and electrowinning circuit to produce gold doré. The crushing circuit will process 12.6 Mt of ore per year resulting in approximately 160,000 ounces of gold produced annually.

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1.5

Mineral Resources Estimates

The estimates were made from a 3D block model utilizing commercial mine planning software. Projects limits, in UTM coordinates, are 686295 to 688655 East, 4260615 to 4262955 North, and 0 to +1110 m elevation. Block model cell size was 20 m east x 20 m north x 10 m high.

Eldorado used new data from the mining and the 2014-16 drilling campaign to update the geologic model described. The resource and reserve work incorporated lithology and alteration models, all constructed in 3D in Leapfrog Geo software. To constrain gold grade interpolation for the Kışladağ deposit, Eldorado created 3D mineralized envelopes or shells. These were based on initial outlines derived by a method of probability assisted constrained kriging (PACK). The threshold value of 0.20 g/t Au was determined by inspection of histograms and probability curves as well as by indicator variography. Shell outline selection was done by inspecting contoured probability values. These shapes were then edited on plan and section views to be consistent with the lithology model and the drill assay data so that the boundaries did not violate data and current geologic understanding of mineralization controls.

The block size for the Kişladağ model was selected based on mining selectivity considerations (open pit mining). It was assumed the smallest block size that could be selectively mined as ore or waste, referred to the selective mining unit (SMU) and was approximately 20 m x 20 m x 10 m. The assays were composited into 5 m fixed-length down-hole composites. The composite data were back-tagged by the mineralized shell and lithology units (on a majority code basis). Risk posed by extreme gold grades were checked; the gold distributions at Kışladağ do not indicate a problem with such grades.

Grade modelling consisted of interpolation by ordinary kriging (OK) for all domains inside the mineralized shell and inverse distance weighting to the second power (ID2) for background model blocks. Nearest-neighbor (NN) grades were also interpolated for validation purposes. Blocks and composites were matched on estimation domain.

A two-pass approach was instituted for interpolation. The first pass required a grade estimate to include composites from a minimum of two holes from the same estimation domain, whereas the second pass allowed a single hole to place a grade estimate in any uninterpolated block from the first pass. This approach was used to enable most blocks to receive a grade estimate within the domains, including the background domains. Blocks received a minimum of 2 to 3 and maximum of 3 to 4 composites from a single drill hole (for the two-hole minimum pass). Maximum composite limit ranged from 9 to 12. Searches had 95 to 500 m ranges for the estimation domains. The gold model was validated by visual inspection, checks for global bias and local trends and for appropriate levels of smoothing (change-of-support checks).

The mineral resources of the Kişladağ deposit were classified using logic consistent with the CIM Definition Standards for Mineral Resources and Mineral Reserves referred to in National Instrument 43-101. The mineralization of the project satisfies sufficient criteria to be classified into measured, indicated, and inferred mineral resource categories.

Inspection of the Kişladağ model and drillhole data on plans and cross-sections, combined with spatial statistical work and investigation of confidence limits in predicting planned annual and quarterly production, contributed to the setup of various distance to nearest composite protocols to help guide the assignment of blocks into measured or indicated mineral resource categories. Reasonable grade and geologic continuity is demonstrated over most of the Kişladağ deposit, which is drilled generally on 40 m to 80 m spaced sections. Blocks were classified as indicated mineral resources where blocks containing an estimate that resulted from samples spaced within 80 m and from two or more drill holes. Where the sample spacing was about 50 m or less, and the grade estimated were from at least three drill holes, the confidence in the grade estimates and lithology contacts were the highest and were thus permissive to be classified as measured mineral resources. All remaining model blocks containing a gold grade estimate were assigned as inferred mineral resources.

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The Kişladağ mineral resources as of January 17, 2020 are shown in Table 1-1. The Kişladağ mineral resource is reported at a 0.25 g/t Au cutoff grade (open pit) and a 0.60 g/t Au cutoff grade (underground) for measured, indicated and inferred mineral resources.

Table 1-1: Kişladağ Mineral Resources, as of January 17, 2020

Mineral Resource Category	Resource (t x 1,000)	Grade Au (g/t)	Contained Au (oz x 1,000)
Measured	345,440	0.63	6,975
Indicated	54,779	0.52	913
Measured & Indicated	400,219	0.61	7,888
Inferred	29,933	0.60	575

1.6

Mineral Reserves and Mining Methods

The operation uses conventional open pit techniques to feed crushing and heap leaching circuits to process the ore. The mineral reserves reported in this section are based upon continuation of heap leaching and a modification in the crushing circuit by the addition of a high-pressure grinding roll (HPGR) in place of the existing tertiary crushers from mid-2021 onwards.

The open pit optimization and pit design was completed using MineSight® software with comparative checks using Whittle® software. No dilution was included in the conversion of mineral resources to mineral reserves as the block modelling methodology (probability assisted constrained kriging) already accounts for dilution. Wall slope design incorporated inter-ramp slope angles by the usage of 15 sectors, created from analysis and modeling of multiple years geotechnical data.

The mineral reserves for the deposit were estimated using a gold price of US$1,250/oz and are effective January 17, 2020. The mineral reserves are reported using a cut-off grade of 0.19 g/t recoverable gold grade for ore that will be processed by heap leaching. This is roughly equivalent to a cut-off of US$7.29/t net smelter return (NSR) or an insitu gold cut-off grade of approximately 0.356 g/t. Mineral reserves are summarized in Table 1-2. The mineral reserves as reported are derived from and are included in the mineral resources.

Table 1-2: Kişladağ, Mineral Reserves Effective January 17, 2020

Mineral Reserves Category	Ore (t x 1,000)	Grade Au (g/t)	Contained Au (oz x 1,000)
Proven	164,531	0.73	3,851
Probable	8,644	0.57	159
Proven & Probable	173,175	0.72	4,010

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The mine is an open pit delivering 12.6 Mtpa ore to the three-stage crushing circuit followed by heap leaching. Life of mine strip ratio will be approximately 1.12:1. Mining methods are by conventional open pit techniques with unit operations consisting of drilling, blasting, loading, and hauling by truck.

The mine fleet includes seven diesel drills, two electric drills, one 29 m3 electric hydraulic shovel, two 21 m3 diesel hydraulic shovels, two 21.4 m3 wheel loaders, one 12 m3 wheel loader, fourteen 136 tonne trucks and ten 219 tonne trucks. The major equipment is supported by a fleet of graders, dozers, a backhoe and water trucks. No additional primary production equipment will be needed and no complete replacements will be required for this reserve.

Ore and waste are mined on 10 m benches. Ore will be hauled to the primary crusher for processing and waste rock will be placed in the south rock dump (SRD) and the north rock dump (NRD). The current SRD can accommodate 158.2 Mt of additional material from Dec 31, 2019. The north rock dump location has capacity far in excess of the remaining 34.9 Mt of waste that will be extracted with the ore reserves.

Mining has been designed in five phases with mining of the first two phases already completed. The mining of phase 3 is ongoing. The final pit will be approximately 1,650 m (east-west) x 1,300 m (north-south) x 565 m deep.

1.7

Metallurgical Testwork and Recovery Methods

In order to better understand gold leach recovery characteristics of remaining mineralization, a comprehensive PQ sized diamond core drilling program in the open pit was designed and executed in early 2019. The program comprised 117 holes with a total of 18,387 m. A total of 101 holes were drilled for metallurgical testwork on site and 16 holes were drilled to generate bulk samples for HPGR testing. Out of the 101 PQ diamond drill holes, 118 composite samples were created to represent different locations, depths, rock types, alteration types, and contacts between units. Throughout the deposit, the materials from 2 to 4 holes were combined to make up the composite samples. Bulk samples for HPGR testing were created on PQ sized holes where each sample was made out of 4 holes to create enough sample weight (~ 5 tonnes). Three bulk composite ore samples were created,

Each of the composite samples was crushed and screened in the metallurgical laboratory at the mine site to a particle size distribution, which closely mimics that of the plant crusher circuit. Parcels of each crushed composite sample were subjected to 2 m high, column leach tests for a 220 day duration. Given the lengthy column test duration requirements, small-scale IBRT tests were conducted in parallel with the columns to a) determine if a more rapid recovery value could be obtained for future recovery testing, and b) act as validation of the columns results. Typical duration for these tests is 5 to 10 days, however, the tailored approach, to better relate to the deeper Kışladağ mineralization, entailed increasing the duration to 45 days.

Testing revealed that the gold leach cycle time requirement has increased with depth in the open pit. Whereas previous, shallower mineralization required approximately 90 to 120 days of irrigation to meet “maximum” gold extraction, the deeper mineralization will require approximately 300 days of irrigation. Mean averaged gold recovery of the 2 m column tests was 51.8%; the 45 day IBRT tests returned averaged gold recoveries of 51%, confirming the 220-day column recoveries. The IBRT result confirms the procedure as valid for returning quicker recovery estimates.

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Products from the HPGR testwork were tested by IBRT for ultimate recovery. Averaged results show an additional 3.9 percentage points recovery were achieved at the highest operating pressure (4.4 N/mm2) giving a total of 54.8% in comparison to the base test 50.9%.

The column recovery data was modeled in 3D using Leapfrog Geo, an implicit modelling software. This created a base leach recovery model that mimicked the variation of the samples. An HPGR recovery model was made by simply adding 3.9% recovery units onto the interpolated column recovery values.

The ore is processed in a conventional heap leach facility which consists of a three-stage crushing plant. The process plant will be modified with the construction of an HPGR unit to replace the tertiary crushing circuit. Typical crushed product size is 80% passing 6.5 mm. An overland conveyor moves the ore from crushing plant to heap leach pad where mobile conveyors and a radial stacker places the crushed ore onto the leach pad. The radial stacker is designed to place the ore in 10 m high lifts in an 80 m wide sweep, the width of one cell. Dissolved gold is recovered in a carbon adsorption facility onto activated carbon. The gold-loaded carbon is then stripped on site in a refinery and the final product is a gold doré bar.

The existing south heap leach pad is a series of 30 cells each 80 m wide by 800 to 1,000 m long with a total size of approximately 1 km x 2.4 km. The south heap leach pad has a capacity of 234 Mt and is planned to be in service until 2029, when its maximum capacity is reached. A new north heap leach pad will be constructed directly north of the existing heap leach pad to provide additional 90 Mt of stacked ore capacity.

Material is now placed on interlift liner and is mixed with cyanide solution while being placed on the heap leach pad.

1.8

Project Infrastructure

The project does not have to upgrade the existing access road, power or water supplies.

A north leach pad facility, process and collection ponds will be constructed approximately 600 m north of south heap leach and will be accessed by a new overland conveyor connected to the current conveyor running along the east side of the south pad.

The SRD is centered about 1 km southwest of the open pit and currently holds approximately 210 Mt of waste rock with additional capacity of 158.2 Mt within the permit boundaries. A new NRD on the mountain west of the leach pads will need to be created. Designed to a capacity of 110 Mt, combined with the remaining SRD capacity will ensure sufficient capacity to hold the waste rock generated in the current mine plan.

The site is bounded by a series of collection ditches to divert non-contact water around the site to reduce the volume of contact water. All contact water is collected from the mine site and pit inflows and sent to collection ponds at the treatment plant. The treatment plant is located north of the existing ADR plant with a capacity of 625 m3/hr. On site there are numerous ponds to collect process streams (barren and pregnant solutions at the ADR plant), contact water, non-contact water, and surge ponds for storm events. The ponds were sized based on a 100-year storm event with additional capacities for storage and process surges.

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1.9

Market Studies and Contracts

Eldorado has not performed any formalized marketing studies in respect to Kişladağ gold production. Gold is currently sold on spot market via Turkish refiners by Tüprag’s internal sales department.

During 2019 Kişladağ sold gold at an average realized selling price of US$1,417 per troy ounce. The Turkish Central Bank has the right to purchase all gold produced at the site at LME spot prices.

Contracts and purchase agreements are currently in place for cyanide supply, diesel fuel and lube, explosives, leases of state lands, security and meal catering.

Corporate taxation for Turkish businesses is currently 22% through tax year 2020. In year 2021, the rate will be reduced to 20%. Depreciation is based mostly on a unit-of-production calculation under international financial reporting standards (IFRS). Turkish lira depreciation is based on the government’s depreciation list and this is mainly 10% for mine assets.

1.10

Environmental

Tüprag conducted baseline studies in 2000, 2001 and 2002 prior to development. An EIA was submitted January 2003, which was approved with Environmental Positive Certificate being granted in June 2003. Since mining began in 2005, Kışladağ mine operations have routinely collected environmental data outlined in the Environmental Management Plan (EMP) and submitted data to the relevant government agencies. Tüprag applied and received subsequent EIA amendments in 2011 and 2014 to increase the Kişladağ operations throughput to 12.5 Mtpa and 35 Mtpa respectively.

The Kişladağ gold mine employs approximately 82% of its labour force from Uşak and villages surrounding the mine. As an active part of the surrounding communities, the mine has completed numerous infrastructure programs within the region including primary schools, water works including the Gedikler Dam, and a 42 classroom building for Uşak University.

1.11

Capital and Operating Costs

The total growth capital cost includes the life-of-mine capitalized waste stripping costs, as well as the initial investment cost to obtain commercial product of a new HPGR circuit. Mining costs at Kişladağ are very well understood, and as such, actual productivities and costs were used as the basis of the mining costs. Growth capital costs are summarized in Table 1-3.

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Table 1-3: Growth Capital Cost Summary

Area	Growth Capital (US$ x 1,000)
Waste Stripping	254,970
HPGR	13,468
Other Plant Equipment and Platework	4,577
Civil Site Services	25
Concrete and Structural Steel	2,791
Electrical & Instrumentation	826
Sub-Total Direct Costs	276,657
Indirects and Owners	4,504
EPCM	2,386
Contingency	7,253
Total Installed Cost	290,800

The basis of the sustaining capital cost estimates vary based on the nature of the costs. Kişladağ has been in operation since 2006. Costs associated with earthworks, leach pad construction, and maintenance of capital equipment are very well understood. Ongoing contract rates, actual costs data, and other in-house data have been utilized as the basis of large majority of sustaining capital costs (summarized in Table 1-4).

Table 1-4: Sustaining Capital Cost Summary

Area	Sustaining Capital (US$ x 1,000)
Mining Equipment Rebuild	58,379
Miscellaneous Mining Capital	3,030
North Rock Dump	5,117
North Leach Pad	116,976
Interlift Liners	25,260
Other Process Capital	32,487
General and Administrative	1,101
Other Sustaining Construction	2,898
Total Cost	245,247

The operating cost estimate was developed based on a combination of first principle calculations, reliance on historical productivities, ongoing contract rates, past purchase orders, and actual annual operating costs for the crushing circuit, heap leach pad, plant infrastructure, and G&A.

The life-of-mine operating cost estimate has been benchmarked against actual operating cost data collected since the start of operations. Operating costs were calculated for each year of operation, totalling an average of US$111 M per annum and US$9.16/t ore over life-of-mine, summarized in Table 1-5.

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Table 1-5: Operating Costs

Category	LOM Average (US$/t)	LOM Expenditure (US$ x 1,000)
Mining	2.76	477,592
Processing	4.92	851,359
General and Administration	1.43	247,794
Transport and Refining	0.05	8,891
Operating Cost	9.16	1,585,727

1.12

Economic Analysis

The Kişladağ economics were analyzed using US$1,400/oz Au and a discount rate of 5%. The test of economic extraction for the Kışladağ mineral reserves is demonstrated by means of a sensitivity analysis. At the mineral reserve metals price of US$1,250/oz Au, the Kişladağ operations shows positive economics. Silver credit was assumed to be 0.50/oz silver payable per 1.0/oz gold payable, based on historical averages, the model assumes a silver price of US$18.00/oz; 100% of the gold recovered is payable.

Transport and refining costs of US$3.80/oz was used for economic analysis, based on historical averages at the Kişladağ operations.

The model has been prepared on a yearly life of mine basis. The LOM is 15 years from the beginning of 2020.

The after-tax net present value (NPV) of Kişladağ is estimated to be US$582 M using a gold price of US$1,400/oz and a discount rate of 5%. All capital investments planned for Kişladağ are self-funding.

The economic model was subjected to a sensitivity analysis to determine the effects of changing metal prices on the Project financial returns. Results are summarized in Figure 1-2.

Figure 1-2: Sensitivity Analysis on Gold Price

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1.13

Interpretations and Conclusions

It is concluded that the work to support this technical report indicate that the mineral resource and mineral reserve estimates and mine economics are sufficiently defined to indicate that the Kışladağ gold mine, with its new lengthier time leach recovery model is technically and economically viable as a gold heap leach operation. HPGR studies should advance to the basic engineering phase.

Eldorado has a high degree of confidence in the contents of this technical report.

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SECTION ● 2

Introduction

Eldorado Gold Corporation (Eldorado), an international gold mining company based in Vancouver, British Columbia, owns and operates the Kişladağ gold mine in Turkey through its wholly owned Turkish subsidiary, Tüprag Metal Madencilik Sanayi Ve Ticaret Limited Sirketi (Tüprag). Eldorado has prepared this technical report of the Kişladağ gold mine in support of a material change in ore processing relative to the existing NI43-101 Technical Report whose Effective Date was March 16, 2018 (Eldorado, 2018). After a period of successful test work on extended gold leaching, the Kişladağ mine resumed crushing in 2019 and continued the heap leaching operation. This report describes the new metallurgy and field leaching procedures that form the basis of updated economics of the operation and its new mineral reserves.

Information and data for this report originated from the Kişladağ gold mine. The work entailed review of pertinent geological, mining, process and metallurgical data in sufficient detail to support the preparation of this technical report.

When preparing reserves for any of its projects, Eldorado uses a consistent prevailing gold price methodology that is in line with the CIM Guidance on Commodity Pricing used in Resource and Reserve Estimation and Reporting. This was set for gold, as of September 2019 for Eldorado’s current mineral reserve work at US$1,250/oz Au. All cut-off grade determinations, mine designs and economic tests of extraction used this price in the mineral reserves work discussed in this technical report. To demonstrate the potential economics of a project, Eldorado may elect to use metal pricing closer to the current prevailing spot price and then provide some sensitivity around this price. For the Kişladağ gold mine, the gold price used for this evaluation was US$1,400/oz Au. This analysis (in Section 22 of this report) generally provides a better ‘snapshot’ of the project value at prevailing prices rather than limiting it to reserve prices, that might vary somewhat from prevailing spot prices. Eldorado stresses that only material that satisfies the mineral reserve criteria is subjected to further economic assessments at varied metal pricing.

David Sutherland, Project Manager, was responsible for overall preparation of the technical study and sections related to infrastructure and environment (report sections 1, 3, 4, 5, 6, 18, 20, 24, and 27). He most recently visited the Kişladağ gold mine on July 10 to 12, 2018.

Sean McKinley, Senior Geologist Resource Development was responsible for the preparation of the sections in this report concerned with geological information, exploration and drilling (report sections 7, 8, 9, 10 and 23). He most recently visited the Kişladağ gold mine on September 16 to 18, 2019.

Stephen Juras, Director, Technical Services, was responsible for the mineral resources and the preparation of related sections on sample preparation and analyses, data verification and mineral resource estimation (report sections 2, 11, 12, 14, 25 and 26). He most recently visited the Kişladağ gold mine on January 13 to 17, 2020.

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Richard Miller, Director of Mine Engineering (Open Pit), was responsible for the mineral reserves and the preparation of related sections on mineral reserves calculation, mining methods and sections related to costs (report sections 15, 16, 21 and 22). He most recently visited the Kişladağ gold mine on January 13 to 24, 2020.

Paul Skayman, Special Advisor to the COO, was responsible for the preparation of the sections in this report that dealt with metallurgy and process operations and related costs and payability (report sections 13, 17 and 19). He most recently visited the Kişladağ gold mine on January 13 to 17, 2020.

This document presents a summary of the current and forecast operation at the mine.

Currency used is US$ throughout, unless stated.

Turkish names frequently include Turkish characters. In some cases, the names may have been written using a standard US keyboard. The following table Table 2-1 is provided as a cross reference list.

Table 2-1: Cross-reference List

Standard US Keyboard Name	Turkish Name
Kisladag	Kışladağ
Kisla	Kışla
Usak	Uşak
Tuprag	Tüprag
Gokgoz Tepe	Gökgöz Tepe
Canakkale	Çanakkale
Gumuskol	Gümüşkol
Sogutlu	Söğütlü
Katrancilar	Katrancılar
Karapinar	Karapınar
Esme	Eşme
Sayacik	Sayacık
Dag	Dağ
TEDAS	Tedaş
Efemcukuru	Efemçukuru

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SECTION ● 3

Reliance on Other Experts

The qualified persons did not rely on a report, opinion or statement of another expert who is not a qualified person, concerning legal, political, environmental, or tax matters relevant to the technical report.

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SECTION ● 4

Property Description and Location

4.1

Introduction

The Kişladağ gold mine is an operating open pit mine in commercial production since 2006 with surface facilities consisting of a crushing plant, heap leach pads and an adsorption, desorption, regeneration (ADR) plant, along with ancillary buildings.

4.2

Property Location

Figure 4-1: Location Map showing Project Location in Western Turkey

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Approximate Project co-ordinates are:

●

UTM 06 87500E and 42 61600N

●

UTM Zone 35S

●

Map Sheet Uşak-L22 (1:100,000 scale)

●

Longitude 29° 08' 58" E

●

Latitude 38° 28' 56" N

Land use within the concession area falls into three categories: inhabited (villages and dwellings), agricultural land (cropping and grazing) and barren lands (not suitable for agriculture). Forestry land makes up about 54% of the Project area (2,509 ha) and treasury land makes approximately 11%. The remaining area belongs to private land. As of December 31, 2019, Tüprag is the owner of 86.95% of the private land within the concession.

There are no permanent water bodies in the area and water supply is limited to ephemeral streams and shallow seasonal stock ponds. Volcanic rocks with generally poor aquifer characteristics dominate the geology of the area. The villages in the area are supplied with potable water piped from a source located approximately 5 km to the west of Kişlaköy village.

The soil depth in the process plant area is less than two meters deep and subsurface conditions are characterized by weathered bedrock suitable for economical construction of equipment foundations.

The Kişladağ site is located approximately 250 km south of the major North Anatolian Fault zone and is located between the first and second-degree seismic zones as defined in the Turkish code. This is equivalent to an earthquake Zone 4 in the American Uniform Building Code. The effective ground acceleration coefficient is 0.4 g.

4.3

Land Tenure

The Kişladağ Project land position shown on Figure 4-2 consists of a single operating licence, number 85994, with a total area of 17,192.48 ha. According to Turkish mining law, Tüprag retains the right to explore and develop any mineral resources contained within the licence area provided fees and taxes are maintained. The licence was issued on April 9, 2003 and renewed on May 10, 2012 and is currently set to expire on May 10, 2032. Duration of mining licence can be extended if the mine production is still going on at the end of licence period.

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Figure 4-2: Kişladağ Land Position

4.4

Royalties

Royalties are discussed in Section 22.

4.5

Environmental Liabilities

No environmental liabilities have been assumed with the Project.

4.6

Permits and Agreements

The process of obtaining the necessary permits for a mining operation in Turkey is similar to the European Union EIA Directive. Table 4-1 lists key Project permits obtained to date, including the date and the governmental authority that issued them.

The project has received approval of supplementary EIA for the expansion of up to 35 Mtpa in 2014. The scope of the existing EIA is sufficient to accommodate the envisioned heap leach pad project. Applications will be made to amend existing forestry permits (Ministry of Forestry and Water), the GSM permit (Uşak Governor’s office) and Environmental Permit according to detailed facility designs.

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Table 4-1: Key Project Permits

Name of Permit	Issue Date	Issuer
Mining Operation Licence	2012-10-05	Ministry of Energy and Natural Resources
Mining Operation Permit	2012-10-05	Ministry of Energy and Natural Resources
EIA Permit	2014-06-24	Ministry of Environment
1Forestry Permit	2004-06-30	Directorship of Forestry
Opening Permit	2019-05-09	Provincial Administration of Uşak
2Grazing Land Permit	2019-01-16	Ministry of Agriculture and Forestry
3Environmental Permit and Licence	2019-03-22	Ministry of Environment

Notes:

1There are multiple forest permits. Permit durations are determined by the duration of the Mine Operation license (2032). After the mining operation licence and permit was renewed on 2012-10-05, the older 17 different forestry permits (starting from 2004) were combined under 3 permits and they were all renewed on 2017-05-04.

2 There are multiple grazing land permits. Permit durations are determined by the duration of the Mine Operation License (2032).

3 The Environmental Permit and Licence is renewed every five years from the date of issue.

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SECTION ● 5

Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1

Site Topography

The mine site sits on the western edge of the Anatolian Plateau at an elevation of approximately 1,000 m, in gentle rolling topography. Local elevations range from a peak of 1,300 masl (Kişladağ) to the adjacent valley of 700 masl.

5.2

Accessibility

The Kişladağ mine is situated 180 km west of the port city of Izmir in Uşak Province. Site is accessed from Izmir by traveling eastward on the D300 (E96) for approximately 185 km; the highway accesses Uşak, the provincial capital, then on to Ankara. The ED300 secondary highway accesses Eşme 32 km to the south. Travelling east from Esme along another secondary highway for 17 km to a 5.3 km private mine access road near the village of Gümüşkol, connecting the mine to the highway (Figure 5-1).

Figure 5-1: Project Road Map

The major cities of Izmir and Ankara are serviced by international airlines and there are regular internal flights by Turkish airlines to most major centers in the country. There is also an airport at Uşak for internal commercial flights servicing Istanbul.

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5.3

Physiography and Climate

The site surface consists of a thin soil on weathered bedrock. The area is dominated by a volcanic sequence with stony outcrops and rocky fields. There are areas forested by low pines but the region is predominantly grasslands and shrubs.

The climate in this region is arid with warm dry summers and mild wet winters. The Project site is located in a transition zone between the Continental and Mediterranean weather regimes. Temperature ranges from -5°C to 35 °C with extremes to -15°C and 40 °C.

5.4

Local Resources

There is a direct route to Uşak which accommodates a majority of the work force along a paved road 35 km north east of the site.

A 154 kV high voltage power line was commissioned in September 2016 from Uşak to site and has a capacity of 100 MVA. The original 34.5 kV line from Uşak with 10 MVA capacity is still in use servicing a portion of the overland conveying system.

Water is pumped from various well fields local to the site. The annual allotment from the wells is 2,482,234 m3/year (78.7 L/s), the pumping and pipeline system has a capacity of 324 m3 per hour.

A dam near Gedikler village was constructed in partnership with the water authority in 2016. A pumping and pipeline system connecting the dam to a site non-contact water pond was commissioned in 2018 with a capacity of 103 m3 per hour as reservoir contingency to support operations.

Contact and non-contact water collection systems and ponds are in place around the site to both collect water and mitigate against storm events. New ponds will be constructed at the north leach pad facility.

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SECTION ● 6

History

Eldorado acquired the Kişladağ property from Gencor Limited of South Africa in July 1996, as part of their portfolio of assets in Brazil and Turkey. The original prospect was identified by Tüprag geologists in 1989 from satellite image interpretations and confirmed through ground reconnaissance and geochemical sampling programs.

Since 1996, Eldorado’s exploration activities at Kişladağ have focused primarily on the zone known locally as Gökgöz Tepe, using principally geochemical soil and rockchip sampling, coupled with geological mapping. On the basis of this work, a gold anomaly was identified along the north slope of Gökgöz Tepe extending approximately 1,200 m on strike by 600 m wide. This work was followed in 1997 by 2,745 m of trench sampling, and 1,638 m of percussion drilling, which confirmed the mineralization in the shallow subsurface down to approximately 50 m over the footprint of the soil anomaly.

In 1998, a six hole HQ (96 mm outer diameter (OD) and 63.5 mm inner diameter (ID)) diamond drilling program (1,059 m) probing the main anomaly target intersected gold mineralization to depths of greater than 250 m and effectively confirmed the potential for large low grade bulk tonnage gold deposit. In 1999, an additional 5,000 m of HQ core drilling and 1,600 m of trenching extended the strike length and depth of the deposit. Based on the trenching, percussion drilling and core drilling data available to that date, Micon International (Micon) and Eldorado identified a measured and indicated resource of 42.8 Mt of 1.49 g/t, plus an inferred resource of 31.1 Mt at 1.35 g/t (all based on a 0.8 g/t cutoff grade).

In 2000, a reverse circulation (RC) drill program totaling 7,580 m (and 577 m of diamond drill hole (DDH)) led to a revised resource estimate and a significant increase in the deposit’s contained metal content. That year, Micon reported a measured and indicated resource of 125.97 Mt for the deposit at an average grade of 1.20 g/t Au. This is equivalent to 4.85 Moz of contained gold (using a cutoff grade of 0.4 g/t Au).

In 2002, a combined total of 10,582 m (RC, DDH and Percussion) was completed.

In 2003 to 2004, the drilling campaigns continued and a total of 8,499 m (RC and DDH) were drilled including 1,384 m for open pit geotechnical purposes. These geotechnical holes were also assayed and results were used for resource-reserve calculations later on.

Metallurgical test work initiated during 1999 and 2000 by Eldorado indicated that the ore would be amenable to heap leaching, and in 1999 Eldorado was granted a Site Selection Permit by the Turkish authorities for a gold mining operation at the Kişladağ Project site. Early receipt of this permit was made possible by the high level of support the Project has received from within the Uşak province as well as at the central government level.

In 2001 Eldorado commissioned a prefeasibility study with Kilborn Engineering Pacific Limited (Kilborn), based on the concept of recovering gold by heap leaching. This study considered an operation to treat 3.4 Mtpa of material based on an owner operated mining fleet and a three-stage crushing circuit generating a final crush size of 100% minus 8 mm. The objective of this approach was to minimize capital expenditure in the early years and allow for expansion to develop the total resource at a later date. Initial capital cost was estimated to be US$47.4 M with a cash operating cost estimated at US$154/oz and an average annual gold production of 103,600 troy ounces.

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Subsequent to issuing the prefeasibility study, Kilborn was asked to review the Project conditions in light of devaluation of the Turkish currency and to incorporate contract mining and utilizing used crushing equipment. An Addendum to the prefeasibility study was issued in December 2001, presenting a revised initial capital cost estimate of US$29.6 M and a cash operating cost estimate of US$149/oz.

In April 2003, a bankable feasibility study was completed by Hatch. The study envisaged a staged increase in production over a five year period from an initial production target of 5 Mtpa increasing to 10 Mtpa in Year 5. An optimization study was subsequently completed in July 2003, which generated a total life of mine capital cost estimate for the project of approximately US$138.5 million.

A technical report with a proven and probable mineral reserves equal to 115 Mt at a grade of 1.23 g/t Au (oxide cut-off 0.35 g/t Au; sulphide cut-off 0.50 g/t Au; gold price – US$325/oz) was declared and supported by Hatch (Technical Report, Hatch, 2003). Subsequent to the Hatch report additional drilling information was received and a new resource report was completed with declared measured and indicated mineral resources of 215 Mt grading 1.04 g/t Au (0.40 g/t Au cut-off) and supported by Micon (Technical Report, Micon, 2003).

Construction work started with access road construction in 2004. Work continued into 2006 with leach pad area preparation, construction of crushing, screening and ADR plants and ancillary buildings. Open pit production started in 2005. All construction work for the first phase was completed in early 2006 and commercial production was declared in July 2006.

Expansion of the crushing-screening plant to 10 Mtpa followed commercial production, with completion of the additional capacity in April 2007.

The operation was shut down in August 2007 after the Environmental Positive Certificate for Kişladağ had been challenged by a third party. The injunction was lifted in February 2008 and operation resumed in March 2008.

Owner operation in the pit replaced the mine contractor in September 2008.

Kişladağ operations further expanded during 2010 and in February 2011 the plant throughput increased to a rate of 12.5 Mtpa, with further optimizations to a rate of 13.1 Mtpa achieved in 2015 and 2016.

In 2015, detailed engineering was largely completed to expand the Kişladağ operations to process 20 Mtpa. The proposed design was to add a new 7.5 Mt/a, three-stage crushing circuit and a new primary crushing system to feed the new and existing plants. As a result of falling gold prices and other corporate capital projects the project was indefinitely cancelled.

In anticipation of the expansion, several infrastructure improvement projects were completed, notably, a new 154 kV substation with the capacity of 100 MVA and additional mining equipment.

In 2017, Eldorado completed an internal concept study followed by an internal scoping study to assess options for a milling circuit. These studies were supported by a large metallurgical testwork program. A prefeasibility level design of the milling circuit in February 2018 to support to support a NI43-101 Technical Report published March 2018. The report supported furthering the engineering and basic engineering for the mill was completed December 2018.

Crushing operations were suspended in April 2018 to conserve ore for milling and further evaluate the future options for the mine site. Heap leaching of existing placed material continued during this period. Prior to the suspension, a decision was taken to place some high grade material on a lined section of the pad, to study the overall recovery without the effect of the large leach pad, which was making metallurgical control difficult. This material leached as expected, and after 120 days, the predicted 42% recovery was achieved. Extended leaching and rest periods with no irrigation allowed us to recover nearly 60% of the gold with a leach cycle that was extended to 250 days. This formed the basis for the extended testwork program of extended time column tests and the longer-term IBRT test.

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SECTION ● 7

Geological Setting and Mineralization

7.1

Regional Geology

Western Turkey is host to several major porphyry and epithermal gold deposits including the Kişladağ porphyry gold mine, Efemçukuru intermediate sulfidation gold mine and Ovaçik low sulfidation gold mine. The gold-rich region is part of the Western Tethyan orogeny defined by a series of magmatic belts that have a strike length of over 3,600 km extending from Romania through Turkey, Iran, and continuing to the east through Pakistan into Central and Eastern Asia. The magmatic belts in Turkey trend younger in age north to south. In the north, Cretaceous to Paleogene subduction-related arc magmatism along the Pontide range transitions to post-collision, extension related Neogene magmatism in central and western Anatolia (Agostini et al., 2010; Jolivet et al., 2013). Deposits associated with the magmatism include a wide variety of porphyry, epithermal and sediment-hosted base and precious metal deposits (Richards, 2015).

The Kışladağ porphyry gold deposit occurs within the extension-related Neogene Uşak-Güre basin (Figure 7-1). The NE–SW-trending basin formed upon the Menderes metamorphic basement, which was exhumed during extension (Karaoğlu et al., 2010). Basement units of the massif include augen gneisses, schist and marble, and the structurally overlying Upper Cretaceous ophiolitic mélange rocks of the İzmir–Ankara zone (Ercan et al., 1978; Şengör et al., 1981). Basin fill units of the Uşak–Güre basin comprise the fluvio-lacustrine sedimentary packages of the Lower Miocene Hacıbekir Group, the Lower–Middle Miocene İnay Group and the Upper Miocene Asartepe Formation.

Within the basin there are three volcanic centers, Elmadağ, İtecektepe and Beydağı, with the latter hosting the Kışladağ deposit (Karaoğlu et al., 2010; Karaoğlu and Helvacı, 2012). The volcanic centers are stratovolcanoes, and the Beydağı stratovolcano in the southwest is the largest at ~16×9 km in diameter. Within the stratovolcanoes three distinct volcanic successions have been identified (Karaoğlu et al., 2010): (1) the Beydağı volcanic unit, composed of shoshonite, latites and rhyolitic lavas followed by dacitic and andesitic pyroclastic deposits; (2) the Payamtepe volcanic unit composed of potassic-intermediate lavas (latites and trachytes); and (3) the Karaağaç dikes composed of andesite and latite. 40Ar/39Ar dating by Karaoğlu et al. (2010) on biotite indicates that volcanic activity occurred between 17.3 Ma and 12.2 Ma, with the older ages from the northern Elmadağ volcanic center (17.3 to 16.3 Ma) and two younger ages of 15.0 Ma and 12.2 Ma from İtecektepe and Beydağı volcanoes respectively.

7.2

Local Geology

The Kışladağ deposit occurs mainly within intrusive rocks of the eroded Miocene Beydağı stratovolcano, which was emplaced within and overlies regional pre-Cretaceous Menderes metamorphic rocks. A minor amount of mineralization occurs within the volcaniclastic rocks of the Beydağı volcanic sequence and metamorphic basement rocks near the contacts with the mineralized intrusions. Menderes metamorphic rocks comprise schist and gneiss, and are exposed in erosional windows to the north and northwest of the deposit (pCM; Figure 7-2). The contact between the metamorphic basement and the volcanic sequence is a gently dipping unconformity, but within the deposit it is an irregular contact where deeper level, intrusive portions of the volcanic sequence cut the metamorphic rocks. Foliation and compositional layering in the metamorphic sequence is subhorizontal to gently dipping, with local variations related to small-scale folds.

The Beydağı volcanic sequence contains a variety of rock types within and surrounding the deposit that display rapid lateral and vertical facies changes. Six primary map units are defined (Figure 7-2): (1) volcanic conglomerate with monolithic porphyritic latite clasts (PBcg); (2) porphyritic quartz latite flows (PBq); (3) porphyritic latite flows commonly with flow-banded texture (PBf); (4) volcaniclastic rocks including lithic tuffs, volcaniclastic breccia, sandstone, siltstone, and mudstone or ash-fall tuff (PBvc); (5) monolithic porphyritic latite clast-supported breccia and minor mud-silt breccia to conglomerate (PBb); and (6) porphyritic hypabyssal monzonite intrusions, which in the mine are further divided into four subunits.

Deformation within the Beydağı volcanic sequence is minor in and around the Kışladağ deposit. Stratigraphic layering dips gently radially outward from the eroded center of the volcanic system, with no evidence of fault-related tilting; however, anomalously steep bedding (up to 45˚) occurs locally adjacent to intrusions. The overall geometry of the volcanic sequence and topographic features surrounding Kışladağ suggest that pre-erosion, the volcanic edifice rose ~1 km above the current erosional level.

There are no mappable fault offsets of stratigraphic or intrusive contacts, although mesoscopic structures consisting of low-offset fractures, joint sets, and veins are common. Fracturing is most intense in rocks adjacent to intrusive contacts and is also well developed in a NE-NNE striking corridor that occurs just east of the deposit characterized by joints, silicified outcrops, and silicified sheeted fractures and breccia zones. This fracture zone has not experienced significant offset since emplacement of the volcanic rocks because in several locations lithologic contacts can be mapped across it without displacement.

The Kışladağ deposit is centered on a set of nested subvolcanic porphyritic intrusions that were emplaced through the underlying Menderes metamorphic rocks into the Beydağı volcanic sequence. The intrusions are all monzonites based on their mineralogy and chemistry (Baker et al., 2016) and have been subdivided into Intrusions #1, #2, #2A, and #3, based on cross cutting relationships. Their ages range from 14.76 ± 0.01 to 14.36 ± 0.02 Ma (U-Pb zircon) that brackets the age of mineralization and is consistent with a rhenium-osmium radiometric age of 14.49 ± 0.06 Ma on Au-associated molybdenite (Baker et al. 2016). The highest gold grades in the Kışladağ deposit occur within the potassic core of the deposit centered on Intrusion #1. Surrounding and partly overlapping the potassic zone is a distinct white mica-tourmaline-(pyrite-albite-quartz) alteration. A poorly mineralized advanced argillic alteration (quartz-alunite-dickite-pyrophyllite-pyrite) post-dates the tourmaline-white mica assemblage and the most widespread alteration is argillic comprising kaolinite-smectite-pyrite-quartz.

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Figure 7-1: Geological Map of Uşak-Güre Basin showing the Location of the major Volcanic centers and Kışladağ Mine (modified after Karaoğlu et al., 2010)

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Figure 7-2: Geological Map of the Kışladağ Deposit and surrounding Area (modified from Baker et al., 2016)

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SECTION ● 8

Deposit Types

8.1

Deposit Geology

The Kışladağ deposit is hosted by a suite of subvolcanic monzonite porphyry intrusions that are subdivided into Intrusions #1, #2, #2A, and #3. Intrusion #1 is the oldest, and generally best mineralized phase. It forms the core of the system and is cut by the younger porphyritic intrusions. It is an E-W oriented elongate elliptical body in map view (~1,300 m x ~500 m), and in the subsurface has a sill-like form intruding along the contact of the basement and volcanic package (Figure 8-1). At depth the main body extends beyond the current limit of drilling (~1,000 m). Contacts between Intrusion #1 and the surrounding volcanic rocks are generally obscured by alteration. Contacts with younger intrusions, particularly Intrusion #3, are better preserved. Intrusion #1 has a K-feldspar-dominant groundmass with plagioclase phenocrysts (up to 30% of the rock by volume), occurring as tabular crystals ranging in size from 0.1 – 5 mm. Biotite is the second most abundant phenocryst phase (up to 10% of the rock) whereas blocky megacrystic K-feldspar phenocrysts, up to 1 cm, are a characteristic of this unit, but are low in abundance compared to plagioclase and biotite phenocrysts (< 2%). Quartz phenocrysts are rare, and are generally rounded or embayed where present. Intrusion #1 lacks amphibole or pyroxene phenocrysts and primary magnetite.

Figure 8-1: Geological Cross Section of the Kışladağ Deposit (modified from Baker et al., 2016)

Intrusion #2 occurs as a WNW-oriented elongate body at depth that splits into two apophyses that form semi-circular stocks (both approximately 150-200 m in diameter) at shallower levels where they cut Intrusion #1. Both apophyses are in contact with and cut by Intrusion #3. The rock is a fine- to medium-grained porphyry with abundant (20-30%) plagioclase phenocrysts up to 2 mm in length in a dominantly K-feldspar groundmass. Rare primary biotite and amphibole phenocrysts occur but the unit lacks quartz phenocrysts.

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Intrusion #2A occurs in the southeast corner of the pit and is characterized by a very intense clay (kaolinite-smectite-pyrite) alteration throughout that forms a distinct textural and rheological argillic altered sub-domain termed the friable domain. Intrusion #2A forms a circular stock (250-300 m across) that tapers at depth. It intrudes the margin of Intrusion #1 but contact relationships with Intrusion #3 are not observed. It is a fine- to medium-grained porphyritic rock, but the intense pervasive clay alteration obscures the primary mineral assemblage.

Intrusion #3 is the youngest large intrusive body. It forms a semi-circular stock near the center of Intrusion #1, west of the central Intrusion #2 stock and at depth to the west extends into a WNW-elongated, subvertical dike-like body. Intrusion #3 is a fine-grained porphyritic unit with 20-30% plagioclase phenocrysts up to 4 mm in length, and lesser quartz and biotite phenocrysts (both < 5%). Amphibole phenocrysts (5-10%) are more abundant than in the other intrusions but are commonly altered to chlorite. This intrusion is magnetic due to the presence of very fine-grained disseminated primary magnetite in the groundmass.

The oldest stage of alteration is a potassic assemblage characterized by the presence of secondary red-brown biotite and abundant pale pink-buff to nearly white K-feldspar. The biotite alteration is most intense in Intrusion #1 where it is associated with the highest gold grades. There are restricted occurrences of secondary magnetite associated with the potassic alteration, occurring as veinlets and as extremely fine-grained crystals intergrown with secondary biotite. A sodic-calcic sub-domain comprising actinolite-albite-magnetite alteration occurs locally within but overprints the potassic alteration and may form a deeper mineralized sodic-calcic core.

The tourmaline alteration is most intense immediately surrounding the potassic zone, however, the associated white mica alteration is more widespread and is particularly abundant on the west side of the deposit and spatially overlaps advanced argillic alteration. Within the intrusions tourmaline commonly occurs as envelopes around quartz ± pyrite veinlets, grading into black quartz-tourmaline matrix-supported hydrothermal breccias containing angular wallrock fragments. Tourmaline alteration of the volcanic rocks is more pervasive, though cross cutting breccia bodies are also present.

Quartz-alunite ± dickite ± pyrophyllite alteration is most abundant as a lithocap and as an alteration halo on the eastern side of the deposit. The advanced argillic alteration generally occurs in thick tabular zones that are localized along either stratigraphic contacts, within favorable volcaniclastic host rocks, or along fracture systems both within intrusions and volcanic rocks. Stratigraphically-controlled zones tend to form subhorizontal to gently dipping lithocaps peripheral to the deposit whereas structurally-controlled zones are steep-sided and form linear outcropping ridges. The advanced argillic alteration is typically poorly mineralized, and commonly oxidized with jarosite replacing pyrite.

Argillic alteration is pervasively developed throughout the deposit but is particularly dominant in the western upper levels and throughout much of the surrounding volcanic sequence. Within the deposit the largest zone of intense kaolinite alteration is focused in Intrusion #2A and a second smaller zone is focused in the southwest corner of the pit within Intrusion #1. Smectite, mainly montmorillonite and locally nontronite, commonly overprints biotite in the potassic alteration zone.

Porphyry-style sheeted to stockwork quartz veins occur with the potassic and white mica-tourmaline alteration zones. Veinlets range in width from 0.1 mm to 1 cm, with most being 1-3 mm. Gold occurs as non-refractory, very fine free gold grains (typically less than 10 microns in diameter) that are associated with pyrite, and less commonly other sulfide phases (chalcopyrite and sphalerite), as well as free grains attached to quartz, K-feldspar and albite. Both native gold and electrum (with up to 18 % Ag) have been identified. Other opaque minerals include pyrite, molybdenite and sphalerite, with minor occurrences of tennantite, tetrahedrite, bournonite, chalcopyrite and gold- and bismuth-telluride. The average copper grade of the deposit is low (~ 200 ppm) but increases to typically between 300 and 500 ppm within potassic alteration (Baker et al., 2016).

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Metallurgical testwork was carried out within the remaining mined mineral reserves on five alteration domains, namely argillic (ARG), potassic (POT), white mica tourmaline (WMT), friable (FRB) and Intrusion #3 (INT3; see Section 13). In addition, gold deportment studies were done in the main gold bearing alteration domains of argillic, white mica-tourmaline and potassic (Table 8-1). Thirty-one gold grains have been identified using scanning electron microscope based techniques (Figure 8-2). The gold grains have an average diameter of 3.8 microns with a range in size from 1.1 to 9.6 microns.

Gold in the argillic alteration occurs primarily with pyrite whereas in the WMT alteration the gold grains occur with pyrite and muscovite. In the potassic samples, the majority of gold is hosted in K-feldspar.

8.2

Deposit Model

The Kışladağ deposit is classified as a gold-only porphyry deposit due to its exceptionally low Cu%/Au ppm ratio (≈ 0.03; Baker et al., 2016). Significant analogues include the Maricunga porphyry deposits (9.8 Moz Au) in Chile and La Colosa (33.2 Moz Au) in Columbia. The low Cu/Au ratio may in part be related to the shallow level of emplacement (< 1 km) and volcanic setting, but also reflects the post-collisional extensional setting of the Miocene in western Turkey (Baker et al., 2016). Nonetheless, the deposit shares many characteristics with typical porphyry systems including: (i) multi-stage porphyry intrusions; (ii) a zoned alteration system that contain a high temperature potassic core, an outer white mica-tourmaline zone (analogous to phyllic alteration in typical porphyry deposits) and late, high level advanced argillic alteration; and (iii) porphyry style stockwork veins.

Table 8-1: Summary of Au Deportment Work at Kışladağ

Minerals around the Gold Grain: % of Gold Grain Boundary

Alteration Zone

No. of Samples

No. of Au Grains

Au Diameter Mean (Microns)

Pyrite

Clay

Muscovite

K-feldspar

Albite

Quartz

Chalcopyrite

Galena

Unknown

Argillic	1	2	6.68	94	6	-	-	-	-	-	-	-
WMT	3	2	7.46	20	-	50	-	-	-	-	23	7
Transitional WMT-Potassic	1	7	3.62	29	-	14	43	14		-	-	-
Potassic	5	20	3.12	14	-	2	58	15	7	3	-	1

Figure 8-2: Scanning Electron Microscope Images of Au located within Pyrite in Argillic Altered Sample and K-feldspar in Potassic altered Sample

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SECTION ● 9

Exploration

Eldorado has not undertaken any recent exploration works at the Kışladağ Project.

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SECTION ● 10

Drilling

Diamond-drill holes are the principal source of geologic, grade and metallurgical data for the Kışladağ mine since the start of mining in 2006. All diamond drilling in Kışladağ was done with wireline core rigs. Drillcores were mostly of HQ size (63.5 mm nominal core diameter). Drillers placed the core into wooden core boxes with each box holding about 4 m of HQ core. Geology and geotechnical data were collected from the core and core was photographed (wet) before sampling. Specific gravity (SG) measurements were done approximately every 5 m. Core recovery in the mineralized units was usually between 95% and 100%. The entire lengths of the diamond-drill holes were sampled (sawn in half by diamond saw). The core library for the Kışladağ deposit is kept in core storage facilities on site and used for subsequent technical investigations. Some intervals have been completely sampled during technical studies and physical core no longer exists.

Drilling totals are shown in Table 10-1. The pre-2004 campaigns were covered in an earlier Technical Report, Micon, 2003, and the 2004 to 2009 programs are described in a pre-existing Technical Report, Eldorado, 2010. Campaigns from 2010 to the present are the focus of this section and report.

Table 10-1: Summary of Kışladağ Mine Drilling

	Diamond Drilling		Reverse Circulation Drilling		Rotary Drilling
Period	# of holes	(m)	# of holes	(m)	# of holes	(m)
Pre-2004	53	12,269	145	21,298	44	2,264
2004-2006	9	862	8	1,329	-	-
2007-2009	81	34,603	14	3,558	-	-
2010	73	40,705	-	-	-	-
2011-2012	32	18,169	-	-	-	-
2014-2016	154	51,198	121	11,591	-	-
2018*	27	3,249	-	-	-	-
2019*	117	18,387	-	-	-	-
Total	546	179,442	288	37,775	44	2,264

*these drillholes were used only for metallurgical testing and for the estimation of the gold recovery model

The 2010 campaign comprised solely of diamond-drill holes which ranged in length from 305 to 902 m averaging 558 m. The 2011 to 2012 program consisted of diamond-drill holes ranging in length from 25 to 930 m and averaging 568 m. The 2014 to 2016 campaign comprised both reverse circulation (RC) and diamond drill programs. The 2014-16 diamond-drill holes ranged in length from 130 to 568 m and averaged 332.5 m. RC holes ranged from 23 to 145 m and averaged 96 m. The location of these drillholes are shown on a collar plan map in Figure 10-1.

Drilling was done by wireline method. Drillcores were most commonly HQ size and, less commonly, NQ size (47.6 mm nominal core diameter). Up to four drill rigs were used. Upon completion, the collar and anchor rods were removed and a polyvinyl chloride (PVC) pipe was inserted into the hole. The drillhole collars were located respective to a property grid. Proposed hole collars and completed collars were surveyed by the mine survey group.

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Figure 10-1: Kışladağ Mine Drillhole Location Map

The drillholes were drilled at an inclination of between 45° and 90°, with the majority between 60° and 70°. The holes were drilled mostly along 0°, 90°, 135°, 180°, and 310° azimuths. Down-hole surveys were taken approximately every 50 m by the drilling contractor mostly using a Reflex single-shot survey instrument.

Standard logging and sampling conventions were used to capture information from the drill core. The core was logged in detail onto paper logging sheets, and the data was then entered into the project database. The core was photographed before being sampled.

Eldorado reviewed the core logging procedures at site, and the drillcore was found to be well handled and maintained. Material was stored as stacked pallets in an organized “core farm”. Data collection was competently done. Core recovery in the mineralized units was excellent, averaging 95%. Overall the Kışladağ drill program and data capture were performed in a competent manner.

All of the drilling in the 2018 and 2019 campaigns was carried out to obtain fresh core samples for metallurgical testing (primarily gold recovery tests). In 2018, 3,249 m of HQ-size diamond drillcores were obtained from 27 drillholes. In 2019, 18,387 m of PQ-size diamond drillcores were obtained from 117 drillholes (described in Section 13). The 2018-19 metallurgical drillholes were not used in any mineral resource estimation, but the data obtained from this program was interpolated to create a block model of gold recovery, as discussed in Section 14.

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SECTION ● 11

Sample Preparation, Analyses and Security

Sample numbers were written on wooden core boxes allowing gaps in numbering sequence for control sample insertion. The entire lengths of the diamond-drill holes were sampled. Specific gravity measurements were done approximately every 5 meters. From 2007 to 2016 all core cutting and sampling was done on site at Kışladağ. The cut samples were sent to Eldorado’s sample preparation facility near Çanakkale in northwest Turkey.

Sampling on RC holes was done on 2.5 m intervals along the entire length of the drillholes. A cloth or perforated sample bag was used for RC samples. When drilling dry, drill cuttings were collected from the cyclone directly and split using a Jones splitter. After each sample was taken, the cyclone and drill string were blown clean. A small (~1 kg) sample of RC cuttings from each interval was collected for logging, spectral analysis and chipboard preparation. Wet RC sampling was done using a rotating wet splitter mounted on the rig. If the ground water flow was insufficient, extra water was injected through the rods to maintain the necessary flow rate for the rotating wet splitter. Wet samples were left to drain in a safe place and were then shipped to the company’s prep lab facility near Çanakkale. After samples were oven-dried, the same sample preparation protocol was applied to the RC samples as the diamond-drill samples.

The entire core library for the Kışladağ deposit is kept in core storage facilities on site. The coarse reject samples were stored off-site at the Çanakkale sample preparation lab. Samples from the core library and rejects were used over time to prepare single and composite metallurgical samples. From some intervals material is no longer available.

11.1

Sample Preparation and Assaying

Split drillcore and RC samples were prepared at Eldorado’s in-country preparation facility near Çanakkale in northwestern Turkey according to the following protocol:

●

The entire sample was crushed to 90% minus 3 mm (or 75% minus 2 mm).

●

A 1 kg subsample was split from the crushed, minus 3 mm sample and pulverized to 90% minus 75 µm (200 mesh). Prior to 2014, splitting was done by a riffle unit. After 2014, a Boyd rotary splitter was utilized.

●

A 110 g subsample was split off by taking multiple scoops from the pulverized 75 µm sample.

The 110 g subsample was placed in a Kraft paper envelope, sealed with a folded wire or glued top, and prepared for shipping. The rest of the pulverized sample was stored in plastic bags for later use.

All equipment was flushed with barren material and blasted with compressed air between each sample preparation procedure. Regular screen tests were done on the crushed and pulverized material to ensure that sample preparation specifications were being met.

A standard reference material (SRM), a duplicate and a blank sample were inserted into the sample stream at every 8th sample in order to monitor precision, possible contamination and accuracy respectively.

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The sample pulps were sent from Çanakkale to ALS Chemex Laboratories’ sample preparation facility in Izmir and were then shipped under the supervision of ALS Chemex to their Analytical Laboratory in North Vancouver, BC. After April 2015, Bureau Veritas (formerly Acme Labs) in Ankara was used to analyze Kışladağ drill samples. All samples were assayed for gold by 30-g fire assay with an atomic absorption (AA) finish and for multi-element geochemistry using fusion digestion and inductively coupled plasma (ICP) analysis.

11.2

Quality Assurance/Quality Control (QA/QC)

Assay results are provided to Eldorado in electronic format and as paper certificates. Upon receipt of assay results, values for SRMs and field blanks are tabulated and compared to the established pass-fail criteria as follows:

●

Automatic batch failure if the SRM result is greater than the round-robin limit of three standard deviations.

●

Automatic batch failure if two consecutive SRM results are greater than two standard deviations on the same side of the mean.

●

Automatic batch failure if the field blank result is over 0.03 g/t Au.

If a batch fails, it is re-assayed until it passes. Override allowances are made for barren batches. Batch pass/failure data are tabulated on an ongoing basis, and charts of individual reference material values with respect to round-robin tolerance limits are maintained.

Assay performance of data collected prior to 2010 has been described in detail in previous technical reports (Technical Report, Micon, 2003 and Technical Report, Eldorado, 2010).

The following sections will focus on the two drilling campaigns that contributed to major updates to the resource model: the 2010-2011 and 2015-2016 drilling campaigns.

11.3

Sample Counts for QA/QC

Table 11-1 shows the number of diamond drill core assay samples, blanks, duplicates and SRMs used in 2010-2011 and 2015-2016 drill campaigns.

Table 11-1: Number of Samples used for 2010-2011 and 2015-2016 Drill Campaigns

	2010-2011		2015-2016
Type of Sample	Sample Count	(%)	Sample Count	(%)
Core	20,241	87.15	19,676	86.94
Duplicate	987	4.25	971	4.29
Blank	999	4.3	979	4.33
SRM	998	4.3	1,005	4.44
Total Assayed	23,225	100	22,631	100

11.4

Blank Sample Performance

Assay performance of field blanks for gold is presented on Figure 11-1 for the 2010-2011 period and Figure 11-2 for the 2015-2016 drill campaign. The analytical detection limit (ADL) for gold is 0.005 g/t. The rejection threshold was chosen to equal 0.03 g/t. The results show no evidence of contamination. Rare higher values were investigated and found to be caused by sample mix-ups.

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Figure 11-1: Kışladağ Blank Data – 2010 to 2011 Standard Blank COB05

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Figure 11-2: Kışladağ Blank Data – 2015 to 2016 Standard Blank COB07

11.5

Standards Performance

Eldorado continuously monitors the performance of the SRM samples as the assay results arrive on site. Assaying during 2010-2015 used five different SRMs whereas 10 different SRMs were used during 2015-2016. The SRMs covered a grade range from 0.13 g/t Au to 3.20 g/t Au. Charts of the individual SRMs are shown on Figure 11-3 and Figure 11-4 for 2010-2011 period and on Figure 11-5 and Figure 11-6 for 2015-2016 period. All samples are given a “fail” flag as a default entry in the project database. Each sample is re-assigned a date-based “pass” flag when assays have passed acceptance criteria. At the data cutoff date of December 31, 2016, all samples had passed acceptance criteria. Some failures, marked with yellow boxes in the charts below, represent SRMs that upon investigation were found to have been inserted amongst unmineralized samples. These were deemed ignored and not used in any trend analysis of that SRM sample.

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Figure 11-3: Standard Reference Material Chart, 2010 to 2011, Standard COS053 (KIS-14)

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Figure 11-4: Standard Reference Material Chart, 2010 to 2011, Standard COS055 (KIS-16)

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Figure 11-5: Standard Reference Material Chart, 2015 to 2016, Standard COS058 (KIS-19)

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Figure 11-6: Standard Reference Material Chart, 2015 to 2016, Standard COS081 (SLGR05)

11.6

Duplicate Performance

Eldorado implemented a program that monitored data from regularly submitted coarse reject duplicates and pulp duplicates. These data showed good results. The duplicate data are shown in a relative difference chart in Figure 11-7 and percentile rank chart in Figure 11-8 for the 2010- 2011 period, and in Figure 11-9 and Figure 11-10 for the 2015-2016 period. Patterns observed in the relative difference plot are symmetric about zero suggesting no bias in the assay process. For the 90th percentile of the population, as shown on the percentile rank plot, a maximum difference of 20% is recommended for the coarse reject duplicates, whereas a maximum difference of 10% is recommended for the pulp duplicate data. The Kışladağ data shows 14% difference in the coarse reject data for 2010-2011 period and 8% difference in the pulp duplicate data for 2015-2016 period.

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Figure 11-7: Relative Difference Plot of Kışladağ Coarse Reject Duplicate Data, 2010 to 2011

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Figure 11-8: Percentile Rank Plot, Kışladağ Coarse Reject Duplicate Data, 2010 to 2011

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Figure 11-9: Relative Difference Plot of Kışladağ Pulp Duplicate Data, 2015 to 2016

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Figure 11-10: Percentile Rank Plot, Kışladağ Pulp Duplicate Data, 2015 to 2016

11.7

Specific Gravity Program

Samples taken for assay from drillcores were also measured for specific gravity and used to create a specific gravity 3-D model. The specific gravity for non-porous samples (the most common type) is calculated using the weights of representative samples in water (W2) and in air (W1). The bulk density is calculated by W1 / (W1-W2).

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11.8

Concluding Statement

Since the start of production in 2006, the entire drillhole database was reviewed in detail. Checks were made to the original assay certificates and survey data. Any discrepancies were corrected and incorporated into the current resource database. Eldorado therefore concludes that the data supporting the Kışladağ resource work are sufficiently free of error to be adequate for estimation. In Eldorado’s opinion, the QA/QC results demonstrate that the Kışladağ assay database is sufficiently accurate and precise for resource estimation.

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SECTION ● 12

Data Verification

The drillhole database created from drill campaigns from 2010 to 2016 was reviewed in detail. Cross-checks were made between the original assay certificates and downhole survey data and the digital database. Also, the descriptive information (lithology and alteration) was reviewed through relogging and pit mapping, and mineral identification by PIMA™ (a field portable, infrared spectrometer) analyses. Any discrepancies found were corrected and incorporated into the current resource database.

Another form of verification is the reconciliation to production of mined portions of the resource model. Results shown in Table 12-1 demonstrate that, to the end of 2017, the last full year of mining, excellent agreement exists between the mined production, crushed and placed on leach pad production, and the resource model.

Eldorado therefore concludes that the data supporting the Kışladağ resource work are sufficiently free of error and verifiable to support resource estimation.

Table 12-1: Annual Reconciliation Summary

Year	Resource Model tonnes	Resource Model Au Grade	Resource Model Au Oz’s	Mined Tonnes	Mined Au Grade	Mined Au Oz’s	Crushed and Placed on Pad Tonnes	Crushed and Placed on Pad Au Grade	Crushed and Placed on Pad Au Oz’s
	(x1,000)	(g/t)	(x1,000)	(x1,000)	(g/t)	(x1,000)	(x1,000)	(g/t)	(x1,000)
Pre-2008	10,246	1.14	377	10,756	1.19	412	10,504	1.21	409
2008	7,357	1.33	315	8,048	1.25	323	7,556	1.27	308
2009	10,525	1.20	406	10,550	1.13	383	10,717	1.11	383
2010	9,754	1.050	328	10,046	1.070	346	10,373	1.060	353
2011	13,449	0.980	424	12,522	0.970	391	12,430	0.950	380
2012	12,003	1.180	454	12,515	1.250	503	12,607	1.200	485
2013	13,340	1.136	487	13,432	1.120	484	13,297	1.120	478
2014	17,511	1.013	570	15,974	1.046	537	15,502	1.010	505
2015	20,508	0.724	477	19,515	0.719	451	19,147	0.700	430
2016	17,848	0.785	450	16,765	0.824	444	16,565	0.800	428
2017	14,997	0.987	476	13,952	1.067	478	13,062	1.030	434
TOTAL	147,537	1.005	4,765	144,075	1.002	4,752	141,760	0.969	4,594
% of Resource Model				98%	100%	100%	96%	96%	96%

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SECTION ● 13

Mineral Processing and Metallurgical Testwork

13.1

Introduction

The Kışladağ deposit is a low grade gold bearing porphyry amenable to heap leaching for gold recovery. Changes noted in the performance of the leach pad operations, along with corresponding testwork have indicated gold is being recovered at a slower-than-historical rate.

To investigate the metallurgical response of remaining mineralization in the deposit, a large testwork program has been conducted at the on-site metallurgical laboratory involving gold recovery tests on 118 bulk samples of PQ diamond drill cores.

Testing revealed that the gold leach cycle time requirement has increased with depth in the open pit. Whereas previous, mineralization required approximately 90 to 120 days of irrigation to meet “maximum” gold extraction, the deeper mineralization will require approximately 300 days of irrigation.

An ancillary testing program involved investigation of implementing HPGR (High Pressure Grinding Rolls) technology.

13.2

Ore Characterization

The mineralogy of Kişladağ remaining ore shows that gold occurs in fine grains (typically less than 10 µm in diameter) that are associated with pyrite and less commonly other sulfide phases (chalcopyrite, galena and sphalerite), as well as free gold grains included within K- feldspar, muscovite, albite and quartz. Both native gold and electrum (with up to 18 % Ag) have been identified.

The mineralization is hosted in a sheeted to stockwork quartz veined, intermediate intrusion and ancillary hydrothermal breccias and volcanoclasitic rocks. All show pervasive, often overlapping types of alteration, including potassic, argillic, tourmaline and white mica

13.3

Drill Program

In order to better understand gold leach recovery characteristics of remaining mineralization, a comprehensive PQ sized diamond core drilling program in the open pit was designed and started in January 2019. In a grid of approximately 50m by 50m, from accessible locations in the open pit, this PQ drilling program conducted 117 holes with a total of 18,387 m (Figure 13-1).

101 holes were drilled for metallurgical testwork on site and 16 holes were drilled to generate bulk samples for HPGR testing. Design of the drill hole lengths considered various large pit designs and grade distributions to ensure thorough coverage. The lengths varied between 70 and 270 m.

Out of the 101 PQ diamond drill holes, 118 composite samples were created to represent different locations, depths, rock types, alteration types, and contacts between units. Throughout the deposit, the materials from 2 to 4 holes were combined to make up the composite samples. In order to control the sample weights, one in every three meters of the core was selected from each hole and combined to make up composite samples. Each composite sample represented a 30 m thickness. Most holes were vertical; however, a few holes were inclined due to accessibility constraints.

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Bulk samples for HPGR testing were created on PQ sized holes where each sample was made out of 4 holes to create enough sample weight (~ 5 tonnes). Three bulk composite ore samples were created, two of which were composed of Potassic altered samples, and one of which comprised mixed Potassic and WMT alterations.

Figure 13-1: Metallurgical Drill Hole Location relative to the Kışladağ Open Pit

13.4

Sample Preparation

Each of the composite samples was crushed and screened in the metallurgical laboratory at the mine site to a particle size distribution, which closely mimics that of the plant crusher circuit. Seven months of plant crusher product sizing data (shift basis) were collated to determine the average size distribution. The months evaluated were those immediately prior to cessation of crushing operation in 2018, thereby best representing the future plant feed material.

Current Kışladağ crushed product size is 100% passing 12.5 mm for an 80% passing value (P80) of 6.5 mm. Approximately 40% of the crushed product is less than 1.7 mm.

In the laboratory, a three-stage control-crush procedure was developed, which involved screening of total product after individual crush stages enabling further crush manipulation of overly coarse material to generate a target fines component.

Due to the large mass of each composite ore sample, initial coarse primary jaw crushing was carried out and the product was rotary divided to yield approximately 150kg for secondary/tertiary jaw crushing on site. The coarse reject product was stored.

Final crushed product was homogenized and divided into test portions utilizing rotary splitter devices.

●

Head sample (Au, Ag, Cu, Stotal)

●

Particle size by size analysis (Au)

●

Column leach test (2 m)

●

Bottle roll leach test

Head sample analyses were conducted at the site assay laboratory. Gold, silver and copper were analysed by aqua regia acid digest followed by atomic adsorption spectroscopy (AAS). Total sulphur content was analyzed by Leco induction furnace.

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13.5

Column Leach Testing

Parcels of each crushed composite sample were subjected to column leach tests. Column tests are accepted as the best approach to simulate heap leach response and were operated under the following conditions:

●

Column dimensions are 2 m height by 0.15 m diameter. The column diameter is 12 times the largest particle size thereby minimizing any wall effect within the column i.e. solution channeling.

●

45 kg crushed sample for each column.

●

8 kg/t quicklime (CaO) blended with the sample prior to column loading.

●

Barren process solution from the current heap leach operation dosed at 550 ppm free NaCN was used.

●

Solution application rate was 7 L/h/m2 for 4 weeks, then reduced to 5 L/h/m2 for remaining duration. The reduction was implemented since actual heap operation generally involves a reduction of irrigation rate as leach pad surface deterioration over time leads to lower rates of solution acceptance.

●

Closed circuit operation with carbon adsorption, carbons were changed out 4 times during the entire duration of each test to ensure barren solution recycle.

●

220 days duration, an allocated duration to fit reporting time-frame requirements.

●

Products (pregnant and barren solutions) were monitored for Au, Ag and Cu, pH and cyanide concentration.

Daily pregnant/barren solution samples were assayed to assess ongoing gold dissolution rate and to develop gold recovery kinetics curves. The final mass balance and gold recovery calculations were based on the solid tails and loaded carbon assays along with the inclusion of gold removed in solution sub-sampling. All loaded carbon samples were assayed by an independent laboratory, ALS. The final solid tails were also assayed by ALS with additional verification by Bureau Veritas.

Given the long test duration and large number of resultant daily solution assays (pregnant and barren), to minimize any effect of compounding inaccuracies, it was deemed best to base the final mass balance on the carbon assays. The solution recovery kinetics curves were normalized to agree with the final carbon based recovery values

Mean averaged gold recovery was 51.8% However to better reflect local variations in the recovery data, a 3D recovery block model was created to use with the gold grade model. This is discussed in sections 14 and 15. Figure 13-2 depicts the average gold recovery kinetics over the 220-day duration. It is estimated that this will translate to approximately 300 leach days in the field.

Sodium cyanide (NaCN) consumption averaged 0.87 kg/t.

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Figure 13-2: Averaged Column Gold Leach Kinetics

13.6

Intermittent Bottle Roll Tests (IBRT)

Given the lengthy column test duration requirements, small-scale IBRT tests were conducted in parallel with the columns to a) determine if a more rapid recovery value could be obtained for future recovery testing, and b) act as validation of the columns results.

IBRT procedure is commonly used as a proxy to columns as they indicate maximum recovery values in a shorter time since they are operated under what are considered optimum conditions, i.e. material is treated as a slurry under intermittent agitation with a controlled pH and NaCN environment. Typical duration for these tests is 5 to 10 days, however, the tailored approach, to better relate to the deeper Kışladağ ore, entailed increasing the duration to 45 days. The 45 days was selected based on performance of the IBRT on a known sample that was placed on an interlift liner and leached in a large 800,000 tonne program.

General test conditions involved:

●

1.5 kg ore sample

●

3 L jars fitted with lifters and a hole in the lid to allow air ingress

●

55% solids pulp density with fresh water

●

pH controlled at 10 to 11 with quicklime (CaO)

●

750 ppm free NaCN maintained throughout test

●

Agitation was supplied by slow, intermittent rolling to avoid any autogenous grinding. The first 10 days involved rolling for 1 minute per hour which was subsequently reduced to 5 seconds per hour (1 revolution) until completion

●

Periodic solution sampling allowed for gold leach kinetics evaluation

●

Final solid tails and solutions were assayed for gold

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Averaged gold recoveries were 51%, which is very much in line with the 220-day column recoveries. This confirms the procedure as valid for returning quicker recovery estimates, and also validates the column results.

Figure 13-3 displays the averaged IBRT gold recovery kinetics in comparison with those of the columns.

Figure 13-3: Averaged IBRT vs Column Gold Leach Kinetics

13.7

High Pressure Grinding Rolls Investigation

HPGR (high pressure grinding rolls) is an established technology, and considered as an alternative to conventional crushing options i.e. cone crushers. When implemented at Kışladağ, a single HPGR unit will replace 5 tertiary cone crushers.

Heap leach operations globally are embracing HPGR due to the benefits of micro-fracturing induced by the rolls which are constantly imparting high pressures as the material passes. This micro-fracturing enables better permeation of the leachate into the particles, enhancing contact with gold grains.

Three world-leading producers of HPGR units were supplied with bulk composite samples (5 tonnes each) of Kışladağ PQ drill cores for pilot plant HPGR testing. The products were then returned to Kışladağ mine site for gold recovery testing.

The test composite samples were each comprised of 4 PQ drill holes spanning the base of the then-smallest pit design to the base of the then-largest, therefore typical of the materials from the deepest zones of the orebody. Two of the composite samples were comprised of purely potassic alteration, which is the hardest and most dominant alteration, accounting for greater than 50% of the remaining gold mineralization. The third composite sample was a blend of potassic and tourmaline alterations, tourmaline itself accounting for approximately 17% of the gold mineralization.

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Each vendor was supplied with a different composite sample to test under the same instructions i.e. three single-pass runs through their pilot scale HPGR unit at three different pressures. The discharge from each test was separated into centre product and edge product with both being subjected to gold recovery testing.

As the IBRT gold recovery procedure had been proven as a valid and more rapid technique, this method was utilized for the recovery testing of the HPGR products. Each IBRT test was completed in duplicate with results reported herein being the averaged values.

Recoveries from both streams were proportionally combined to determine if single-pass operation would be possible. The results, not included herein, clearly indicated that recycle of centre product would be beneficial due to larger particles resulting from the reduced pressure imparted at the rolls peripheries. It is considered that the pilot plant centre products, for the purpose of this investigation, best represented the final product from a recycle operation. It is anticipated that under operational conditions, the steady-state centre product will in fact yield higher recovery due to the recycle stream having additional contact with the rolls pressure (additional micro-fracturing).

Figure 13-4 shows centre product gold recovery results and a single result with material recycled from the edge product (in red). This clearly indicates that the specific press force (rolls pressure) is the determining operating parameter. An additional 3.9 percentage points recovery was achieved at the highest operating pressure (4.4 N/mm2) giving a total of 54.8% in comparison to the base test 50.9%. The recycled result in clearly shows a benefit in recirculating the edge product. This has not been factored into the additional recovery of 3.9% used in this analysis.

The single pass HPGR crush sizing was coarser (P80) than the base non-HPGR test with the high-pressure test returning a P80 of 7.9 mm compared to the normal crush of 6.5 mm. There is opportunity to further optimize both crush size and amount of recirculated load to further increase recovery.

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Figure 13-4: Averaged IBRT Tests Gold Recoveries

HPGR Crush with Increasing Rolls Pressures vs non-HPGR Base Test

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SECTION ● 14

Mineral Resource Estimates

14.1

Geologic Models

Eldorado used data from the mining and the 2014-16 drilling campaign to update the geologic model. The resource and reserve work incorporated new lithology and alteration models, all constructed in 3D in Leapfrog Geo software. Generally, there were no significant changes to the principal gold-hosting unit, Intrusion #1. The basement Schist unit was slightly enlarged in the West, South and South East directions. Intrusion #2 was modeled as a single entity while the contact between Intrusion #3 and Intrusion #1 became more irregular.

To constrain gold grade interpolation for the Kışladağ deposit, 3D mineralized envelopes, or shells were created. These were based on initial outlines derived by a method of probability assisted constrained kriging (PACK). The threshold value of 0.20 g/t Au was determined by inspection of histograms and probability curves as well as by indicator variography. Shell outline selection was done by inspecting contoured probability values. These shapes were then edited on plan and section views to be consistent with the lithology model and drill assay data so that the boundaries did not violate data and current geologic understanding of mineralization controls. Figure 14-1 shows the relationship between the PACK or mineralized shell and the lithology units.

All generated 3D shapes were checked for spatial and geological consistency on cross-section and plan views and were found to have been properly constructed. The shapes honoured the drill data and appear well constructed.

14.2

Data Analysis

The lithologic and mineralized domains were reviewed to determine appropriate estimation or grade interpolation parameters. Several different procedures were applied to the data to discover whether statistically distinct domains could be defined using the available geological objects. The lithology categories were investigated within and outside the mineralized shell.

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Figure 14-1: Relationship between the PACK or Mineralized Shell and Lithology Units

Descriptive statistics, histograms and cumulative probability plots, box plots and contact plots have been completed for gold assay data. The results were used to guide the construction of the block model and the development of estimation plans. These analyses were conducted on 5 m downhole composites of the assay data. The statistical properties from this analysis are summarized in Table 14-1.

14.2.1

Estimation Domains

Gold grades are highest and most prevalent in Intrusion #1. Younger units, Intrusions #2 and #2A, are also mineralized but at more uniform lower values with means of 0.58 g/t Au and 0.50 g/t Au respectively. Intrusion #3 and the pyroclastic rock unit generally contain weak to no gold mineralization. Gold mineralization above background levels within these two units occurs along the contact area with Intrusion #1. Generally, the coefficient of variance (CV) values of all units are relatively low reflecting the porphyry style mineralization of the deposit.

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Table 14-1: Kişladağ Deposit Statistics for 5 m Composites – Au g/t Data

Lithology	Mean	CV	Q25	Q50	Q75	Max	No. of Comps
Within PACK Shell
Intrusion #1	0.90	0.88	0.43	0.66	1.08	20.30	14,080
Intrusion #2	0.58	0.67	0.33	0.50	0.72	4.91	1,983
Intrusion #2A	0.50	0.65	0.29	0.43	0.61	4.79	1,681
Intrusion #3	0.42	0.85	0.23	0.33	0.50	4.42	1,931
Pyroclastics	0.37	0.84	0.21	0.29	0.43	9.20	3,394
Schist	0.40	0.77	0.24	0.32	0.45	3.13	999
Outside PACK Shell
Intrusion #1	0.15	1.02	0.08	0.12	0.16	1.41	511
Intrusion #2	0.18	1.28	0.09	0.13	0.17	1.50	95
Intrusion #2A	0.13	0.64	0.07	0.12	0.16	0.48	109
Intrusion #3	0.11	0.75	0.06	0.09	0.14	1.08	1,654
Pyroclastics	0.07	1.24	0.01	0.04	0.11	1.56	926
Schist	0.05	1.53	0.01	0.01	0.07	2.27	12,825

14.3

Evaluation of Extreme Grades

Extreme grades were examined for gold, mainly by histograms and cumulative probability plots. Generally, the distributions do not indicate a problem with extreme grades for gold. Less densely drilled areas of the deposit required the use of outlier-restricted grades to prevent the possibility of grade smearing. This is described in the estimation section below.

14.4

Variography

Variography, a continuation of data analysis, is the study of the spatial variability of an attribute. Eldorado prefers to use a correlogram, rather than the traditional variogram, because it is less sensitive to outliers and is normalized to the variance of data used for a given lag. Correlograms were calculated for gold in the mineralization shell. Variogram model parameters and orientation data of rotated variogram axes are shown in Table 14-2 and Table 14-3.

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Table 14-2: Au Correlogram Parameters for Kişladağ Deposit

Model

Nugget

Sills

Rotation Angles

Ranges

X1'

Y1''

X2'

Y2”

X1'

Y1''

X2'

Y2''

Inside the PACK Shell

SPH

0.25

0.377

0.373

-59

-11

713

187

248

Models are spherical (SPH). The first rotation is about Z, left hand rule is positive; the second rotation is about X', right hand rule is positive; the third rotation is about Y", left hand rule is positive.

Table 14-3: Azimuth and Dip Angles of Rotated Correlogram Axes, Kişladağ Deposit

	Axis Azimuth						Axis Dip
	Z1	X1	Y1	Z2	X2	Y2	Z1	X1	Y1	Z2	X2	Y2
Inside the PACK Shell	206	301	36	187	124	38	44	46	5	68	-10	19

Azimuths are in degrees. Dips are positive up and negative down.

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14.5

Model Setup

The block size for the Kişladağ model was selected based on mining selectivity considerations (open pit mining). It was assumed the smallest block size that could be selectively mined as ore or waste, referred to the selective mining unit (SMU), was approximately 20 m x 20 m x 10 m. In this case, the SMU grade-tonnage curves predicted by the restricted estimation process adequately represented the likely actual grade-tonnage distribution.

The assays were composited into 5 m fixed-length down-hole composites. The composite data were back-tagged by the mineralized shell and lithology units (on a majority code basis). The compositing process and subsequent back-tagging was reviewed and found to have performed as expected.

Bulk density data were assigned to a unique assay database file. These data were composited into 10 m fixed-length down-hole values. This compositing honoured the lithology domains by breaking the composites on the domain code values.

Various coding was done on the block model in preparation for grade interpolation. The block model was coded according to lithologic domain and mineralized shell (on a majority code basis). Percent below topography was also calculated into the model blocks.

A near surface oxidation of sulphide minerals has occurred at Kişladağ. Since leaching recoveries differ between the oxidized and primary mineralized rock, the boundary needs to be known for reserve conversion work. This model used an interpreted oxide surface. Since the start of production, the oxide – primary boundary has been defined through modeled total sulfur (S) % and cobalt (Co) ppm. The abundance of the latter element was found to be sensitive to the destruction of pyrite thus correlative to identifying oxidized areas. Modeled Co values also helped to distinguish between S values due to sulphide (i.e., pyrite) and S concentrations due to sulphates (alunite and barite).

14.6

Estimation

The search ellipsoids were oriented preferentially to the orientation of the respective domain as defined by the attitude of the gold grade shell and structures defined in the spatial analysis. Searches had 95 to 500 m ranges for the estimation domains. Block discretization was 4 m x 4 m x 2 m.

These parameters were based on the geological interpretation, data analyses, and correlogram analyses. The number of composites used in estimating grade into a model block followed a strategy that matched composite values and model blocks sharing the same ore code or domain. The minimum and maximum number of composites were adjusted to incorporate an appropriate amount of grade smoothing. This was done by change-of-support analysis (Discrete Gaussian or Hermitian polynomial change-of-support method), as described in the validation section below.

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In all domains, an outlier restriction was used to control the effects of high-grade composites in local areas of less dense drilling, particularly in background domains and poorly mineralized units (e.g., Intrusion #3). The restricted distance was 50 m meaning that beyond this distance from a model block center, composites exceeding the outlier values are not used in estimation. The threshold grades were generally set close to the threshold grade of the PACK shell in the case of the background domains or through inspection of the cumulative probability plots for the mineralized units. Mineralized domains in Intrusion #1, #2 and #2A and Pyroclastics used an outlier restricted grade limit of 7.0 g/t Au, whereas mineralized Intrusion #3 unit used an outlier limit equal to 1 g/t Au. All background domains used a 0.5 g/t Au outlier restricted grade except for Intrusion #3 and Schist, where the outlier grade equaled 0.3 g/t Au and 1 g/t Au respectively.

Bulk density values were estimated into the resource model by an averaging of 10 m composites of individual density measurements that were carried out on each assay interval. A maximum of six and minimum of two 10 m composites were used for the averaging. A rectangular search was used, measuring 200 m north x 125 m east x 50 m elevation. In the event a block was not estimated, default density values were assigned based on lithology and oxidation code.

14.7

Modelling of Gold Recovery from 2019 Column Testwork Program

A major diamond drill campaign consisting of 117 PQ diameter drill holes was executed in early 2019 to provide sufficient samples for a significant leach recovery testwork program whose results were to form the basis of a new gold leach recovery model. The drilling campaign, sample philosophy and testwork are described in Section 13. The main testwork comprised 117 two-meter columns that stayed under leached for at least 220 days. The carbon calculated recoveries for each of the columns are used in the gold recovery block model. Values in that recovery model were created by 3-D numerical modeling in Leapfrog Geo software. Weighting was added to the interpolation in the X and Y directions relative to the Z direction (1.75 x 1.75 x 1.00) in order to yield results that more closely mimicked known geological trends. A spheroidal numerical model with a base range of 250 m was chosen as the interpolant. All blocks in the mineral resource block model were ‘evaluated’, or assigned an estimated gold recovery, using the interpolated values from the Leapfrog model.

3-D modelling of the column recovery data used Seequent’s Leapfrog Geo (version 5.0.3) software. Leapfrog Geo is an implicit modelling package that utilizes their Fast Radial Basis Function (FastRBF™) algorithms for rapid data interpolation.

This modeling effort created the base leach recovery model. Installation of an HPGR unit in place of the tertiary crushing circuit is planned for Kişladağ. Metallurgical testwork (Section 13) has shown that the unique micro-fracturing generated by this type of crushing unit did have a positive effect on the expected recovery. However, since this testwork is only carried out at a pilot plant level of scale (~5 tonnes per test), the identified recovery enhancement is simply treated as a constant value that is added on to the base modeled recovery. The calculated HPGR recovery upgrade value was 3.9% recovery units.

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14.8

Validation

14.8.1

Visual Inspection

Eldorado completed a detailed visual validation of the Kişladağ resource model. The model was checked for proper coding of drillhole intervals and block model cells, in both cross-section and plan views. Coding was found to be accurate. Grade and gold recovery interpolation was examined relative to drill hole composite values by inspecting cross-sections and plans. The checks showed good agreement between drill hole composite values and model cell values. The hard boundaries appear to have constrained grades to their respective estimation domains. The addition of the outlier restriction values succeeded in minimizing grade smearing in regions of sparse data and, in general, all background domains. Examples of representative sections and plans containing block model grades, drill hole composite values, and domain outlines are shown in Figure 14-2 to Figure 14-4.

Figure 14-2: West – East Cross Section 4261400 N of Kişladağ modeled Gold Grades (g/t). Measured+Indicated Blocks are Full Size; Inferred Cells are the smaller Set

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Figure 14-3: Plan view of Kişladağ modeled gold grades (g/t), 750 m Plan. Measured+Indicated Blocks are Full Size; Inferred Cells are the smaller Set

Figure 14-4: West – East Cross Section 4261400 N of Kişladağ Modeled Leach Recovery Values (%). Measured+Indicated Blocks are Full Size; Inferred Cells are the smaller Set

14.8.2

Model Check for Change-of-Support

An independent check on the smoothing in the grade estimates was made using the Discrete Gaussian or Hermitian polynomial change-of-support method. This method uses the “declustered” distribution of composite grades from a NN or polygonal model to predict the distribution of grades in blocks. The histogram for the blocks is derived from two calculations:

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●

The block-to-block or between-block variance

●

The frequency distribution for the composite grades transformed by means of Hermite polynomials (Herco) into a less skewed distribution with the same mean as the declustered grade distribution and with the block-to-block variance of the grades.

The distribution of hypothetical block grades derived by the Herco method is then compared to the estimated grade distribution to be validated by means of grade-tonnage curves.

The distribution of calculated 20 m x 20 m x 10 m block grades for gold in the mineralized domain is shown with dashed lines on the grade-tonnage curves in Figure 14-5. This is the distribution of grades obtained from the change-of-support models. The continuous lines in the figures show the grade-tonnage distribution obtained from the block estimates. The grade-tonnage predictions produced for the model show that grade and tonnage estimates are validated by the change-of-support calculations over the range of mining grade cutoff values (0.3 g/t to 0.5 g/t Au).

14.8.3

Model Checks for Bias

The block model estimates were checked for global bias by comparing the average metal grades (with no cutoff) from the model with means from NN estimates. The NN estimator declusters the data and produces a theoretically unbiased estimate of the average value when no cutoff grade is imposed and is a good basis for checking the performance of different estimation methods. Results, summarized in Table 14-4, show no global bias in the estimates.

The model was also checked for local trends in the grade estimates by grade slice or swath checks. This was done by plotting the mean values from the NN estimate versus the kriged results for benches (in 5 m swaths) and for northings and eastings (both in 20 m swaths). The kriged estimate should be smoother than the NN estimate, thus the NN estimate should fluctuate around the kriged estimate on the plots. The observed trends behave as predicted and show no significant trends of gold in the estimates in Kişladağ model.

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Figure 14-5: Herco Plots for Mineralization Shell

Table 14-4: Global Model Mean Gold Values by Mineralized Shell Domain

Domain	NN Estimate	Kriged Estimate	Difference (%)
Domain	NN Estimate	Kriged Estimate	Difference (%)
Intrusion #1	0.69	0.69	0.0
Intrusion #2	0.55	0.56	-1.6
Intrusion #2A	0.52	0.50	3.3
Pyroclastics	0.16	0.15	10.5
Intrusion #3	0.33	0.34	-3.6
Schist	0.34	0.34	0.6

14.8.4

Mineral Resource Classification

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Inspection of the Kişladağ model and drillhole data on plans and cross-sections, combined with spatial statistical work and investigation of confidence limits in predicting planned annual and quarterly production, contributed to the setup of various distance to nearest composite protocols to help guide the assignment of blocks into measured or indicated mineral resource categories. Reasonable grade and geologic continuity is demonstrated over most of the Kişladağ deposit, which is drilled generally on 40 m to 80 m spaced sections. Blocks were classified as indicated mineral resources where blocks containing an estimate that resulted from samples spaced within 80 m and from two or more drill holes. Where the sample spacing was about 50 m or less, the confidence in the grade estimates and lithology contacts were the highest and were thus permissive to be classified as measured mineral resources. This was facilitated where such blocks contained an estimate from samples of three or more drill holes.

All remaining model blocks containing a gold grade estimate were assigned as inferred mineral resources.

A test of reasonableness for the expectation of economic extraction was made on the Kişladağ mineral resources by developing a series of open pit designs based on optimal operational parameters and gold price assumptions. A pit design based on $1800/oz Au and heap leaching was chosen to constrain mineral resources likely to be mined by open pit mining methods (Figure 14-2). Eligible model blocks within this pit shell were evaluated at an open pit resource cut-off grade of 0.25 g/t Au. For interpolated blocks lying outside this pit design, likely mining would be by underground methods. The necessary economic threshold would be higher; thus a cut-off grade of 0.60 g/t Au was chosen (consistent with what Eldorado used to declare similar underground mineral resources at its Skouries and Bolcana projects). The cut-off grade alone was not sufficient to assign outside pit mineral resources. Appropriate volumes of contiguous model blocks needed to be also identified. Guiding polygons were drawn defining likely volumes of mineable above cut-off grade material (see Figure 14-2). As the mining method would be bulk in nature, selectivity will be limited. Consequently, all blocks within these polygons, regardless of gold grade, were tabulated as the underground permissive mineral resource. Finally, due to increased uncertainty introduced by invoking a different mining method on these outside pit resources, all measured eligible blocks were downgraded to Indicated.

14.9

Mineral Resource Summary

The Kişladağ mineral resources as of January 17, 2020 are shown in Table 14-5. The Kişladağ mineral resource is reported at a 0.25 g/t Au cutoff grade (open pit) and a 0.60 g/t Au cutoff grade (underground) for measured, indicated and inferred mineral resources.

Table 14-5: Kişladağ Mineral Resources, as of January 17, 2020

Mineral Resource Category	Resource (t x 1,000)	Grade Au (g/t)	Contained Au (oz x 1,000)
Open Pit Resources
Measured	345,440	0.63	6,975
Indicated	47,645	0.49	749
Measured & Indicated	393,086	0.61	7,724
Inferred	9,354	0.44	131
Underground Resources
Measured	--	--	--
Indicated	7,133	0.72	164
Measured & Indicated	7,133	0.72	164
Inferred	20,579	0.67	434
Total Resources
Measured	345,440	0.63	6,975
Indicated	54,779	0.52	913
Measured & Indicated	400,219	0.61	7,888
Inferred	29,933	0.60	575

Au cut-offs: Open Pit = 0.25 g/t; Underground = 0.60 g/t

Only material within an $1800/oz pit design was cast as open pit resources; all other material was considered as underground mineral resources tabulated to the December 31, 2019 open pit surface.

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SECTION ● 15

Mineral Reserve Estimates

The Kışladağ gold mine historically processed ore through a three-stage crushing circuit and heap leaching facility. The mineral reserves reported in this section are primarily based upon a modification in processing methods that includes the addition of an HPGR in place of the existing tertiary crushers from mid-2021 onwards.

This section describes the open pit optimization process including key assumptions and economic considerations leading to pit limit selection and the reporting of mineral reserves used for mine planning and scheduling as described in Section 16.

The open pit optimization was completed using MineSight® software with comparative checks using Whittle® software. The detailed pit design was also primarily designed using MineSight® software with some final adjustments completed using GEMS® software.

The mineral reserves have been estimated and classified in compliance with the CIM Definition Standards for Mineral Resources and Mineral Reserves.

The mineral resource model as referenced in Section 14 of this report was used as input for the mineral reserve estimates. The modelling methods, grade models, resource classification, and density model were reviewed by Eldorado and found appropriate for mineral reserve estimation. Only measured and indicated resources were used in the pit optimization and reserve reporting.

The open pit optimization was performed using a Lerchs-Grossmann algorithm as described in Section 16. An optimum pit shell was selected to provide a basis for the pit design used to report mineral reserves.

15.1

Mineral Reserve Classification and Summary

The mineral reserves are the measured and indicated resource blocks that are within the reserve pit design and above the ore cut-off grade. The total proven and probable mineral reserves estimate is 173.2 Mt at a grade of 0.72 g/t Au. Table 15-1 provides further details of the mineral reserve estimates. The mineral reserves are effective January 17, 2020.

Table 15-1: Kışladağ, Mineral Reserve Estimates Effective January 17, 2020

Reserve Classification	Ore (t x 1,000)	Grade Au (g/t)	Contained Au (oz x 1,000 )
Proven	164,531	0.73	3,851
Probable	8,644	0.57	159
Proven & Probable	173,175	0.72	4,010

The mineral reserves as reported are derived from and are included in the mineral resources.

No dilution was included in the conversion of mineral resources to mineral reserves. The block modelling methodology (probability assisted constrained kriging with 20m by 20m by 10m block size) already accounts for dilution.

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The cut-off grade for the mineral reserves is 0.19 g/t recoverable gold grade for ore that will be processed by heap leaching. This is roughly equivalent to a cut-off of US$7.29/t net smelter return (NSR). Taking into account the variable recovery on a block by block basis, and the improved recovery once the HPGR circuit is commissioned, the reserve cut-off grade represents an insitu cut-off of approximately 0.356 g/t (on average for the life of the mine).

Two recovery calculations have been used in this study. Up until June, 2021 all ore comminution will be handled through the existing three-stage crushing plant. The three-stage crushing recovery factors are used in the NSR and recoverable grade calculations for all ore placed on the leach pad during this period. From July 2021 onwards all ore comminution will be handled through two stages of crushing followed by HPGR. The HPGR recovery factors are used in the NSR and recoverable grade calculation for all ore placed on the leach pad in this period.

The reference point at which Kışladağ’s mineral reserves are defined is the point where the ore is delivered to the processing facility.

15.2

Open Pit Optimization

15.2.1

Introduction

The open pit optimization was carried out using Minesight® mine planning software. A series of unsmoothed pit shells were created using a Lerchs-Grossmann algorithm with revenue factors declining from unity. The unsmoothed pit shells were then used as a guide for developing a detailed design to be used in production scheduling.

15.2.2

Economic Parameters applied to Mine Design

15.2.2.1

Metal Prices

Base case pit optimization metal prices were as follows:

●

Gold: US$1,250/ounce

●

Silver: US$16/ounce

15.2.2.2

Refining and Royalties

Gold will be refined on site and shipped as doré. The basis for pit optimization was the net mine gate revenue per tonne calculated for each block in the resource model. Metal prices described above and refining costs of US $3.80/ounce were used in the NSR value determination. The refining cost is based on averaging of historical costs. A silver credit of US$7.48/ounce of produced gold was applied based upon historic credits factored to a future silver price of US$16.00 per troy ounce.

NSR calculations have been conducted on a block by block basis in the block model. The calculations used allow for the accounting of:

●

Au grade thus taking into account the variability in the metal content of the deposit

●

Au metallurgical recovery for:

Three-stage crushing heap leaching for all ore to the leach pad from January 2020 to June 2021, inclusive

HPGR heap leaching for all ore to the leach pad from July 2021 onwards

●

Metal prices (silver as a credit)

●

Royalties

●

Refining charges

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15.2.2.3

Onsite Operating Costs and Increments

The onsite operating costs used for pit limit analysis include general and administration, processing and mining costs. The general and administration costs have been calculated at US1.74/t and US$1.66/t for the periods of three-stage crushing and HPGR, respectively. The processing operating costs from primary crushing to doré production were estimated to be US$5.00/t and US$5.08/t for three-stage crushed ore processing and HPGR ore processing, respectively. Sustaining capital costs for all ore processing was estimated to be US$0.55/t. Preliminary operating costs for mining ore and waste were US$1.42/t and US$1.47/t, respectively. Additionally sustaining capital costs of US$0.17/t and US$0.19/t were added for ore mining and waste mining, respectively. An incremental haulage cost of US$0.02/t/bench for ore where 150 ton class trucks are exclusively used and US$0.02/t/bench for waste was added for each 10 m bench below the open pit entrance at 960 masl.

Kışladağ is a mature mine with detailed historical costing data. To date more than 360 Mt have been excavated from the open pit and over 150 Mt have been leached. The operating and maintenance costs and performances for the mining, crushing and gold recovery operations are well understood and have been incorporated into the cost modelling used for the reserve estimate.

15.2.3

Metallurgical Parameters

15.2.3.1

Process Selection

The existing processing method at Kışladağ is primary crushing followed by secondary and tertiary crushing and conveying to heap leaching pads. Processing modifications will upgrade the existing infrastructure to replace the third stage of comminution with HPGR. The throughput rate will increase from 12.0 Mt per year to 12.6 Mt per year once the HPGR is commissioned. Cyanide cure will also be added to the circuit to improve the leach kinetics.

15.2.3.2

Process Recovery

The processing recovery used to develop the NSR models for mine planning was based upon column and IBRT test work.

Two recovery models were used to estimate recoverable gold and NSR in the block model. Both recoverable gold and block recovery averages were reported for mine planning and final recoveries in mine schedules carry back calculated recoveries using the recoverable grades. The back calculated average recovery of the mineral reserve is 56%.

Figure 15-1 shows the variability of recovery in the block model on a typical vertical section. The model shown is for the HPGR recovery.

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Figure 15-1: HPGR Ore Heap Leach Recovery Model on Section

On a block by block basis there is a wide distribution of recovery values ranging from a low of 23% to a high of 85%. The values however concentrate around the mean of 56% recovery with 93% of the ore blocks (based on contained gold) having recoveries falling within the range of 44% to 66%, and with the upper and lower tails each representing about 3.5% of the population. The distribution by alteration ore types and recovery within the ore blocks within open pit mine limits are shown in Figure 15-2 and Table 15-2. Averaging of the recovery is based upon the contained Au metal within ore blocks.

Figure 15-2: Distribution of Au Recovery by Alteration

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Table 15-2: Recovery Summary

Alteration	MZONE	Ore	Calculated Average Recovery
Alteration	MZONE	(%)	(R%)
ARG	10	9%	55%
POT	30	53%	55%
WMT	50	17%	57%
INT #3	60	3%	51%
FRB	70	18%	58%
Total		100%	56%

15.2.4

Block Model

15.2.4.1

General

The resource block model developed by the Eldorado technical group is described in Section 14. The block model and surfaces for topography and the geology were imported to a Minesight® mine planning model. The block model limits and block dimensions are shown in Table 15-3.

Table 15-3: Block Model Limits 2020

Parameter	Minimum	Maximum	Length (m)	Block Size (m)	No. of Blocks
X	686,335 E	688,455 E	2,120	20	106
Y	4,260,675 N	4,262,355 N	1,680	20	84
Z	0	1,130	1,130	10	113

Total number of blocks in the model is 1,006,152

Block model items transferred from the geology model for mine planning included estimated grade for gold, alteration, density, rock unit, metallurgical domains and recovery for leaching of both three-stage crushed ore and HPGR ore, as well as resource classification.

Additional items were populated in the Minesight® model for pit optimization and design purposes, as well as possible scheduling destinations, such as NSR and various wall slope codes.

15.2.4.2

Resource Classification

Resource Class: The mineral resource model includes measured, indicated, and inferred resources. Measured and indicated resources have been used to define the pit limits and for reporting of mineral reserves for scheduling. Inferred resources were not used in the mine plan.

Mining Recovery: Mining recovery is assumed to be 100%. No mining losses were applied to the ore reserves for the following reasons:

●

The deposit shows good lateral and vertical continuity at the cut-off grades applied for scheduling.

●

There is a broad width to the ore zones on individual benches.

●

A detailed grade control program has been implemented.

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Mining Dilution: Internal dilution was incorporated in the resource model by virtue of the compositing and interpolation method used to obtain the block grades. No additional dilution was applied in optimization.

15.2.5

Wall Slope Design

Inter-ramp wall slopes angles were assigned by sector that were further subdivided into “Pyroclastic Oxide” zone, “Pyroclastic Sulphide” and “Intrusives & Schist”. An additional zone was coded for the “Friable Zone” where single benching and reduced geotechnical berm spacing need to be implemented in the final design. The slope sector locations and design parameters applied for pit optimization and design presented in Table 15-4 and shown on Figure 15-3 and Figure 15-4.

Figure 15-3: Primary Slope Sector Locations

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Figure 15-4: Slope Coding in the Block Model

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Table 15-4: Slope Sector Parameters

MSEP Sector	SLOPC	Friable Code SLOP1	Design Sector SLOP2	STRAT	Oxide Sulphide	OXIDE	Bench Face Angle Pit Design Tool	Bench Height	Berm Spacing	Berm Interval	Berm Width BERM	Calc. Inter-ramp Angle	Geo-tech. Berm Width	Stack Height	Adjusted Inter-ramp Angle	Flat. Effect	Pit Opt. Input
NE	1		1	500	PYCL Oxide	10 to 20	65	10	2	20	7	50.8	20	80	43.2	7.6	39.4
			2	500	PYCL Sulphide	30	75	10	2	20	7	58.3	20	100	50.7	7.6	49
			3	60-400	INT-1, INT-2, INT-3, Schist	20-30	80	10	2	20	8	60	20	100	52.2	7.9	49
SE	2		4	500	PYCL Oxide	10 to 20	65	10	2	20	7	50.8	20	80	43.2	7.6	39.4
			5	500	PYCL Sulphide	30	75	10	2	20	7	58.3	20	100	50.7	7.6	47.5
			6	60-400	INT-1, INT-2, INT-3, Schist	20-30	80	10	2	20	8	60	20	100	52.2	7.9	47.5
SW	3		7	500	PYCL Oxide	10 to 20	65	10	2	20	7	50.8	20	80	43.2	7.6	39.4
			8	500	PYCL Sulphide	30	75	10	2	20	7	58.3	20	100	50.7	7.6	47
			9	60-400	INT-1, INT-2, INT-3, Schist	20-30	80	10	2	20	8	60	20	100	52.2	7.9	47
NW	4		10	500	PYCL Oxide	10 to 20	65	10	2	20	7	50.8	20	80	43.2	7.6	39.4
			11	500	PYCL Sulphide	30	75	10	2	20	7	58.3	20	100	50.7	7.6	45
			12	60-400	INT-1, INT-2, INT-3, Schist	20-30	80	10	2	20	8	60	20	100	52.2	7.9	45
	1 to 4	1	13	Friable		10 to 30	65	10	1	10	7	40.6	20	40	31	9.6	36
External			14				65	10	2	20	8	49.1	20		0	49.1	39.4
Fill			15				35	10	1	10	0	35	20		0	35	35

The Lithology (STRAT) codes are as follows:

●

60 is schist. It is a waste rock at the base and to the north of the main orebody.

●

100 (or anything in the 100’s) is Intrusive-1. This is the main ore bearing lithology and was the first of the intrusions.

●

200’s are Intrusive-2 (220 & 240). This is also mostly mineralized in the 0.2 to 1.0 g/t range and consists of two separate masses, only one of which reaches surface.

●

300’s are Intrusive-3 and similar Dykes (300, 330 & 350). This was probably the last intrusive and it is not well mineralized. Generally, the rock is more silicified and therefore has some better geotechnical properties.

●

400’s are Intrusive-2A. It is a partially mineralized mass in the south east and it has poor geotechnical properties. It grades in the half gram range.

●

500 is pyroclastics. They are mineralized at the Int-1 contact but beyond that are almost barren. They make up most of the waste rock.

15.2.6

Pit Limit Analysis

Unsmoothed pit limits were developed using a Minesight® variable slope Lerchs-Grossmann algorithm. The preliminary estimates of net mine gate revenue and operating costs were used to determine the value of each regular block in the model. A series of 30 nested pit limits were defined using revenue factors between 0.10 and 1.00. Those nested pit limits used to guide pit design are shown in Figure 15-5, Figure 15-6 and Figure 15-7. The highest NPV shell, pit shell 30, is shown as a thick blue line (HPGR model at a gold price of $1,250/ounce). The thick black line is a reference line representing the previous reserve pit.

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Figure 15-5: Bench Plan NSR & Lerchs-Grossmann Pit Limits

Figure 15-6: Cross Section Looking North

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Figure 15-7: Cross Section Looking West

The resources within the unsmoothed nested Lerchs-Grossmann pit limits are summarized in Table 15-5 at US$7.29/t NSR cut-off based on HPGR recoveries.

Pit optimization results are shown graphically in Figure 15-8 and Figure 15-9.

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Table 15-5: Lerchs-Grossmann in-Pit Resources

											Gold
Shell	Factor	Ore (kt)	Waste Total (kt)	Total (kt)	Strip Ratio (w:o)	Au (g/t)	Recoverable Au (g/t)	HPGR NSR (US$/t)	Calculated Recovery (%)	Years	Contained (oz x 1,000)	Recoverable (oz x 1,000)
1	0.1	34,326	1,932	36,258	0.06 : 1	0.94	0.54	21.11	57.3%	2.7	1,032	592
2	0.13	36,716	2,910	39,626	0.08 : 1	0.92	0.53	20.85	57.3%	2.9	1,091	624
3	0.16	40,537	4,561	45,098	0.11 : 1	0.90	0.52	20.37	57.3%	3.2	1,176	674
4	0.19	42,961	5,624	48,585	0.13 : 1	0.89	0.51	20.08	57.2%	3.4	1,232	704
5	0.22	47,045	8,322	55,367	0.18 : 1	0.88	0.50	19.72	57.3%	3.7	1,323	758
6	0.26	52,984	12,564	65,548	0.24 : 1	0.86	0.49	19.30	57.2%	4.2	1,458	835
7	0.29	55,339	14,260	69,599	0.26 : 1	0.85	0.49	19.11	57.2%	4.4	1,509	863
8	0.32	72,912	31,973	104,885	0.44 : 1	0.82	0.47	18.39	57.1%	5.8	1,918	1,095
9	0.35	74,885	34,040	108,925	0.45 : 1	0.81	0.47	18.31	57.1%	5.9	1,960	1,120
10	0.38	75,401	34,330	109,731	0.46 : 1	0.81	0.46	18.27	57.1%	6.0	1,971	1,125
11	0.41	83,521	42,890	126,411	0.51 : 1	0.80	0.46	17.98	56.8%	6.6	2,156	1,224
12	0.44	89,967	52,993	142,960	0.59 : 1	0.80	0.45	17.89	56.9%	7.1	2,308	1,313
13	0.47	90,495	53,564	144,059	0.59 : 1	0.80	0.45	17.87	56.8%	7.2	2,319	1,318
14	0.5	94,747	60,376	155,123	0.64 : 1	0.79	0.45	17.79	56.9%	7.5	2,416	1,374
15	0.53	99,172	64,544	163,716	0.65 : 1	0.79	0.45	17.60	56.7%	7.9	2,506	1,422
16	0.57	104,196	72,293	176,489	0.69 : 1	0.78	0.44	17.48	56.7%	8.3	2,616	1,484
17	0.6	105,239	73,251	178,490	0.70 : 1	0.78	0.44	17.43	56.7%	8.4	2,636	1,496
18	0.63	130,643	110,539	241,182	0.85 : 1	0.75	0.43	16.81	56.8%	10.4	3,159	1,794
19	0.66	130,973	110,854	241,827	0.85 : 1	0.75	0.43	16.80	56.6%	10.4	3,167	1,794
20	0.69	141,094	127,664	268,758	0.90 : 1	0.75	0.42	16.63	56.6%	11.2	3,380	1,914
21	0.72	172,444	173,344	345,788	1.01 : 1	0.72	0.41	16.07	56.4%	13.7	4,014	2,262
22	0.75	176,774	180,093	356,867	1.02 : 1	0.72	0.41	16.02	56.3%	14.0	4,109	2,313
23	0.78	178,589	182,487	361,076	1.02 : 1	0.72	0.41	15.99	56.2%	14.2	4,146	2,331
24	0.81	184,400	191,907	376,307	1.04 : 1	0.72	0.40	15.92	56.1%	14.6	4,269	2,395
25	0.84	187,392	197,998	385,390	1.06 : 1	0.72	0.40	15.90	56.1%	14.9	4,332	2,428
26	0.88	194,657	212,418	407,075	1.09 : 1	0.72	0.40	15.84	56.1%	15.4	4,487	2,516
27	0.91	195,607	214,518	410,125	1.10 : 1	0.72	0.40	15.83	56.1%	15.5	4,509	2,528
28	0.94	196,708	217,019	413,727	1.10 : 1	0.72	0.40	15.82	56.0%	15.6	4,528	2,536
29	0.97	203,820	232,635	436,455	1.14 : 1	0.72	0.40	15.76	55.9%	16.2	4,685	2,621
30	1	216,899	266,222	483,121	1.23 : 1	0.71	0.40	15.72	56.0%	17.2	4,972	2,782

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Figure 15-8: Pit Optimization Shells

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Figure 15-9: Pit Optimization Shells

15.3

Pit Design

The open pit Year End 2019 topography surface is shown below in Figure 15-10. Ore mining is currently active in the Phase 3 pit bottom. Active Phase 4 pushback on upper benches of the west, south and east sides of the pit is primarily in waste rock.

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Figure 15-10: Topography Surface Year End 2019

The final pit configuration plan view is shown in Figure 15-11. A section looking southwest through the Friable Zone is shown in Figure 15-12. The friable zone is a weaker geotechnical zone than all other zones. The design closely follows the Lerchs-Grossmann pit limits for PIT21 of the nested pit series. This pit limit was selected after detailed cash flow analysis of a high level schedule was completed. The conversion of ore tonnes from the unsmoothed pit to the final design was 100.4% with an 11.4% increase in reported waste.

The pit shell PIT21 captures 98.1% of the net present value of the maximum pit shell PIT30 but requires movement of only 71.6% of the total indicated tonnage and processing of 79.5% of the indicated leach pad feed.

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Figure 15-11: Final Pit Limits

Figure 15-12: Cross Section Looking Southwest

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15.4

Mineral Reserves

The mineral reserves for the deposit were estimated using a gold price of US$1,250/oz. The mineral reserves are reported using a 0.19 g/t recoverable Au three-stage crushing cut-off for ore that will use the three-stage crushing circuit and 0.19 g/t recoverable Au HPGR cut-off for ore that will be will use HPGR for final comminution, which commences in July 2021. The mineral reserves are constrained by the December 31, 2019 mined surface. The reference point at which Kışladağ’s mineral reserves are defined is the point where the ore is delivered to the processing facility. The proven and probable mineral reserves are 173.2 Mt with an average grade of 0.72 g/t Au. Mineral reserves effective January 17, 2020 are summarized in Table 15-6.

Table 15-6: Kışladağ, Mineral Reserves Effective January 17, 2020

Reserve Classification	Ore (t x 1,000)	Grade Au (g/t)	Contained Au (oz x 1,000)
Three-stage Crushing Ore (January 2020 to June 2021, inclusive)
Proven	15,942	0.9	464
Probable	513	0.91	15
Proven & Probable	16,455	0.9	479
HPGR Ore (July 2021 onwards)
Proven	148,589	0.71	3,387
Probable	8,131	0.55	144
Proven & Probable	156,720	0.7	3,531
Combined (Total Reserves)
Proven	164,531	0.73	3,851
Probable	8,644	0.57	159
Proven & Probable	173,175	0.72	4,010

Notes:

CIM Definition Standards (2014) were used for reporting the mineral reserves.

Mineral reserves are estimated based on the following assumptions: metal prices of $1,250/oz Au; cut-off of 0.19 g/t recoverable Au (equivalent to a NSR cut-off of $7.29/t); recovery is variable throughout the block model with average life of mine metallurgical recovery being 56% for all ore; and no dilution and mining recovery of 100% (both already accounted for in the resource block model).

The mineral reserve is derived from the measured and indicated mineral resources.

The mineral reserve estimation is constrained by the December 31, 2019 topo surface.

15.5

Risk Factors

The results of the economic analysis to support mineral reserves represent forward-looking information that is subject to a number of known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from those presented here. Uncertainty that may materially impact mineral reserve estimation include realized prices, market conditions, capital and operating cost estimates, foreign exchange rates, resource model performance, recoveries, and the timely and successful implementation of recommended actions.

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SECTION ● 16

Mining Methods

16.1

Introduction

The Kişladağ open pit mine currently provides ore at a rate of 12.0 Mtpa for three-stage crushing followed by heap leaching. In 2021, a HPGR will replace the tertiary crushers for final comminution. Once the HPGR is commissioned, the ore production rate will increase to 12.6 Mtpa. Annual mine production, inclusive of ore and waste, will peak at 42.5 Mtpa of total material movement in 2022, and will continue until 2034. The life of mine (LOM) stripping ratio is 1.12:1 based on a cut-off of 0.19 recoverable Au g/t. The major mining equipment at Kişladağ is summarized in Table 16-1.

Table 16-1: Major Mining Equipment

Description	Make	Model	Size	Units
Drilling
Wall Control drill	Atlas Copco	ROC L6	110 mm	1
Wall Control drill	Atlas Copco	ROC D65	110 mm	1
Blasthole drill	Atlas Copco	DM45	165 mm	2
Blasthole drill	Atlas Copco	PV-235-D	165 mm	3
Blasthole drill	Atlas Copco	PV-235-E	165 mm	2
Blasting
Blasting crew flatbed truck	M-Benz	Sprinter	110 kW	1
ANFO truck	M-Benz	3028K	280 kW	2
ANFO truck	M-Benz	3029K	210 kW	1
Emulsion truck	Volvo	420	315kW	1
Blasters crew truck	Nissan	Navara 4x4	140kW	1
Loading
Front end loader	Caterpillar	993K	12 m3	1
Front end loader	Le Tourneau	L-1350	21.4 m3	2
Hydraulic shovel	Hitachi	EX3600	21 m3	2
Hydraulic shovel	Hitachi	EX5600-6	29 m3	1
Hauling
Haul truck1	Caterpillar	785C/D	136 t	14
Haul truck	Hitachi	EH4000AC	219 t	10
Other mine operations
Track dozer	Caterpillar	D9T	306 kW	2
Track dozer	Caterpillar	D10T	433 kW	2
Wheel dozer	Caterpillar	834H	372 kW	1
Wheel dozer	Caterpillar	854K	597 kW	2
Motor grader	Caterpillar	16M	221 kW	4
Water truck	M-Benz	4140B	20,000L	3
Excavator	Caterpillar	330DL	200 kW	1
Excavator	Hitachi	ZX350LCH-3	202 kW	1

1 One Caterpillar 785C has been converted to a water truck – it is now interchangeable from water truck to haulage truck as needed.

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No additional primary production equipment will be needed and no complete replacements will be required for this reserve. The existing complement of equipment on a two shift per day roster produced 51.1 million tonnes of material movement in 2014. Since then the mobile equipment has been utilized at a lesser capacity to suit the needs. This has however demonstrated that the drilling, loading and auxiliary work capabilities of the existing equipment thoroughly meets the annual material movement requirements of the remaining reserve. As the pit gets deeper and parts of the SRD get higher, and as the longer haul to the NRD becomes part of the production requirement, the haulage cycle times will lengthen. Detailed haul profiles were calculated for the various production periods and haul truck requirements were determined accordingly. Material movements were constrained such that at no point in time do we require more haul trucks those currently in place. Equipment productivities were also used to predict equipment hours required by period. It is determined that ten of the Caterpillar 785 haul trucks will be nearing 150,000 hours at the time of retirement. The other four Caterpillar 785 haul trucks will be nearing 120,000 hours and all other equipment will be under 100,000 hours at the time of retirement. Based on historical availability records and the established maintenance standards, including certification of 5 Stars in Contamination Control by Caterpillar, these equipment hours will be manageable.

Ore and waste are mined on 10 m benches. Ore will be hauled to the primary crusher for processing and waste rock will be placed in the south rock dump (SRD) and the north rock dump (NRD). A total of 158.2 Mt of waste rock will be placed in the SRD as the primary location from 2020 until 2026. A total of 34.9 Mt of waste rock will be placed in the NRD over the mine life starting in 2022. Between 2022 and 2026 both rock dumps will be used as best suited.

The general arrangement of the open pit and SRD are shown in Figure 16-1.

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Figure 16-1: General Arrangement (photo taken early December 2019)

16.2

Mine Design

16.2.1

Geotechnical Wall Slope Design Sectors

As described in Section 15, there are four major slope design sectors that have been further subdivided according to lithology and oxidation state. A total of thirteen (13) design sectors have been coded into the mine block model. Bench face angles and berm width codes have been used to develop the final design. Specific geotechnical berm width input was used to over-ride the general design criteria. Key elements incorporated in the design are shown in Figure 16-2.

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Figure 16-2: Design Elements

16.2.2

Ultimate Pit Dimensions

The final pit dimensions are summarized in Table 16-2. The final depth of the pit will be to the 520 m bench with a final wall height of 565 m to the highest point on the pit rim.

The footprint of all remaining mining of the open pit is shown as the darkened areas in the aerial photograph in Figure 16-3. The pit dimensions are inclusive of already mined features in the south and west.

Table 16-2: Final Pit Dimensions

Final Pit		Dimensions/Elevations
East – West		1,650 m
North – South		1,300 m
Ramp Exit	Northeast exit (ore)	1,010 masl
	North exit (waste)	970 masl
	South exit (waste)	1,030 masl
Pit Bottom		520 masl
Maximum Highwall Crest Elevation		1,085 masl
Maximum Depth		565 m

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Figure 16-3: Footprint of Remaining Open Pit Production

16.2.3

Haulage Roads

The ultimate pit haulage road allowances have been designed for 35 m width at 10% grade. This width will provide adequate room for ditches, outside berm and travel width for 219 tonne trucks. In the “Friable Zone” haulage road width has been increased to 40 m to allow for potential bench face instability. Haulage road width was reduced to 25 m on the final 8 benches (tapering in from the 590 bench and extending to the bottom of the 520 bench) for single lane access during the last phase of ore recovery.

16.2.4

Phase Development

Mining has been designed in five phases with mining of the first two phases already completed. The current mine plan considers mining the Phase 3 pit until completion in 2022. Concurrently Phase 4 is being stripped back from 2020 to 2026. Phase 5 will start to be stripped back in 2021 and mining will continue in Phase 5 until the end of the mining part of the operation in 2034.

The phase designs have suited the need to manage stripping within the constraints of available equipment. Further analysis and design work will be conducted with the intent of improving the timing of gold production and, to a lesser extent, deferring some stripping by sinking Phase 5 quicker into the ore production areas. Bench plots of these phases are shown in Figure 16-4 and Figure 16-5.

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Figure 16-4: 800 m Bench Pit Phase Contours

Figure 16-5: 700 m Bench Pit Phase Contours

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Figure 16-6 shows the ore grades and recoveries, averaged by bench for the reserve. Tonnages have been broken down by phase to give a clear understanding of where the ore is situated.

Figure 16-6: Ore Qualities by Bench

16.3

Mine Production Schedule

16.3.1

Mining Plan

The LOM plan has been developed for a 15 year operation with reduced ore tonnages in the last two years. All ore will be extracted from a single open pit mine. The open pit has been phased to smooth the scheduling and improve the NPV. The pit bottom is currently in the lower benches of Phase 3 and capitalized waste stripping is occurring in Phase 4. Phase 3 will continue being the prime supplier of ore until the end of 2020. It will still supply ore until the end of Q2, 2022 however; from early 2021, Phase 4 becomes the prime supplier of ore. Phase 4 continues to supply ore until late 2026. From 2027 until the end of the mining life in 2034, all ore will come from Phase 5.

Capitalized waste stripping from Phase 4 and Phase 5 will occur from 2020 until 2025. During this period, the volume of material movement will be forcing a high demand on haul truck utilization. That tails off quickly corresponding to reduced stripping ratios slightly offset by longer haul routes from 2026 until the end of the pit operation in 2034.

A total of 173.2 Mt of ore will be processed during the 15 year LOM plan; 16.5 Mt of the mined ore will be processed using the existing crushing plant and the remaining 156.3 Mt of ore will be processed after converting the crushing plant to include HPGR.

The average strip ratio over the LOM is 1.12:1. A total of 193.2 Mt of waste will be mined, including 141.1 Mt of capitalized waste in Phase 4 and Phase 5. The average LOM gold grade is forecast to be 0.72 g/t Au. The recoverable grade is estimated to be 0.40 g/t for a LOM metallurgical recovery of 56%.

The mine schedule is shown in Figure 16-7 and Table 16-3.

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Figure 16-8 shows the number of trucks that will be required (utilized) each year of the remaining mine life. There are two sizes of trucks at site. These being Caterpillar 785, which is used for ore and waste as needed, and Hitachi EH4000, which are currently being used exclusively for waste haulage. The current number of trucks are sufficient for the material movement requirements of this reserve.

Figure 16-7: Mine Material Movement Schedule

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Figure 16-8: Haul Truck Requirement

Table 16-3: Mine Material Movement Schedule

Material	Units	2020	2021	2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	TOTAL
Ph3 Ore Mined	Kt	10,776	3,689	2,533	0	0	0	0	0	0	0	0	0	0	0	0	16,998
Au	g/t	1.02	0.95	0.87	0	0	0	0	0	0	0	0	0	0	0	0	0.98
Insitu Au Ounces	k oz	354	113	71	0	0	0	0	0	0	0	0	0	0	0	0	538
Recovery	%	53%	51%	51%	0%	0%	0%	0%	0%	0%	0%	0%	0%	0%	0%	0%	52%
Ph4 Ore Mined	Kt	1,220	6,923	9,545	10,979	8,556	2,493	619	0	0	0	0	0	0	0	0	40,335
Au	g/t	0.48	0.63	0.67	0.76	0.82	0.72	0.71	0	0	0	0	0	0	0	0	0.72
Insitu Au Ounces	k oz	19	139	205	270	224	58	14	0	0	0	0	0	0	0	0	929
Recovery	%	48%	51%	56%	60%	61%	53%	50%	0%	0%	0%	0%	0%	0%	0%	0%	57%
Ph5 Ore Mined	Kt	0	0	525	1,624	4,048	10,111	11,983	12,602	12,607	12,605	12,607	12,603	12,602	9,088	2,836	115,841
Au	g/t	0.00	0.51	0.47	0.56	0.60	0.58	0.62	0.67	0.73	0.73	0.77	0.74	0.71	0.63	0.63	0.68
Insitu Au Ounces	k oz	0	0	8	29	79	188	238	271	294	296	310	299	289	185	58	2,544
Recovery	%	0%	53%	52%	52%	52%	55%	59%	59%	57%	57%	55%	55%	56%	52%	54%	56%
Total Ore Mined	Kt	11,996	10,612	12,603	12,603	12,604	12,604	12,602	12,602	12,607	12,605	12,607	12,603	12,602	9,088	2,836	173,174
Au	g/t	0.97	0.74	0.70	0.74	0.75	0.61	0.62	0.67	0.73	0.73	0.77	0.74	0.71	0.63	0.63	0.72
Insitu Au Ounces	k oz	373	252	284	299	303	246	252	271	294	296	310	299	289	185	58	4,011
Recovery	%	52%	51%	55%	60%	59%	55%	58%	59%	57%	57%	55%	55%	56%	52%	54%	56%
Ph3 Waste Mined	Kt	3,094	709	203	0	0	0	0	0	0	0	0	0	0	0	0	4,006
Ph4 Waste Mined	Kt	26,959	16,421	3,971	860	1,159	1,138	517	0	0	0	0	0	0	0	0	51,025
Ph5 Waste Mined	Kt	0	12,946	25,708	25,749	24,919	22,235	9,911	4,026	2,336	1,595	2,319	2,563	1,626	1,833	388	138,154
Total Waste Mined	Kt	30,053	30,076	29,882	26,609	26,078	23,373	10,428	4,026	2,336	1,595	2,319	2,563	1,626	1,833	388	193,185
Total Material Mined	Kt	42,049	40,688	42,485	39,212	38,682	35,977	23,030	16,628	14,943	14,200	14,926	15,166	14,228	10,921	3,224	366,359
Strip ratio	w:o (t:t)	2.51 :1	2.83 :1	2.37 :1	2.11 :1	2.07 :1	1.85 :1	0.83 :1	0.32 :1	0.19 :1	0.13 :1	0.18 :1	0.20 :1	0.13 :1	0.20 :1	0.14 :1	1.12 : 1
Ph3 Bottom Bench	masl	720	690	660	-	-	-	-	-	-	-	-	-	-	-	-	660
Ph4 Bottom Bench	masl	900	850	810	770	710	670	640	-	-	-	-	-	-	-	-	640
Ph5 Bottom Bench	masl	-	980	930	890	840	800	770	750	730	710	680	650	620	570	520	520

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16.3.2

Annual Mining Plans

Annual mine plans were created to reflect the advance of mine development. Selected plans are presented in the figures below.

From the start of 2020 until the end of June, 2021 all ore from the open pit will be processed using existing equipment. The pit configuration at the end of June 2021 is shown in Figure 16-9. By this time stripping of Phase 4 has progressed to bench 870 and Phase 5 stripping to bench 1030, while ore is still being extracted from bench 710 of Phase 3.

Figure 16-9: Mine Development Mid-Year 2021

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Mine development to year end 2023 is shown in Figure 16-10. Phase 3 has terminated at Bench 660. Phase 4 has progressed to Bench 770 and Phase 5 has been opened up to Bench 890.

Figure 16-10: Mine Development 2023

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Mine development to year end 2027 is shown in Figure 16-11. Phase 4 has terminated at Bench 640 (in Q2-2026), and Phase 5 has advanced to Bench 750. All mining after this point is within Phase 5 with the very low strip ratio of 0.17 tonnes of waste per tonne of ore. Mining progresses bench by bench at 12.6 Mt of ore per year from 2028 to 2032. Thereafter the advance is constrained by five benches per year, which resultantly constrains the rate of ore extraction to less than 12.6 Mtpa.

Figure 16-11: Mine Development 2027

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The final pit configuration is shown in Figure 16-12.

Figure 16-12: Mine Development 2034

16.3.3

Waste Rock Management

The current SRD can accommodate 158.2 Mt of additional material from December 31, 2019. The north rock dump location has capacity far in excess of the remaining 34.9 Mt of waste that will be extracted with the ore reserves. A 110 million tonne capacity design from earlier work on the milling option (Eldorado, 2018), as shown in the general layout in Section 18, was conservatively used for calculating haulage profiles for costing and productivity estimation. Optimization work will commence to improve the scheduling of waste material movement, which may change the final size and/or configuration of the individual dumps while maintaining the same overall capacity. The current configuration of the SRD is shown in Figure 16-13. The general layout figure in Section 18 also shows the final shape and size of the SRD, and the location of these dumps with respect to the open pit.

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Figure 16-13: Current Waste Dump Configuration (as of December 31, 2019)

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SECTION ● 17

Recovery Methods

17.1

General Description

Kişladağ is an open pit mine and heap leach operation with a 3-stage crushing plant. The process plant will be modified with the construction of an HPGR unit to replace the tertiary crushing circuit. Concentrated cyanide solution will be added to the crushed ore prior to its stacking. Subject to further investigation, the crushed ore may be agglomerated to enhance percolation in the heap. The upgraded plant will maintain its current capacity to process 12.6 Mt per year, producing approximately 160,000 ounces of gold annually.

17.2

Recovery Methods

A previous technical report entitled “Technical Report for the Kişladağ Gold Mine, Turkey”, published by Eldorado Gold in March 2010, was based on the ore being processed in a conventional heap leach facility which consists of a three-stage crushing plant, an overland conveyor from crushing plant to heap leach pad, mobile conveyors, a radial stacker for placing the crushed ore onto leach pad, and a carbon adsorption facility for recovering dissolved gold onto activated carbon. The gold-loaded carbon is then stripped on site in a refinery and the final product is a gold doré bar.

The initial production capacity was 5 Mtpa for the first two years of operation. Predominantly oxide material was processed during this period. The facilities were then expanded to process 10 Mtpa and subsequently to 12.5 Mtpa. As mining progressed deeper into the pit, the proportion of sulphide ore has increased and become dominant. Since 2016, oxide material quantities have become negligible. Typical crushed product size after three stages of crushing is 80% passing 6.5 mm.

The existing process plant consists of:

●

Primary crushing and coarse ore stockpile

●

Secondary screening and crushing

●

Tertiary crushing and screening

●

Crushed ore overland conveying and stacking

●

Heap leaching

●

Adsorption, desorption, regeneration (ADR) plant

●

Electrowinning and gold smelting

●

Reagent and air services

●

Water services (fresh water, process water, potable water)

Subsequent evaluation for treatment of sulphide ore through heap leach process necessitated changes in the process flowsheet to improve gold recovery over the remaining Kişladağ ore body. Consequently, in 2018 and 2019, testwork was undertaken to identify if alternative crushing methods and cyanide cure would improve overall leaching recoveries. The resulting crushing and recovery testwork program serve as the basis for the current evaluation.

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17.3

Process Selection

Subsequent changes in leaching characteristics of the sulphide ore at depth necessitated changes in the process to achieve satisfactory gold recovery for the remaining mineralization in the deposit.

Cyanide leach testwork using HPGR crushed materials showed a modest increase in gold recoveries. An internal scoping study was started in November 2019 to evaluate the technical and economic viability of incorporating an HPGR to the existing crushing circuit. Metallurgical testwork outlined in Section 13 was used as the basis to support the revised recovery values to support the economic evaluation.

The product from the secondary crusher will report to a new HPGR circuit. The HPGR circuit effectively replaces the tertiary crushing and screening system and greatly simplifies the overall materials handling system. The HPGR product will be separated into two streams, namely the center cut product and the edge product. The center cut product will report to the overland conveying system for transfer to the heap leach facility. The edge product will be recycled back to the HPGR feed for a second pass of crushing.

The HPGR circuit will be installed within the footprint of the existing crushing circuit and utilize parts of the existing conveying systems with minor modifications. New structures and equipment will be erected in free spaces of the crushing circuit area to minimize disruption to production during construction.

A simplified flowsheet of the revised flowsheet is depicted in Figure 17-1.

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Figure 17-1: Kişladağ Revised Flow Sheet

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17.4

Plant Design Basis

The modified process plant was designed on the basis of overall plant operating time of 80% and 365 days per year for a total operating time of 7,008 hours per annum. The process plant has been designed to produce up to approximately 160,000 oz/a gold as doré bar although historically, the ADR has produced in excess of 300,000 ounces annually on occasion.

Key criteria selected for the modified process plant design are:

●

Annual ore throughput of 12.6 Mtpa

●

Replacement of tertiary crushing system by HPGR circuit

●

Addition of cyanide cure circuit

●

Recirculation of HPGR edge product

●

Final product size will be controlled by the amount of HPGR crushing that can be utilized but a P80 of 6.5mm is what is currently being generated and would represent a minimum size.

●

Plant operating time of 80%

●

Average ore head grade of 0.72 g/t for gold

●

LOM average gold recovery of 56%

17.5

Process Description

17.5.1

Crushing

17.5.1.1

Primary Crushing

The primary crusher is a 1,270 mm by 1,651 mm gyratory crusher capable of processing up to 2,000 t/h. Run-of-mine ore is hauled from the open pit and directly dumped into the primary crusher. Initially, contract miners were hired to deliver the ore in 37-tonne trucks and the layout provided for two trucks dumping simultaneously. Owner operated mining training and transition with Caterpillar 136-tonne haul trucks started in May 2008. In October 2008, the mine started to use its own equipment. The dump pocket has a capacity of 300 tonnes and is suitable for direct tipping by the Caterpillar haul trucks. The crushed ore is conveyed to a 20,000-tonne coarse ore stockpile. This is used as surge capacity to feed the secondary crushing circuit. The material is then conveyed to the secondary crushing circuit.

17.5.1.2

Secondary Screening and Crushing

The secondary crushing system operates as an open circuit where the crushed product from primary crushing is reclaimed by two apron feeders onto a conveyor belt feeding a 2.4 x 6.4 m double deck screen. The screen undersize product of less than 65 mm) is directed to the HPGR circuit and the screen oversize is conveyed to a 200-tonne secondary crusher feed bin. The aperture of the bottom deck of the screen will be adjusted so that the secondary crusher generates a product suitable as feed for the HPGR.

The bin is discharged by an apron feeder onto a conveyor belt feeding an MP1000 standard secondary cone crusher.

The top size of the product from the secondary crushing will be compatible with the allowed maximum feed size for the HPGR.

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17.5.1.3

Tertiary Screening and Crushing (HPGR)

The secondary screen undersize and secondary crusher product will combine and then feed the HPGR via an existing conveyor. The center cut product of the HPGR is directed to the overland crushing system. The edge product of the HPGR is recycled via a new set of conveyors back to the HPGR for a second pass through the system. Recirculating load can be adjusted in the field and the circuit will be designed to handle up to 100% recirculating load.

17.5.2

Overland Stacking

The HPGR center cut product is conveyed from the crushing plant to the heap leach pad via overland conveyors. A bypass is allowed to divert to a 20,000 tonne stockpile to allow maintenance for overland conveyors. The stockpile has a reclaim system feeding back onto the first overland conveyor. Material is dozed as required to feed the reclaim system.

The last overland conveyor features a mobile belt tripper that can move along its length which follows the width of the cells at the bottom of the heap leach pad. The belt tripper feeds the portable ramp conveyors to convey up the side slope of the heap leach pad. Once onto the top of the leach pad, the last portable ramp conveyor feeds a series of portable grasshoppers which load a sequence of conveyors: feed/loader conveyor, horizontal conveyor, and finally the radial stacking conveyor. The radial stacker places the ore in 10-m high lifts in an 80 m wide sweep, the width of one cell. Concentrated cyanide solution is added to the crushed ore before stacking.

Installation of interlift liners started in 2018 in order to shorten the distance of pregnant solution to flow out, to reduce dilution to the pregnant solution, and to decrease inventory of dissolved gold. The interlift liners are constructed using high-density polyethylene membranes and direct the pregnant solution to a system of collection pipes equipped with flowmeters.

The existing south heap leach pad is a series of 30 cells each 80 m wide by 800 to 1,000 m long with a total surface area of approximately 1.0 km x 2.4 km. The south heap leach pad has a maximum capacity of 234 Mt and is planned to be in service until 2029, when this maximum capacity is reached. A new north heap leach pad will be constructed directly north of the existing heap leach pad to provide additional 90 Mt of stacked ore capacity.

17.5.3

Irrigation and Water Management

The leach pad is irrigated by a series of pumps, pipes, control valves and drip emitters to feed dilute cyanide solution from the barren pond to the top surface of the heap leach pad. Solution application rate ranges from approximately 5.0 to 6.5 L/h/m2.

There are three main process ponds to contain the heap leach solutions: the pregnant solution pond, barren solution pond, and intermediate solution pond. Live capacities of these ponds are 8,600 m3, 17,400 m3, and 7,200 m3 respectively. A spare fourth pond of 8,600 m3 is available as well. All ponds are lined with double HDPE membrane and fitted with leak detection pipes and pumps. Ponds are covered with floating HDPE plastic balls to reduce evaporation and to ensure wildlife do not drink the cyanide solution.

The water management system has been designed to accommodate a 100 year, 24 hour storm event. An 116,000 m3 storm water event pond lined with HDPE membrane is provided to contain excess overflow solution from the pregnant solution pond. A second storm water event pond of 98,000 m3 capacity is also available to cater for the storm event that exceeds the one in a 100 year estimate. An emergency hydrogen peroxide detoxification circuit is installed to reduce cyanide content of any overflowing solution containing cyanide to a safe level, in the event that discharge into environment becomes absolutely necessary.

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17.5.4

Gold Adsorption and Smelting

The gold adsorption facility (ADR plant) consists of five trains of carbon columns with each train consisting of five columns. Gold from the pregnant solutions is loaded onto the activated carbon and then the carbon is removed periodically for recovery of gold. The gold is stripped from the carbon in a standard Zadra process consisting of pressure stripping, followed by electrowinning, and smelting. The final product is a gold doré bar suitable for final refining to 99.999% purity in by a domestic refinery.

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SECTION ● 18

Project Infrastructure

18.1

Site Location

The Project is located on the western edge of the Anatolian Plateau at an elevation of approximately 1,000 m. Local elevations range from the peak of Kişla Dağ (Kişla Mountain) at 1,300 masl, to the valley base at 700 masl. The crushing plant UTM location is N 4262000 and E 688200, approximately 2 km north of the village of Gümüşkol.

18.2

Site Infrastructure

The project area at the end of mine life is presented on Figure 18-1.

18.2.1

Plant

The crushing plant is adjacent to the open pit and the administration buildings are located on ground level between the pit and the crushing plant.

18.2.2

Heap Leach Facilities

The south leach-pad is located on the western toe of Kişla Dağ and bounded on the west side by the main basin drainage course. The south leach pad facility starts 300 m north of the crusher plant and extends northwards, approximately 2.4 km.

Construction of the north leach pad facility will be similar to the existing heap leach facility. The liner system will consist of a compacted clay bed and high-density polyethylene (HDPE) liner to form a double liner. The south heap leach pad has a capacity of 234 Mt with life-of-mine footprint of 2.32 km2. The north heap leach pad is already permitted with a 366 Mt capacity and life-of-mine footprint of 3.63 km2. The two heap leach facilities can accommodate the total reserve tonnage.

18.2.3

Rock Disposal Site

The SRD is centered about 1 km southwest of the open pit, within the headwater area of a small valley drained by an intermittent stream. The rock dump holds approximately 210 Mt of waste rock with additional permitted capacity of 158 Mt. Studies have shown capacity can be increased within the permit boundaries. Other construction will require use of a small portion of the waste rock. Concept studies are ongoing to potentially expand capacity of the SRD.

A new NRD on the mountain west of the leach is planned to hold the 34.9 Mt of excess waste rock in the mine plan. Designs for a one billion tonne capacity facility were completed previously with the footprint incorporated in the approved expansion to the EIA boundary.

The total capacity of the NRD design (capacity 110 Mt) and with the remaining SRD (capacity 158.2 Mt) holds waste rock in excess of the current mine plan requirements.

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Figure 18-1: Project Area

18.2.4

Site Security

A 2 m high fence has been installed along the property boundary, which controls access to the mine site. There is one main access gate with a gatehouse and weighbridge, and a secondary gatehouse near the ancillary buildings, both are manned 24 hours per day. Additional security fencing is erected around the ADR plant and solution ponds, electrical substations, reagent facilities, and explosives storage areas. Additional fencing will be added to encompass the north leach pad facility during construction.

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18.2.5

Access Road

The access road is a paved road approximately 5.3 km long, 10 m wide connecting the mine site to the regional road from Ulubey to Esme. The upper portion of the road east of the pit was relocated to accommodate a larger pit perimeter in 2018.

A portion of a village road connecting the villages of Gümüşkol and Katrancilar was previously replaced by a new road, approximately 1.9 km long with an overpass above a haul road to the leach pads and future north rock dump. The bypass was constructed to bypass crushing facility expansion; the overpass is sufficient to allow two-way traffic of large (219 tonne) mining trucks on the haul road beneath.

18.2.6

Water Supply

Fresh water for the project is supplied from a well field located approximately 13 km to the east of the plant site, in Neogene limestones. Five wells are in operation with a pipe line to site. Two water storage tanks and an underground distribution system at site provide the capacity for process and non-potable water requirements; a portion of the tank capacities are dedicated to fire protection.

In 2017 a new dam (Gedikler Dam) was constructed by Tüprag in conjunction with the General Directorate of State Hydraulic Works (DSI) approximately 6 km south south-west of the process plant. Kışladağ is allotted a portion of the annual collection from the dam to supplement the well systems. A pumping station and pipe line between the dam and the non-contact water collection ponds adjacent to the heap leach pad are in operation as required.

18.2.7

Power Supply

Two power lines distribute power to site. Turkish Electricity Transmission Corporation (TEIAS) distributes electrical power to the Kışladağ site with a 27.7 km long 154 kV transmission line from the Uşak industrial zone. The main substation at site is rated at 100 MVA with three 50 MVA 154 / 34.5 kV power transformers (one cold spare). The substation currently has abundant spare capacity and will require only minor modifications to accommodate the additional load associated with the HPGR installation. Site distribution is at 34.5 kV, and in local areas at 6.6 kV, 4.16 kV or 0.4 kV, which is distributed locally via overhead power lines and underground cables in process areas and near buildings. The original power line from the Turkish Electricity Distribution Corporation (TEDAS), distributes electrical power to the Kışladağ site with a 25 km long 34.5 kV transmission line from the Uşak industrial zone, to a small portion of the facilities adjacent to its alignment.

18.2.8

Ancillary Buildings

The permanent mine buildings were designed and constructed by local Turkish contractors. The architecture of the facilities includes local building materials and methods compatible with the surrounding infrastructure.

18.2.8.1

Warehouse

The original workshop/warehouse (760 m2) was constructed with the first phase of the crushing plant in 2006. Subsequently a separate maintenance workshop was built and a portion of the facility was converted to increased warehouse space, control room, and department offices. An adjacent outdoor fenced area together with covered area has been constructed for storage of large equipment and miscellaneous reagents. A diesel depot for dispensing fuel to small vehicles is installed near the warehouse. A new fabric covered warehouse was constructed in 2016 and provides 4,000 m2 of additional covered storage with added outdoor fenced storage near the crushing plant.

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18.2.8.2

Maintenance Workshop

A maintenance workshop (780 m2) with overhead crane services the site and includes an electrical workshop, an instrument workshop, tool storage, a security store, offices, storage space for maintenance items, washroom, locker and change rooms.. Outside paved areas have been provided for work areas and additional storage.

18.2.8.3

Mine Truckshop

The mine truckshop complex (1,850 m2) services the fleet of mining trucks. An addition was added in 2015 to accommodate the 219 tonne truck fleet. The complex includes five indoor heavy equipment repair bays equipped with an overhead traveling crane, a covered outdoor service bay, and an outdoor wash bay equipped with an oil/water separator. A general repair area and a welding shop are included in the complex. A three-story annex to the truckshop houses a mechanical room and office space. In addition there is another two-storey annex for change rooms and washrooms.

18.2.8.4

Administration Building

The administration building (400 m2) is a single story building and includes general areas for engineering, geology and administration personnel plus seven individual offices for management personnel.

18.2.8.5

Mine Dry and Canteen

The mine dry and canteen (540 m2) is a single story concrete building. The canteen is equipped with a kitchen area and a seating area for 72 people and includes a covered, enclosed patio that can seat a further 60 diners. Washrooms, shower facilities, and clean and dirty lockers are provided in the mine dry area. Additional space is provided for five offices and a meeting room.

18.2.8.6

Assay Laboratory Building

The assay laboratory building (440 m2) houses the assay laboratory rooms, assayers and assistants’ offices, washrooms for personnel, and storage rooms. The assay laboratory is capable of handling 550 samples per day. It includes sample preparation, acid digestion, atomic absorption finish, fire assay, and a wet laboratory. An additional metallurgical laboratory area has been added to support additional testing inside the cyanide storage facility.

18.2.8.7

Health and Security Building

The health and security building (86 m2) provides three consulting and treatment rooms for the mine’s doctor, toilets, and separate attached office for the security contractor’s manager.

18.2.8.8

Environment and Safety Building

The environment and safety building (470 m2) provides office accommodation, meeting room, small laboratory, toilets and tearoom facilities for environment, safety and process engineering management personnel.

18.2.8.9

Public Relations Building

At the main entrance gate a public relations building (450 m2) provides office accommodation, reception room, toilets, and tearoom facilities.

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18.2.8.10

Gate House

The gatehouse controlling the main entrance and weigh bridge also provides meeting rooms, washroom, and locker space for security personnel.

18.2.8.11

Miscellaneous Prefabricated Buildings

Miscellaneous prefabricated buildings provide additional office accommodation, washroom facilities, storage rooms, and work areas for the construction management team, geology core logging and storage, safety rescue and demonstration, laundry, archives, and a prayer room.

Operations personnel reside in the surrounding towns and villages and there are no plans to erect a permanent camp for operations personnel or temporary construction camps. Personnel are transported to the site by buses.

18.2.8.12

Sewage

Sewage systems on site include an underground sewer reticulation system, which connects all the buildings to a treatment plant, with a capacity of 120 m3/d.

18.3

Water Management

18.3.1

Water Collection and Treatment

The site is bounded by a series of collection ditches to divert non-contact water around the site to reduce the volume of contact water.

All contact water is collected from the mine site and pit inflows and sent to collection ponds at the treatment plant. The treatment plant is located north of the existing ADR plant with a capacity of 625 m3/hr.

On site, there are numerous ponds to collect process streams (barren and pregnant solutions at the ADR plant), contact water, non-contact water, and surge ponds for storm events. The ponds were sized based on a 100-year storm event with additional capacities for storage and process surges.

18.3.2

Water Requirements

The water balance for the heap leaching process requires an inflow of approximately 405 m3/hr. Water for the process will come from the fresh water systems including the well fields and Gedikler dam; and will be supplemented with water from the treatment plants and contact water ponds (Table 18-1). A full water balance will be conducted during the next phase but based on the current systems and operating conditions there are no additional water requirements beyond the existing infrastructure.

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Table 18-1: Water Sources Available

	Dry Season Estimates
Water Sources	L/s	m3/hr
Water Wells	78.7	283.3
Gedikler Dam	28.0	100.8
Open Pit Dewatering and Water Collection Ponds	30.6	110.2
Totals:	137	494

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SECTION ● 19

Market Studies and Contracts

19.1

Markets

19.1.1

Market Studies

Eldorado is currently selling gold from the Kişladağ operation; hence, Eldorado has not performed any formalized marketing studies in respect to future Kişladağ gold production. Gold is currently sold on spot market via Turkish refiners by Tüprag’s internal sales department. During 2019 Kişladağ sold gold at an average realized selling price of US$1,417 per troy ounce.

As per the new mining code put into effect in August 2017, the Turkish Central Bank has the right to purchase all gold produced at the site at LME spot prices.

19.1.2

Price

The price of gold is the largest single factor in determining profitability and cash flow from operations. Therefore, the financial performance of the project has been, and is expected to continue to be, closely linked to the price of gold. Reserves have been determined at a gold price of US$1,250 per troy ounce.

19.2

Contracts

Kişladağ has no contracts for gold sales or hedging in place. Gold is sold at spot price. Currently Kişladağ has contracts and purchase agreements that are in place including cyanide, diesel, and explosives and leasing of forestry land; and service contracts for security and catering.

19.3

Taxes

The current corporate taxation for Turkish businesses is 22% through the year of 2020 and reduces to 20% in the year 2021 and thereafter. Depreciation is based mostly on a unit-of-production calculation in international financial reporting standards (IFRS) recording. Turkish lira depreciation is based on government’s depreciation list and this is mainly 10% for mine assets.

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SECTION ● 20

Environmental Studies, Permitting and Social or Community Impact

20.1

Baseline Conditions

Initial baseline studies were been performed relating to Kışladağ, between 2000 and 2002, the combination of these studies defines the initial environmental and socioeconomic baseline conditions of the study area.

20.2

Environmental Considerations

The Kişladağ Project Environmental Impact Assessment (EIA) study was completed in January 2003 and submitted to the Turkish Authorities at the Ministry of Forest and Environment. An Environmental Positive Certificate for the project was subsequently obtained in June 2003. The EIA document presents the baseline conditions and socio-economic effects associated with the development of the Project, and defines the features and measures to mitigate potential impacts.

The EIA considered the potential impact on the local and regional environment as it relates to the following project areas including:

●

Open pit workings

●

Waste rock impoundment

●

Process plant

●

Heap leaching facility and solution management

●

Infrastructure necessary for the Project's operation

An Environmental Management Plan (EMP) was developed to address the potential impacts of the mining operation addressed in the EIA and additional issues. This plan was put in place prior to pre-production mining starting in 2005 and has been maintained throughout the production phase. The scope of the monitoring program within this plan includes elements of air quality, surface water and ground water monitoring, noise, blast vibration, flora and fauna, social-economy, as well as waste and hazardous waste storage and disposal. Data outlined in the monitoring program has been collected on a monthly basis since the implementation and reported to the relevant government agencies on a quarterly and annual basis.

Tüprag applied and received subsequent EIA amendments in 2011 and 2014 to increase the Kişladağ operations throughput to 12.5 Mtpa and 35 Mtpa respectively.

ENCON Environmental Consultancy Co. authored the full EIA submitted in March 2003 and the subsequent amendments.

20.3

Social Impact

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SECTION ● 21

Capital and Operating Costs

21.1

Introduction

The currency exchange rates used are as per Q4 2019 market conditions. All costs are presented in US Dollars (US$) based on the exchange rates shown in Table 21-1.

Table 21-1: Exchange Rates

Currency Code	Currency Name	Exchange Rate
US$	United States Dollar	US$1.00 = US$1.00
₺	Turkish Lira	US$1.00 = ₺6.20
CAN$	Canadian Dollar	US$1.00 = CAN$1.30
€	Euro	US$1.00 = €0.87

21.2

Growth Capital Costs

Mining costs at Kişladağ are very well understood, and as such, actual productivities and costs were used as the basis of the mining costs. Direct costs for the HPGR construction were developed from a combination of budget quotations, ongoing contract rates, Q3 2018 contractor survey rates, in-house data, historical benchmarks, and material take-offs. Indirect costs and owner’s costs were estimated in accordance with the project execution strategy, first principles calculations, relying on historical data at Kişladağ, and allowances. Contingency was calculated based on the level of project definition by discipline.

Table 21-2: Growth Capital Cost Summary

Area	Growth Capital (US$ x 1,000)
Waste Stripping	254,970
HPGR	13,468
Other Plant Equipment and Platework	4,577
Civil Site Services	25
Concrete and Structural Steel	2,791
Electrical & Instrumentation	826
Sub-Total Direct Costs	276,657
Indirects and Owners	4,504
EPCM	2,386
Contingency	7,253
Total Installed Cost	290,800

In Table 21-2, everything other than the Waste stripping of $ 255 M is associated with the purchase and installation of the HPGR. The costs of the HPGR project in isolation is $ 35.8 M.

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21.2.1

Capitalized Waste Stripping

Capitalized waste stripping schedule is in accordance with the mining methods described in Section 16.

Kışladağ is an operating mine with well-known mining costs. Load and haul productivities and fuel burn rates were calculated on planned haulage profiles, and all other cost factors used actual costs data as a basis for the future cost estimates.

21.2.2

Basis of Estimate (HPGR)

The accuracies of the cost estimates are consistent with the standards outlined by the Association for the Advancement of Cost Engineering (AACE). The cost estimate for the HPGR construction is a prefeasibility-level estimate categorized as AACE Class 4.

21.2.2.1

Labour

Labour rates were developed based on escalated contractor survey data completed in Q4 2018. The all-in crew labour rates include all direct and indirect costs associated with the contractors. Mobilization and de-mobilization costs are captured separately from the labour rates in the indirect costs.

A labour productivity factor is used to account for the overall labour force efficiency. Non-productive time is estimated based on the expected construction conditions. The overall productivity factor was determined to be 1.30 for concrete and steel, and 1.80 for mechanical, piping, electrical, instrumentation (MPEI), based on historical contractor survey data.

21.2.2.2

Materials

Direct unit costs for bulk materials were based on contract rates and quotations from recent projects. For minor items, allowances were carried based on benchmarked historical data. Table 21-3 outlines the primary source of unit costs, by discipline.

Table 21-3: Primary Source for Unit Costs

Commodity	Primary Source
Earthworks	On-going Contract Rates
Concrete	Escalated Q3 2018 Quotations
Steel	Escalated Q3 2018 Quotations
Process & Ancillary Equipment	HPGR – Budget Quotations (three vendors) Conveyors – In-house Data Minor Equipment – In-house Data and Allowances
Platework	Escalated Q3 2018 Quotations
Piping	Allowance
Electrical	Equipment – Budget Quotation, In-house Data Bulk Materials - Allowance
Instrumentation	Allowance

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21.2.2.3

Material Quantities

Quantities were based on preliminary design material take-offs and equipment lists, with allowances for minor items that are not substantial in cost. Table 21-4 summarizes the primary source of quantities by discipline.

Table 21-4: Primary Source of Quantities

Commodity	Primary Source
Concrete	Material take-offs based on dimensions from general arrangement drawings and experience from similar projects, with design consideration for region’s earthquake zone status
Steel	Material take-offs based on dimensions from general arrangement drawings and experience from similar projects, with design consideration for region’s earthquake zone status
Process & Ancillary Equipment	Mechanical Equipment List
Platework	Material take-offs based on dimensions, previous similar designs, and quantity factors
Piping	Allowance
Electrical	Electrical Equipment List Bulk Materials – Allowance
Instrumentation	Allowance

21.2.2.4

Indirect Cost Estimate

Indirect costs were primarily factored from direct costs based on historical experience. Table 21-5 summarizes the basis of indirect costs.

Table 21-5: Basis of Indirect Costs

Area	Primary Source
Q – Construction Indirects	Field indirect costs were allowances based on historical experience, revised for site specific assumptions.
	Freight and logistics costs were calculated as a percentage of equipment and material supply cost.
	Vendor representatives’ costs were calculated based on vendor provided information.
R – Spares/First Fills	Capital and commissioning spares estimated based on vendor provided information.
R – Spares/First Fills	Equipment first fills were factored from the plant equipment supply costs
S – Engineering, Procurement, Construction Management (EPCM)	EPCM costs were factored based on the total direct costs.

21.2.2.5

Contingency

Contingency was calculated at a summary level by each discipline. A percentage was applied for each discipline depending on the level of project definition and reliability of the cost information.

Contingency for the capital cost estimate was calculated to be 25% of direct and indirect costs.

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21.2.2.6

Exclusions

The following costs and scope were excluded from the capital cost estimates:

●

All facilities not identified in the summary description of the project

●

Fees or royalties relating to use of certain technologies or processes

●

Force majeure

●

Working capital other than capital and commissioning spares as well as first fills

●

Operating spares

●

Financing charges and interest

●

Currency exchange fluctuations

●

Recoverable value added taxes (VAT)

●

Sunk costs

●

Escalation

21.3

Sustaining Capital Costs

The basis of the sustaining capital cost estimates (Table 21-6) vary based on the nature of the costs.

Kişladağ has been in operation since 2006. Costs associated with earthworks, leach pad construction, and maintenance of capital equipment are very well understood. Ongoing contract rates, actual costs data, and other in-house data have been utilized as the basis of large majority of sustaining capital costs (Table 21-7).

Table 21-6: Sustaining Capital Cost Summary

Area	Sustaining Capital (US$ x 1,000)
Mining Equipment Rebuild	58,379
Miscellaneous Mining Capital	3,030
North Rock Dump	5,117
North Leach Pad	116,976
Interlift Liners	25,260
Other Process Capital	32,487
General and Administrative	1,101
Other Sustaining Construction	2,898
Total Cost	245,247

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Table 21-7: Basis of Sustaining Capital

Commodity	Primary Source
Mining Equipment Rebuild	Quantities: Maintenance schedule was updated for the life-of-mine fleet demand calculated for the mine plan described in Section 16. Cost: Rebuild/replace costs of mobile equipment fleet at Kışladağ are well understood. Quotations and actual costs data were used as the basis for the cost estimates.
Miscellaneous Mining Capital	Quantities: Miscellaneous mining capital costs include testwork, consultancy costs, and small miscellaneous purchases. Cost: Costs are based on a combination of quotations, in-house data, and allowances.
North Rock Dump	Quantities: Material take-offs for the north rock dump, ponds, and associated roads were scaled based on preliminary design for 110Mt capacity completed for a past study. Cost: Unit rates for earthworks related to the rock dump and ponds were based on ongoing contract rates and in-house data. An allowance for the access road to the north rock dump was factored based on total length.
North Leach Pad	Quantities: Material take-offs for the north leach pad, ponds, and associated infrastructure were based on preliminary design for 150Mt capacity completed for a past study. Cost: Unit rates for earthworks, liners, and heap leach infrastructure are based on ongoing current contract rates and in-house data based on work completed at the currently operating South Leach Pad.
Interlift Liners	Quantities: Material take-offs for the interlift liners are based on the remaining area of South Leach Pad to be lined, and one additional level of interlift liners for the full effective surface area of the final lift. Cost: Unit rates for the interlift liners are based on actual purchase orders and installation costs for the interlift liners currently installed at the South Leach Pad.
Other Process Capital	Quantities: Other process capital costs include for small construction and equipment purchase and/or repair works to support the existing crushing and heap leaching equipment. Cost: Installation rates are based on current ongoing contracts, and supply costs are generally from quotations, past purchase orders, and in-house data. Miscellaneous allowances have been for 2029 to end of mine life to account for likely expenditures that have not yet been planned.
General and Administrative	G&A miscellaneous capital spending requirements are projected to support Administration, Finance, Environmental, and Health and Safety departments, as well as for the general operation of the mine. Costs are based on in-house data and allowances.
Other Sustaining Construction	Other sustaining capital costs account for small miscellaneous projects to support general infrastructure that not captured above. Costs are based on in-house data and allowances.

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21.4

Operating Costs

The life-of-mine operating cost estimate has been benchmarked against actual operating cost data collected since the start of operations.

Operating costs include allocations for:

●

Mining

●

Processing

Crushing

HPGR (from Q3 2021)

Overland Conveying and Stacking

Heap Leaching

ADR and Refining

Ancillaries (Power, Water Treatment, Plant Maintenance)

●

General & administration

●

Transport and refining

Operating costs were calculated for each year of operation, totalling an average of US$111 M per annum and US$9.16/t ore over life-of-mine, summarized in Table 21-8. No contingency was included for the operating cost estimate.

Table 21-8: Operating Costs

Category	LOM Average (US$/t)	LOM Expenditure (US$ x 1,000)
Mining	2.76	477,592
Processing	4.92	851,359
General and Administration	1.43	247,794
Transport and Refining	0.05	8,891
Operating Cost	9.16	1,585,727

21.4.1

Basis of Estimate

21.4.1.1

Open Pit Mining

Open pit mining costs were estimated from first principles by unit operation, based on projected fleet requirements for an annual production schedule. Fleet requirements were calculated based on actual Kişladağ mining productivities and haulage simulations run for each period. Equipment operating cost and fuel consumption rates were estimated from historical data from the Kişladağ operations. Haulage fuel costs were calculated during the haulage simulations.

Labour requirements were developed to support the operation and maintenance of the fleet, and for the general operation of the mine. Actual salaries from the Kişladağ operations was applied.

21.4.1.2

Power

Annual power consumption was calculated based on total installed power, load factor, and overall utilization for each unit operation. Power consumption for non-process facilities such as offices, maintenance shops, laboratory, warehouse, etc. were calculated based on total installed power and operating hours per annum. Power cost of $0.073 per kWh was used for the operating cost estimate.

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21.4.1.3

Process and Maintenance Consumables

Process and maintenance consumable costs were calculated based on a combination of calculations, in-house historical data, and vendor recommendations. Budget quotations and past purchase orders were referred to for the supply of all significant consumables.

21.4.1.4

Reagents

Reagent quantities were calculated from a combination of first principles and present operating parameters. Sodium cyanide, representing a large majority of reagent costs, was calculated based on targeting a specified concentration of cyanide for the barren solution. Caustic, anti-scalant, lime and other miscellaneous reagents consumption rates were based on current operating parameters.

21.4.1.5

Labour

The Kişladağ organization chart is based on current staffing levels for the Process and G&A departments, and based on demand for mining. Labour requirement for the process operation was assumed not to change with the introduction of the HPGR in 2021. Salaries applied in the operating cost estimate were based on current salaries and burdens.

21.4.1.6

Site General and Administrative Costs

General and Administrative (G&A) costs were estimated based on historical actual G&A costs.

21.4.1.7

Transport and Refining

Transport and refining costs of US$3.80/oz Au recovered was included in the operating cost estimate based on actual costs data.

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SECTION ● 22

Economic Analysis

22.1

Summary

The Kişladağ economics were analyzed using US$1,400/oz Au and a discount rate of 5%, which generates an after-tax net present value (NPV) of US$582 million. Kisladag is expected to remain a self-funding operation with all capital expenditures outlined in the report to be funded from operating cash flow. As such, there is no internal rate of return (IRR) calculated for the project as net cash flow is expected to remain positive in each year of operation.

The test of economic extraction for the Kisladag mineral reserves is demonstrated by means of a sensitivity analysis. At the mineral reserve metals price of US$1,250/oz Au, the Kişladağ operation shows positive economics with a NPV of US$396 million. At a price of US$1,550/oz Au (a price closer to spot prices at the time writing this report) the NPV for Kisladag increases to US$756 million.

22.2

Methods, Assumptions and Basis

The economic analysis is based on the mineral reserves as outlined in Section 15, the mining methods and production schedule as outlined in Section 16, the recovery and processing methods in Section 17, and the capital and operating costs as outlined in Section 21.

The project case metal price used in the economic model is US$1,400/oz Au, 100% of the gold recovered is payable. Silver credit was assumed to be 0.50/oz silver payable per 1.0/oz gold payable, based on historical averages, the model assumes a silver price of US$18.00/oz.

Transport and refining costs of US$3.80/oz was used for economic analysis, based on historical averages at the Kişladağ operations.

The model has been prepared on a yearly life of mine basis. The LOM is 15 years from the start of 2020 until the depletion of economic mineral reserves.

22.3

Production Schedule

The Kişladağ circuit will operate at the design capacity of 12.6 Mtpa over the life of the operation. The production schedule is shown in Figure 22-1. The average head grade is 0.72 g/t Au.

Gold recovery is estimated for each year of production, averaging 56% for the LOM.

Figure 22-2 shows annual gold production and net cash flow.

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Figure 22-1: Kişladağ Production Schedule and Grade

Figure 22-2: Kişladağ Gold Production Cash Flow

22.4

Royalties and Other Fees

The Kişladağ Project is subject to a mineral production royalty which is based on a sliding scale to gold price, and is payable to the Turkish government.

The relevant royalties are shown in Table 22-1. The royalties are calculated on revenue with deductions allowed for processing and haulage costs of ore.

The royalty regime incorporates a sliding scale depending on the metal price on the date of sale, ranging from 1% to 15%. The royalty rates in Table 22-1 are for sellers of raw ore. If the ore is processed on site the royalty rate is reduced by a further 40%.

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Table 22-1: Gold Royalty

Gold Price (US$ / oz)		Base Royalty Rate
From	To	(%)	Reduced Royalty Rate (Discounted to 60% of Base)
<800	800	1	0.6%
801	900	2	1.2%
901	1000	3	1.8%
1001	1100	4	2.4%
1101	1200	5	3.0%
1201	1300	6	3.6%
1301	1400	7	4.2%
1401	1500	8	4.8%
1501	1600	9	5.4%
1601	1700	10	6.0%
1701	1800	11	6.6%
1801	1900	12	7.2%
1901	2000	13	7.8%
2001	2100	14	8.4%
2101	>2101	15	9.0%

As Tüprag processes the ore on site and produces gold doré production is eligible for the reduced royalty rate. At the project case price US$1,400/oz Au, the base royalty is 7.0% with a reduced royalty rate of 4.2%; Allowable deductions for processing and haulage costs further reducing the effective rate to approximately 2.8% .

22.5

Closure and Salvage Value

Closure costs are captured by the economic model to account for dismantling the processing plant, ancillary buildings, power lines and roads, and rehabilitation. The estimate used in the economic model is US$23.5 M accounting for salvage value at end of mine life.

22.6

Taxation

The Turkish government implemented a temporary rate increase from 20% to 22% for the periods of 2018-2020. From 2021 onwards, the effective tax rate is expected to return to 20%.

22.7

Financing Costs

Cost of financing the Project, such as interest on loans, are not included in the economic model. The Project is expected to be funded from operating cash flow and any costs or charges relating to potential future funding are beyond the scope of the analysis.

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22.8

Third Party Interests

Tüprag is the 100% owner of the Kişladağ Gold Mine. Eldorado owns 100% interest in Tüprag.

22.9

Sensitivity Analysis

The economic model was subjected to a sensitivity analysis (Figure 22-3) to determine the effects of changing metal prices on the Project financial returns. Results are summarized as follows:

Figure 22-3: Sensitivity Analysis on Gold Price

22.10

Cash Flows

The annual cash flow forecast is built from a first principles financial model. The results are shown in Table 22-2.

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Table 22-2: Annual Cashflow Summary

Parameter	Units	Total	2020	2021	2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034
Metal Production
Payable Gold	koz	2,351	250	143	142	174	178	144	137	155	168	164	161	162	155	150	69
Payable Silver	koz	1,167	125	71	71	87	89	72	69	77	84	82	81	81	78	75	26
Revenue
Gross Revenue	US$M	3,312	352	201	200	245	250	203	193	218	236	231	227	228	219	211	97
Treatment & Refining	US$M	9	1	1	1	1	1	1	1	1	1	1	1	1	1	1	0
Royalties	US$M	92	11	5	5	7	7	5	5	6	7	6	6	6	6	6	3
Net Revenue	US$M	3,211	340	195	194	237	242	197	188	212	229	224	220	221	212	204	94
Operating Costs
Mining Cost	US$M	478	21	30	27	22	27	43	45	36	34	35	37	39	38	33	10
Processing	US$M	851	63	64	59	60	60	60	60	60	60	60	60	60	60	46	23
G&A	US$M	248	15	14	14	14	14	14	20	20	20	20	19	18	17	16	15
Inventory Change	US$M	43	13	-2	-5	3	0	-6	-10	1	3	-2	-6	-3	-4	30	30
Total Operating Costs	US$M	1,619	112	105	94	99	101	111	114	116	117	113	110	114	110	126	78
Cost Per Ounce
C1 - Cash Cost	US$/oz	684	442	733	658	564	562	764	826	747	692	681	678	697	705	836	1,134
C2 - Total Cash Cost	US$/oz	723	487	769	695	605	603	801	860	784	731	720	716	735	743	878	1,173
C3 - Site Sustaining Cost	US$/oz	827	605	1,039	814	773	763	869	934	852	842	851	773	793	787	906	1,211
Capital Costs
Growth Capital	US$M	291	75	44	47	46	48	30	0	0	0	0	0	0	0	0	0
Sustaining Capital	US$M	245	29	38	17	29	28	10	10	11	18	22	9	9	7	4	3
Total Capital Costs	US$M	536	105	82	64	75	77	40	10	11	18	22	9	9	7	4	3
CASH FLOW
EBITDA	US$M	1,591	228	90	100	138	142	86	74	95	112	112	110	108	102	78	16
Inventory Changes	US$M	43	13	-2	-5	3	0	-6	-10	1	3	-2	-6	-3	-4	30	30
Pre-tax Cash Flow	US$M	1,074	136	6	31	66	64	40	54	86	97	89	95	96	91	104	19
Depreciation	US$M	606	63	63	71	69	73	56	26	24	23	25	20	16	15	13	10
Income Tax	US$M	208	36	5	6	14	14	6	10	14	18	17	18	18	17	13	1
After-tax Cash Flow	US$M	866	100	0	25	52	51	34	44	72	79	71	77	77	74	91	18

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SECTION ● 23

Adjacent Properties

There are no mineral properties of importance adjacent to the Kişladağ mine site.

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SECTION ● 24

Other Relevant Data and Information

24.1

HPGR Project Schedule

The HPGR option involves the installation of a HPGR crusher replacing the existing tertiary crushing and screening circuit.

The Kişladağ HPGR project requires approximately 15 months to complete and commissioning will occur at the end of second quarter 2021. Request for quotations (RFQ) and full bid evaluations for HPGR will be completed by the end of Q1 2020.

The Project is not complex but basic engineering will need to evaluate the current design and operational improvements to fully evaluate the new project and tie-ins to the existing systems and with incoming vendor data and is scheduled for 6 months.

Early construction works will be undertaken during basic engineering within the existing permit constraints. This includes preliminary earthworks, upgrades and maintenance of the existing equipment and facilities, and relocation of secondary equipment and extension of some conveyors.

The crushing operation will remain fully functional with the normal maintenance shutdowns utilized for a majority of the smaller modifications to existing equipment and a one month shutdown for the major tie at end of the construction phase.

The summary level schedule for development of the Project is shown in Figure 24-1.

Figure 24-1: Kişladağ HPGR Project, Implementation Schedule

24.2

Manpower Estimate

The existing process and G&A team will remain constant.

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SECTION ● 25

Interpretation and Conclusions

It is concluded that the work completed in the prefeasibility study indicate that the mineral resource and mineral reserve estimates and Project economics are sufficiently defined to indicate the Project is technically and economically viable. Key conclusions are listed below.

25.1

Mineral Resources and Mineral Reserves

The mineral resource and mineral reserve are consistent with the CIM definitions referred to in NI 43-101. It is the opinion of the qualified persons that the information and analysis provided in this report is considered sufficient for reporting mineral resources and mineral reserves.

The mineral resource model was used as input for the mineral reserve estimate. The modelling methods, grade models, resource classification, and density model were reviewed and found appropriate for the mineral reserve estimation.

The mineral reserves for the deposit were estimated using a gold price of US$1,250/oz. The mineral reserves are reported using a 0.19 g/t recoverable Au (approximately $7.29/t NSR) cut-off for all ore, where the recovery factor used is respective of whether the ore is crushed by three-stage crushing or the HPGR circuit. The proven and probable mineral reserves are 173.2 Mt with an average grade of 0.72 g/t Au containing 4.0 Moz Au.

25.2

Mining Methods

Mining will use a conventional fleet consisting of seven diesel drills, two electric drills, one 29 m3 electric hydraulic shovel, two 21 m3 diesel hydraulic shovels, two 21.4 m3 wheel loaders, one 12 m3 wheel loader, fourteen 136 tonne trucks and ten 219 tonne trucks. The major equipment is supported by a fleet of graders, dozers, a backhoe and water trucks.

Ore and waste will be mined on 10 m benches. Mining operations involves conventional open pit mining delivering ore to the primary crusher for processing while all associated waste rock will be placed in either the south rock dump (SRD) or the north rock dump (NRD), as best suited. The life of mine strip ratio is approximately 1.12:1.

A total of 193.2 Mt of waste will be mined (inclusive of 141.1 Mt of capitalized waste). The average LOM gold grade of all ore is forecast to be 0.72 g/t Au. The recoverable grade is estimated to be 0.40 g/t for a LOM recovery of 56%.

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25.3

Metallurgical Testwork

Testing by 2m column test revealed that the gold leach cycle time requirement has increased with depth in the open pit. Whereas previously, mineralization required approximately 90 to 120 days of irrigation to meet “maximum” gold extraction, the deeper mineralization will require approximately 300 days of irrigation. The columns testwork results were modeled to create a 3D recovery model to supplement and use with the gold grade model.

An ancillary testing program involved investigation of implementing HPGR (High Pressure Grinding Rolls) technology. Testing proved positive with an achievement of an additional 3.9 percentage points recovery globally.

25.4

Process Design

The process plant design will continue to utilize the existing primary and secondary crushing circuits with a HPGR circuit installed replacing the tertiary crushing and fine ore circuits. The new configuration will be capable of processing 12.6 Mtpa to an 80% passing product size 6.5 mm.

Cyanide heap leaching process is robust for treatment of Kişladağ sulphide ore.

Bulk flotation was not considered economically viable because test results indicated that a significant mass pull is required and overall recovery would be significantly reduced compared to whole ore leaching.

The existing portable stacking equipment will be utilized for stacking the ore on the south leach pad and reconfigured to the new north leach pad when it is required. A minimal amount of new conveyors will be required to access the north leach pad.

The new process plant was designed on the basis of overall plant operating time of 80% and 365 days per year for a total operating time of 7,008 h/y. The process plant has been designed to produce up to 300,000 oz/a gold as doré bar with average annual production of 160,000 oz/a.

25.5

Project Infrastructure

The HPGR crushing facility will be incorporated into the crushing plant replacing the current tertiary crushers.

No upgrades are required to the existing access road, power or water supplies for the addition of the new facility.

Additional security fencing will be added to encompass the North Leach Pad, North Rock Dump and associated water and process solution management infrastructure. There will be no additional access gates required.

Fresh water will continue to be supplied from the well fields and Gedikler dam. Existing water storage tanks and underground distribution system will continue to provide process, non-potable, and fire protection water as required.

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The main substation at site is rated at 100 MVA with three 50 MVA 154 / 34.5 kV power transformers (one cold spare). The substation currently has abundant spare capacity and will require only minor modifications to accommodate the additional load associated with the conversion to HPGR. Site distribution is at 34.5 kV and in local areas at 6.6 kV or 0.4 kV.

25.6

Waste Rock Dump

Recent studies indicate that the SRD has a remaining capacity of 158.2 Mt within the permitted boundaries. Existing designs for the NRD range 110 Mt to 1,000 Mt. The LOM plan is to place 34.9 Mt in the NRD. A new design will be made for this quantity. Furthermore, an optimization study of the two rock dumps will be done to see if cost savings are achievable.

25.7

Capital and Operating Costs

The total growth capital cost includes the life-of-mine capitalized waste stripping costs, as well as the initial investment cost to obtain commercial product of a new HPGR circuit. Total growth capital cost is US$290.8 M; $ 35.8 M of this is for the HPGR purchase and installation. Sustaining capital costs for the life-of-mine total US$245.2 M.

Operating costs were calculated for each year of operation, totaling average of US$181.4 M per annum for an average of US$14.13/t ore over life-of-mine.

25.8

Economic Analysis

The economic model has been built from first principles and includes all relevant data. The qualified persons have a high level of confidence in the stated economic performance of the Project.

The economic model was subjected to a sensitivity analysis to determine the effects of changing metal prices and capital and operating expenditures on the Project financial returns. It was concluded that the Project economics are robust and will continue to be economic with higher capital and operating costs, or lower gold prices.

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SECTION ● 26

Recommendations

Opportunities may exist to improve the mine plan going forward. These include:

●

Schedule modifications to optimize the truck fleet utilization while avoiding capital equipment additions. This will conceptually include enhancements to the Phase 4 design and advancing higher valued areas of Phase 5. This work will also aim to bring some gold production sooner.

●

Simple modifications of the primary crusher feed bin could allow the larger Hitachi trucks to replace the dependency of using Caterpillar trucks for all ore.

●

Optimize the loading equipment fleet distribution including utilization of the Hitachi 5600 as the primary stripping shovel.

●

Phase in the electric blasthole drills in wide areas of pit development rather than re-building older diesel drills.

●

Continue testing and evaluate the economics of converting the haul trucks from the current conventional diesel engines to dynamic gas blending of liquefied natural gas-diesel powered engines.

●

The new larger pit design and incorporating the NRD makes the use of trolley assisted haulage for the Hitachi trucks more feasible. However, at the current diesel costs the capital cost likely still outweighs the economic benefit. This should however be looked at again if diesel costs rise.

The scheduling and ultimate design of the two rock dumps needs to be optimized with respect to the timing of waste release and where it is coming from. Although the NRD has a further lateral haulage from the open pit centroid it does not have as much of a vertical rise that the last phases of the SRD has. Sequencing their development in harmony with the location of the material in the pit will improve overall economics. There is also scope for a limited amount of in-pit waste disposal towards the end of the mining phase.

Process plant opportunities that will further be investigated include:

●

Alternative systems of fine ore stockpiling and reclaiming system to compare capital and operational availability.

●

Agglomeration studies to assess leach pad percolation and effect on overall recoveries.

●

Optimize HPGR recirculation (edge effect) and screening of fine materials to optimize HPGR circuit design.

●

Optimization studies and debottlenecking studies to increase system capacity beyond 12.6 Mtpa.

An investment tax incentive is available in Turkey when new equipment is purchased to improve operations; the incentive is not included in the economics but will be applied for during future purchase. The incentive allows for 40% of eligible capital costs spent on new production facilities to be applied as an incentive value where the taxable income is taxable at a reduced rate of 4.0% until the total value of tax savings equals the incentive value. The incentive value can be carried forward; after the incentive value is utilized, the 20% corporate taxation rate is applied.

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References

Agostini, S., Doglioni, C., Innocenti, F., Manetti, P., and Tonarini, S., 2010, On the geodynamics of the Aegean rift: Tectonophysics, v. 488, p. 7-21.

Baker, T., Bickford, D., Juras, S., Lewis, P., Oztas, Y., Ross, K., Tukac, A., Rabayrol, F., Miskovic, A., Friedman, R., Creaser, R.A., and Spikings, R., 2016, The Geology of the Kisladag Porphyry Gold Deposit, Turkey: SEG SPECIAL PUBLICATION 19, p. 57–83.

Eldorado Gold Corporation, 2010, Technical Report for the Kışladağ Gold Mine, Turkey, NI 43-101 Technical Report, January 2010.

Eldorado Gold Corporation, 2018, Technical Report, Kışladağ Milling Project, Turkey, NI 43-101 Technical Report, March 2018.

Ercan, T., Dinçel, A., Metin, S., Türkecan, A., and Günay, E., 1978, Geology of the Neogene basins in Uşak region: Bulletin of the Geological Society of Turkey, v. 21, p. 97-106.

Hatch, 2003, Technical Report Kışladağ Project, Feasibility Study, NI 43-101 Technical Report, March 2003.

Jolivet, L., Faccenna, C., Huet, B., Labrousse, L., Le Pourhiet, L., Lacombe, O., Lecomte, E., Burov, E., Denèle, Y.; Brun, J.P., Philippon, M., Paul, A.; Salaün, G., Karabulut, H., Piromallo, C., Monié, P., Gueydan, F., Okay, A.I., Oberhänsli, R., Pourteau, A., Augier, R., Gadenne L., Driussi O., 2013, Aegean tectonics: Strain localization, slab tearing and trench retreat. Tectonophysics, v. 597, p. 1–33.

Karaoğlu, Ö., Helvacı, C., and Ersoy, Y., 2010, Petrogenesis and 40Ar/39Ar geochronology of the volcanic rocks of the Uşak-Güre basin, western Türkiye: Lithos, v. 119, Issues 3-4, p. 193-210.

Karaoğlu, Ö., and Helvaci, C., 2012, Structural evolution of the Uşak–Güre supra-detachment basin during Miocene extensional denudation in western Turkey: Journal of the Geological Society, v. 169, Issue 5, p. 627-642.

Micon, 2003, 2003 Update of Resources, Kışladağ Project, Usak, Turkey, NI 43-101 Technical Report, September 2003.

Richards, J. P., 2015, Tectonic, magmatic, and metallogenic evolution of the Tethyan orogen: From subduction to collision: Ore Geology Reviews, v. 70, p. 323–345.

Şengör, A.M.C., Yılmaz, Y., and Ketin, I., 1981, Remnants of a pre-Late Jurassic ocean in northern Turkey: fragments of a Permian-Triassic Paleotethys: Geological Society of American Bulletin, v. 91.

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Date and Signature Page

The effective date of this report entitled “Technical Report, Kişladağ Gold Mine, Turkey” is January 17, 2020. It has been prepared for Eldorado Gold Corporation by David Sutherland, P. Eng., Stephen Juras, Ph.D.,P.Geo., Paul Skayman, FAusIMM, Richard Miller, P.Eng. and Sean McKinley P.Geo. , each of whom are qualified persons as defined by NI 43-101.

Signed the 28th day of February 2020.

“Signed and Sealed” David Sutherland David Sutherland, P. Eng.	“Signed and Sealed” Stephen J. Juras Stephen J. Juras, Ph.D.,P.Geo.
“Signed” Paul J. Skayman Paul J. Skayman, FAusIMM	“Signed and Sealed” Richard Miller Richard Miller, P.Eng.
“Signed and Sealed” Sean McKinley Sean McKinley, P. Geo.

2020 Final Report

Kişladağ Gold Mine, Turkey

Technical Report

CERTIFICATE OF QUALIFIED PERSON

Stephen J. Juras, Ph.D., P.Geo

1188 Bentall 5, 550 Burrard St.

Vancouver, BC

Tel: (604) 601-6658

Fax: (604) 687-4026

Email: stevej@eldoradogold.com

I, Stephen J. Juras, am a Professional Geoscientist, employed as Director, Technical Services, of Eldorado Gold Corporation and reside at 9030 161 Street in the City of Surrey in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Kişladağ Gold Mine, Turkey, with an effective date of January 17, 2020.

I am a member of Engineers & Geoscientists British Columbia. I graduated from the University of Manitoba with a Bachelor of Science (Honours) degree in geology in 1978 and subsequently obtained a Master of Science degree in geology from the University of New Brunswick in 1981 and a Doctor of Philosophy degree in geology from the University of British Columbia in 1987.

I have practiced my profession continuously since 1987 and have been involved in: mineral exploration and mine geology on gold, copper, lead, zinc and silver properties in Canada, United States, Brazil, China, Greece and Turkey; ore control and resource modelling work on gold, copper, lead, zinc, silver, platinum/palladium and industrial mineral open pit and underground properties and operations in Canada, United States, Mongolia, China, Brazil, Turkey, Greece, Romania, Peru and Australia; and geometallurgical matters on gold, lead, and zinc properties and operations in Turkey and Greece.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I have visited the Kişladağ Gold Mine on numerous occasions with my most recent visit occurring on January 13 to 17, 2020.

I was responsible for reviewing matters related to the geological and geometallurgical data and directing the mineral resource estimation and classification work for the Kişladağ Gold Mine, Turkey. I am responsible for the preparation or supervising the preparation of sections 2, 11, 12, 14, 25 and 26 in the technical report.

I have had continual prior involvement with the property that is the subject of this technical report.

I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.

I have read National Instrument 43-101 and Form 43-101FI and the items for which I am responsible in this report entitled, Technical Report, Technical Report, Kişladağ Gold Mine, Turkey, with an effective date of January 17, 2020, has been prepared in compliance with same.

As of the effective date of the technical report, to the best of my knowledge, information and belief, the items of the technical report that I was responsible for contain all scientific and technical information that is required to be disclosed to make the technical report not misleading

Dated at Vancouver, British Columbia, this 28th day of February 2020.

“Signed and Sealed”

Stephen J. Juras

Stephen J. Juras, Ph.D., P.Geo.

Kişladağ Gold Mine, Turkey

Technical Report

CERTIFICATE OF QUALIFIED PERSON

Sean McKinley, P.Geo.

1188 Bentall 5, 550 Burrard St.

Vancouver, BC

Tel: (604) 601-6658

Fax: (604) 687-4026

Email: Sean.McKinley@eldoradogold.com

I, Sean McKinley, am a Professional Geoscientist, employed as Senior Geologist- Resource Development, of Eldorado Gold Corporation and reside at 2231 Bellevue Ave in the City of Coquitlam in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Kişladağ Gold Mine, Turkey, with an effective date of January 17, 2020.

I am a member of the Engineers & Geoscientists British Columbia. I graduated from Queen’s University in Kingston, Ontario with a Bachelor of Science (Honours) degree in geology in 1992 and subsequently obtained a Master of Science degree in geology from the University of British Columbia.

I have practiced my profession continuously since 1996 and have been involved in: mineral exploration (both greenfields and brownfields), mine geology (underground and open pit settings) and geological modelling on gold, copper, lead, zinc and silver projects in Canada, Ireland, Sweden, China, Mexico, Romania, Greece and Turkey.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I have visited the Kişladağ Gold Mine on numerous occasions with my most recent visit occurring on September 16 to 18, 2019.

I was responsible for the preparation of the sections in this report concerned with geological information, exploration and drilling for this technical report. I am responsible for the preparation or supervising the preparation of items 7, 8, 9, 10 and 23 in the technical report.

I have had continual prior involvement with the property that is the subject of this technical report.

I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.

I have read National Instrument 43-101 and Form 43-101FI and the items for which I am responsible in this report entitled, Technical Report, Kişladağ Gold Mine, Turkey, with an effective date of January 17, 2020, has been prepared in compliance with same.

Dated at Vancouver, British Columbia, this 28th day of February 2020.

“Signed and Sealed”

Sean McKinley

Sean McKinley, P.Geo.

Kişladağ Gold Mine, Turkey

Technical Report

CERTIFICATE OF QUALIFIED PERSON

Richard Miller, P.Eng.

1188 Bentall 5, 550 Burrard St.

Vancouver, BC

Tel: (604) 601-6671

Fax: (604) 687-4026

Email: richard.miller@eldoradogold.com

I, Richard Miller, am a Professional Engineer, employed as Director, Mine Engineering (Open Pit), of Eldorado Gold Corporation and reside at 832 Victoria Drive in the City of Port Coquitlam in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Kişladağ Kisladag Gold Mine, Turkey, with an effective date of January 17th, 2020.

I am a member of Engineers & Geoscientists British Columbia. I graduated from the University of British Columbia with a Bachelor of Applied Science degree in Mining and Mineral Process Engineering in 1987.

I have practiced my profession continuously since 1987 and have worked at open pit gold mining operations in Namibia, Guinea and Turkey, including being Mine Manager of Kişladağ Gold Mine from 2004 to 2007 and General Manager at Kisladag Gold Mine from 2014 to 2018. Additionally I have worked at several other types of underground and open pit mines, and have been involved with mine designs in Canada, China, Brazil, Greece, Romania and South Africa.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I have visited the Kişladağ Gold Mine on numerous occasions with my most recent visit occurring on January 13 to 24, 2020.

I was responsible for directing the mineral reserve estimation, mining methods and economic analysis work for the Kişladağ Gold Mine, Turkey. I am responsible for the preparation or supervising the preparation of sections 15, 16, 21 and 22 in the technical report.

I have had continual prior involvement with the property that is the subject of this technical report.

I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.

Dated at Vancouver, British Columbia, this 28th day of February 2020.

“Signed and Sealed”

Richard Miller

Richard Miller, P.Eng.

Kişladağ Gold Mine, Turkey

Technical Report

CERTIFICATE OF QUALIFIED PERSON

Paul J. Skayman, FAusIMM

1188 Bentall 5, 550 Burrard St.

Vancouver, BC

Tel: (604) 601-6658

Fax: (604) 687-4026

Email: paul.skayman@eldoradogold.com

I, Paul J. Skayman, am a Professional Extractive Metallurgist, employed as Special Advisor to the COO, of Eldorado Gold Corporation and reside at 3749 West 39th Avenue in the City of Vancouver, in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Kişladağ Gold Mine, Turkey, with an effective date of January 17, 2020.

I am a fellow of the Australian Institute of Mining and Metallurgy. I graduated from the Murdoch University with a Bachelor of Science (Extractive Metallurgy) degree in 1987.

I have practiced my profession continuously since 1987 and have been involved in operation and management of gold and base metal extraction operations in Australia, Ghana, Tanzania, Guinea, China Turkey and Greece. This work has also included Feasibility Studies, Project Acquisition, Development / Construction and closure of said projects.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I have visited the Kişladağ Gold Mine on numerous occasions with my most recent visit occurring on January 13 to 17, 2020.

I was responsible for the preparation of the sections in this report that dealt with metallurgy and process operations and related costs and payability of the technical report. I am responsible for the preparation or supervising the preparation of items 13, 17 and 19 in the technical report.

I have had continual prior involvement with the property that is the subject of this technical report.

I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.

Dated at Vancouver, British Columbia, this 28th day of February 2020.

“Signed”

Paul J. Skayman

Paul J. Skayman, FAusIMM

Kişladağ Gold Mine, Turkey

Technical Report

CERTIFICATE OF QUALIFIED PERSON

David Sutherland, P. Eng.

1188 Bentall 5, 550 Burrard St.

Vancouver, BC

Tel: (604) 601-6658

Fax: (604) 687-4026

Email: david.sutherland@eldoradogold.com

I, David Sutherland, am a Professional Engineer, employed as Project Manager, of Eldorado Gold Corporation located at 1188 Bentall 5, 550 Burrard St., in the City of Vancouver, in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Kişladağ Gold Mine, Turkey, with an effective date of January 17, 2020.

I am a member of the Engineers & Geoscientists of British Columbia. I graduated from the Lakehead University with a Bachelor of Science (Physics) in 2003 and a Bachelor of Engineering (Mechanical) in 2005.

I have practiced my profession continuously since 2005. Since receiving my profession designation, I have worked exclusively on the design of mineral processing plants, assisted on numerous NI43-101 studies and have directed engineering and procurement on three mineral processing projects through construction. For 30 years I have been working in heavy industry including operations, maintenance and construction. During this time, I have lead the design and construction of major greenfield and brownfield construction projects in Canada, Turkey, and Greece.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I have visited the Kişladağ Gold Mine on numerous occasions with my most recent visit occurring on July 10 to 12, 2018.

I was responsible for coordinating the preparation of the technical report. I am responsible for the preparation or supervising the preparation of items 1, 3, 4, 5, 6, 18, 20, 24, and 27 in the technical report. .

I have had continual prior involvement with the property that is the subject of this technical report.

I am not independent of Eldorado Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.

Dated at Vancouver, British Columbia, this 28th day of February 2020.

“Signed and Sealed”

David Sutherland

David Sutherland, P. Eng.