Exhibit 99.1

Technical Report

 Efemçukuru Gold Mine

 Turkey

Centered on Latitude N 38° 17' 31.59" and Longitude E 26° 58' 30.30"

Effective Date: December 31, 2019

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
Mr. Paul Skayman, FAusIMM.	Eldorado Gold Corporation
Ms. Imola Götz, P.Eng.	Eldorado Gold Corporation
Mr. Sean McKinley, P.Geo.	Eldorado Gold Corporation
Mr. Ertan Uludag, P.Geo.	Eldorado Gold Corporation

Efemçukuru Gold Mine, Turkey Technical Report

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 Resources Estimates	1-4
	1.5	Mineral Reserves and Mining Methods	1-5
	1.6	Mineral Processing and Recovery Methods	1-6
	1.7	Project Infrastructure	1-7
	1.8	Environmental	1-7
	1.9	Capital and Operating Costs	1-8
	1.10	Economic Analysis	1-9
	1.11	Interpretations and Conclusions	1-10
	1.12	Recommendations	1-11
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	Permits and Agreements	4-3
SECTION ● 5     ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			5-1
	5.1	Access and Infrastructure	5-1
	5.2	Physiography	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	7-3
SECTION ● 8     DEPOSIT TYPES			8-1
	8.1	Deposit Geology	8-1
	8.2	Deposit Model	8-3
SECTION ● 9     EXPLORATION			9-1
	9.1	Surface and Subsurface Exploration Work	9-1
	9.2	Mapping	9-3
	9.3	Surface Sampling and Trenching	9-3
	9.4	Geophysics	9-3
	9.5	Mine Exploration	9-3

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SECTION ● 10     DRILLING			10-1
	10.1	Project Drilling	10-1
	10.2	Infill Drilling	10-3
SECTION ● 11     SAMPLE PREPARATION, ANALYSES AND SECURITY			11-1
	11.1	Sample Method	11-1
	11.2	Assay Method	11-1
	11.3	Quality Assurance and Quality Control (QA/QC) Program	11-1
	11.4	Concluding Statement	11-7
SECTION ● 12     DATA VERIFICATION			12-1
	12.1	Introduction	12-1
	12.2	Block Model to Mill Reconciliation	12-1
SECTION ● 13     MINERAL PROCESSING AND METALLURGICAL TESTWORK			13-1
	13.1	Introduction	13-1
	13.2	Efemçukuru Gold Ore	13-1
	13.3	Flotation	13-3
	13.4	Gravity Concentration	13-4
	13.5	Performance Analysis of the Flotation Circuit	13-5
	13.6	Recommendations for Future Testwork	13-6
SECTION ● 14     MINERAL RESOURCE ESTIMATES			14-1
	14.1	Introduction	14-1
	14.2	Mineralization Models	14-1
	14.3	Data Analysis	14-3
	14.4	Variography	14-8
	14.5	Model Set-up	14-9
	14.6	Estimation	14-9
	14.7	Mineral Resource Classification and Summary	14-18
SECTION ● 15     MINERAL RESERVE ESTIMATES			15-1
	15.1	Cut-off Grades	15-1
	15.2	Planning Cut-off Grade	15-1
	15.3	Incremental Cut-off Grade	15-2
	15.4	Marginal Cut-off Grade	15-2
	15.5	Dilution	15-4
	15.6	Mineral Reserves Statement	15-5
SECTION ● 16     MINING METHODS			16-1
	16.1	Introduction	16-1
	16.2	Mining Methods	16-2
	16.3	Underground Mine Desıgn	16-5
	16.4	Geotechnical Assessment	16-6
	16.5	Backfill	16-10
	16.6	Mining Rate	16-11
	16.7	Mine Development and Production Assumptions	16-11
	16.8	Mining Sequence	16-11
	16.9	Mobile Equipment Fleet	16-12
	16.10	Mine Schedule	16-13
	16.11	Ventilation	16-14
	16.12	Mining Services	16-16

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SECTION ● 17      RECOVERY METHODS			17-1
	17.1	Introduction	17-1
	17.2	Process Performance	17-1
	17.3	Process Flowsheet and Equipment	17-5
	17.4	Production	17-10
SECTION ● 18     PROJECT INFRASTRUCTURE			18-1
	18.1	Site Access and Local Roads	18-1
	18.2	Site Access Road	18-1
	18.3	Site Layout	18-1
	18.4	Surface Tailings and Development Rock Management	18-3
	18.5	Fire/Fresh Water Supply Storage and Distribution	18-5
	18.6	Sewage Collection and Treatment	18-8
	18.7	Waste Disposal	18-8
	18.8	Power Supply and Electrical Distribution	18-8
	18.9	Communications	18-12
	18.10	Ancillary Facilities	18-12
SECTION ● 19     MARKET STUDIES AND CONTRACTS			19-1
	19.1	Gold Market	19-1
	19.2	Gold Price	19-1
	19.3	Contracts	19-1
SECTION ● 20     ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT			20-1
	20.1	Introduction	20-1
	20.2	Air Quality	20-1
	20.3	Water Quality	20-3
	20.4	Land Use	20-7
	20.5	Flora and Fauna	20-8
	20.6	Approvals and Permits	20-8
	20.7	Conclusions	20-10
SECTION ● 21     CAPITAL AND OPERATING COSTS			21-1
	21.1	Capital Cost estimate	21-1
	21.2	Operating Cost Estimate	21-1
SECTION ● 22     ECONOMIC ANALYSIS			22-1
	22.1	Basis	22-1
	22.2	Cashflow	22-1
	22.3	Sensitivities	22-3
	22.4	Opportunities and Risks	22-4
	22.5	Conclusions and Recommendations	22-5
SECTION ● 23     ADJACENT PROPERTIES			23-1
SECTION ● 24     OTHER RELEVANT DATA AND INFORMATION			24-1
	24.1	Life of asset strategy	24-1
SECTION ● 25     INTERPRETATION AND CONCLUSIONS			25-1

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SECTION ● 26     RECOMMENDATIONS			26-1
	26.1	Geology - Exploration	26-1
	26.2	Mining – planning and operational	26-1
	26.3	Metallurgy and Processing	26-1
SECTION ● 27      REFERENCES			27-1
SECTION ● 28     DATE AND SIGNATURE PAGE			28-1

LIST OF FIGURES

Figure 1 1: Property Location Map - Çeşme Peninsula Turkey	1-2
Figure 1 2: LOM Plan NPV Sensitivities based on Au Price Variation	1-10
Figure 1 3: LOM Plan NPV Sensitivities based on Cost Variations	1-10
Figure 4 1: Location Map Showing Project Location	4-2
Figure 4 2: Efemçukuru Land Position	4-3
Figure 5 1: Project Accessibility Climate	5-1
Figure 5 2: Efemçukuru Mine Facilities Layout	5-3
Figure 6 1: Current Drilling Program Site	6-2
Figure 7 1: Simplified Tectonic Map of the Western Anatolian Extensional Province, outlining Major Structures, Geologic Units, and Locations of Mineral Deposits (from Boucher, 2016)	7-2
Figure 7 2: Geological Map of the Efemçukuru Deposit	7-3
Figure 7 3: Cross Section through the Bati, Kestane Beleni (MOS) and Kokarpinar Veins	7-5
Figure 8 1: Long Section with Gold Distribution & Illustrating Shoots at Kestane Beleni	8-2
Figure 8 2: Gold deportment at Efemçukuru	8-1
Figure 8 3: Mineral Paragensis at Efemçukuru (from Boucher, 2016)	8-2
Figure 9 1: Exploration drilling at Efemçukuru	9-2
Figure 10 1: Drill Hole Location Map at Efemçukuru Mine Area	10-2
Figure 10 2: Diamond Drill Rig in an Underground Drill Station	10-4
Figure 10 3: Infill Drilling Pattern a) 10 m x 10 m grid b) five dice pattern	10-4
Figure 11 1: CRM Charts, 2017 to 2019, Efemçukuru Mine	11-4
Figure 11 2: Efemçukuru Blank Data – 2017-2019 Drill Programs	11-5
Figure 11 3: Relative Difference Chart, Efemçukuru 2017-2019 Duplicate Data	11-5
Figure 11 4: Percentile Rank Plot, Efemçukuru 2017-2019 Duplicate Data	11-6
Figure 11 5: QQ Plot for Efemçukuru Duplicate Data	11-7
Figure 12 1: Flowchart of the Reconciliation Procedure at Efemçukuru Mine	12-1
Figure 14 1: Modeling domains for the Kestane Beleni, Kokarpinar and Bati vein systems	14-2
Figure 14 2: Modeled voids in the Kestane Beleni Vein System	14-3
Figure 14 3: Average Coefficient of Variation for Capped and Uncapped Data by Domain	14-5
Figure 14 4: Cumulative Probability Plots, Kestane Beleni Vein System	14-6
Figure 14 5: Cumulative Probability Plots, Kokarpinar and Bati Systems	14-7
Figure 14 6: SOS showing modeled Gold Grades and composited Drill Hole Samples (g/t)	14-10

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Figure 14 7: MOS showing modeled Gold Grades and composited Drill Hole Samples (g/t)	14-11
Figure 14 8: NOS showing modeled Gold Grades and composited Drill Hole Samples (g/t)	14-12
Figure 14 9: KBNW showing modeled Gold Grades and composited Drill Hole Samples (g/t)	14-12
Figure 14 10: Kokarpinar Domains (KPS on left; KPM on right) showing modeled Gold Grades and composited Drill Hole Damples (g/t)	14-13
Figure 14 11: Bati veins (#1 on left; #2 on right) showing modeled Gold Grades and composited Drill Hole Samples (g/t)	14-13
Figure 14 12: Model trend plot showing 10m binned averages along elevations for kriged (OK) and nearest neighbour (NN) gold grade estimates, SOS domain	14-14
Figure 14 13: Model trend plot showing 10m binned averages along elevations for kriged (OK) and nearest neighbour (NN) gold grade estimates, MOS domain	14-15
Figure 14 14: Model trend plot showing 10m binned averages along elevations for kriged (OK) and nearest neighbour (NN) gold grade estimates, NOS domain	14-15
Figure 14 15: HERCO plot, SOS domain	14-16
Figure 14 16: HERCO plot, MOS domain	14-17
Figure 14 17: HERCO plot, NOS domain	14-17
Figure 14 18: Mineral Resources Classification, Efemçukuru Mine	14-18
Figure 15 1: Schematic Showing Dilution Encapsulated by MSO Mining Shape	15-4
Figure 16 1: Isometric View Showing Mine Areas and Vein Wireframe	16-1
Figure 16 2: Schematic Cross Section Showing DAF Production Block Layout	16-3
Figure 16 3: Schematic Long and Cross Sections Showing LHOS Production Block Layout	16-4
Figure 16 4: View Looking South-West Showing Production Blocks	16-4
Figure 16 5: Back Analysis of Stope Performance	16-9
Figure 16 6: Schematic Showing DAF Mining Sequence for a Typical Production Block	16-12
Figure 16 7: Schematic Showing LHOS Mining Sequence for a Typical Production Block	16-12
Figure 16 8: Current and LOM Plan Mine Development	16-13
Figure 16 9: Primary Ventilation Components	16-15
Figure 16 10: Power Line	16-17
Figure 17 1: Simplified Process Flow Diagram of Efemçukuru Process Plant	17-2
Figure 17 2: Simplified Process Flow Diagram of Efemçukuru Process Plant with Column Flotation Cells	17-3
Figure 17 3: Crusher availability and utilization between 2013 and 2019	17-5
Figure 17 4: Distribution of Flotation Tailings to TSF and Paste Plant	17-9
Figure 17 5: Availability and utilization of flotation plant between 2013 and 2019	17-10
Figure 18 1: Mine Main Access Road	18-2
Figure 18 2: General View of the Mine Site	18-3
Figure 18 3: View of Tailings Area	18-5
Figure 18 4: WRSP Sedimentation Pond	18-6
Figure 18 5: East Pond	18-6

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Figure 18 6: Water Treatment Plant	18-7
Figure 18 7: Power Line	18-10
Figure 18 8: Process Plant and Crushed Ore Bins	18-12
Figure 18 9: Filtration Plant and North Portal	18-13
Figure 18 10: Mine Building beside South Portal	18-14
Figure 19 1: Gold Price Movement	19-1
Figure 20 1: Air Quality-Monitoring Station	20-3
Figure 20 2: Water Quality Monitoring	20-4
Figure 22 1: LOM Plan Cashflow and Cumulative NPV	22-1
Figure 22 2: LOM Plan NPV Sensitivities based on Au Price Variation	22-3
Figure 22 3: LOM Plan NPV Sensitivities based on Cost Variations	22-3
Figure 24 1: Extended LOM Design	24-1

LIST OF TABLES

Table 1 1: Efemçukuru Gold Mine Mineral Resources, as of December 31, 2019	1-4
Table 1 2: Efemçukuru Mineral Reserves Effective December 31, 2019	1-5
Table 1 3: Capital Cost Summary	1-8
Table 1 4: Operating Cost Summary	1-9
Table 2 1: Cross-reference List	2-2
Table 4 1: Royalties Calculation	4-4
Table 4 2: Key Project Permits	4-4
Table 5 1: Distribution of Annual Climate Data	5-3
Table 6 1: Summary of Drilling on the Efemçukuru Deposit	6-2
Table 9 1: Exploration Activity Summary for Efemçukuru Mining Area	9-1
Table 10 1: Summary of Drilling on the Efemçukuru Deposit	10-1
Table 10 2: Efemçukuru Deposit Drilling between 2017 and 2019	10-2
Table 11 1: QAQC sample frequency, 2017 to 2019	11-2
Table 11 2: Main CRM Samples used between 2017 and 2019	11-3
Table 12 1: Annual Resource Model to Mill Reconciliation data	12-2
Table 13 1: Reports Reviewed	13-1
Table 13 2: Head Assay of typical High-Grade Ores	13-2
Table 13 3: Modal Mineralogy of One Plant Feed Sample in February 2018	13-3
Table 13 4: Comminution Characteristics of Efemçukuru Ore	13-3
Table 13 5: Concentrate Mass Pull, Grade and Recovery of Final Concentrates from Locked Cycle Tests	13-4
Table 13 6: Average Mass Pull, Grade and Recoveries of the Continuous Column Flotation Tests	13-6
Table 14 1: Project Limits and Block Model Properties	14-1
Table 14 2: Efemçukuru Statistics for 1 m Uncapped Composite Au Data (g/t)	14-4
Table 14 3: Efemçukuru Statistics for 1 m Capped Composite Au Data (g/t)	14-4
Table 14 4: Correlogram Parameters for the main Kestane Beleni domains	14-8
Table 14 5: Global Model Mean Grade Gold Values (g/t) by Domain	14-14
Table 14 6: Efemçukuru Mine Mineral Resources, as of December 31, 2019	14-19
Table 15 1: Cut-off Grade Input and Calculation Summary	15-3

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Table 15 2: Planned Mining Dilution and Mining Recovery Factors	15-5
Table 15 3: Efemçukuru Mineral Reserves as of December 31, 2019	15-5
Table 16 1: Development Dimensions	16-5
Table 16 2: Rock Mass Quality	16-7
Table 16 3: Current Equipment Fleet	16-12
Table 16 4: LOM Plan Capital Development	16-13
Table 16 5: LOM Plan Production Schedule	16-14
Table 16 6: Fresh Air Requirement for Diesel Equipment	16-15
Table 17 1: Process Design Criteria of Efemçukuru Process Plant and Actual Plant Performance Data	17-4
Table 17 2: Reagent Types, Addition Points and Dosages in Efemçukuru Flotation Plant	17-7
Table 17 3: Characteristics of the Flotation Cells at Efemçukuru Flotation Plant	17-7
Table 17 4: Summary of Production Data between 2013 and 2019 in Efemçukuru Flotation Plant	17-11
Table 18 1: Surface Areas of the Dry tailings and WRD Storage Areas at Efemçukuru Gold Mine	18-4
Table 18 2: Estimated Load List – Efemçukuru	18-9
Table 20 1: Permits Obtained Prior to Mining and During Mining at Efemçukuru	20-8
Table 21 1: Mine Capital Cost Summary by Year (in US$ x 1,000)	21-1
Table 21 2: Unit costs	21-2
Table 21 3: Efemçukuru Operating Costs by Year (in US$ x 1,000)	21-3
Table 22 1: Cashflow Summary ((in US$ x 1,000)	22-2
Table 24 1: Extended LOM Production Schedule	24-2

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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
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
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

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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
Tonnes per day	t/d or tpd
Tonnes per hour	tph
Tonnes per year	tpa
Turkish Lira	₺
Volt	V
Watt	W
Weight/volume	w/v
Weight/weight	w/w

Abbreviations and Acronyms

Acidity or Alkalinity	pH
Air Pollution Control Regulation	APCR
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
Canadian Institute of Mining, Metallurgy, and Petroleum	CWMS
Cobalt	Co

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Coefficient of Variance	CV
Continuous Wastewater Monitoring System	CM
Copper	Cu
Copper Sulphate	CuSO4.5H2O
Cumulative Distribution Function	CDF
Cut-off Grade	COG
Cyanide	CN
Drift and Fill	DAF
Diamond Drill Hole	DDH
Directorate of State Hydraulic Works	DSI
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
HERCO Discrete Gaussian Model aka HERCO (Hermite Coefficient)	Herco
High Density Polyethylene	HDPE
High Grade	HG
Hydrochloric Acid	HCl
Hydrogen Oxide	H2O
Incremental cut-off Grade	ICOG
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
Kestane Beleni North West Shoot	KBNW
Kişladağ Concentrate Treatment Plant	KCTP
Kopinar Vein System Middle Domain	KPM
Kopinar Vein System South Domain	KPS
Lead	Pb
Lerchs-Grossman	L-G
Life-of-mine	LOM
Locked Cycle Test	LCT
London Metal Exchange	LME
Longitudinal Longhole Open Stoping	LLHOS
Longhole Open Stoping	LHOS
Manganese	Mn
Mechanical, Piping, Electrical, Instrumentation	MPEI
Measured & Indicated	M&I

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Mercury	Hg
Middle Ore Shoot	MOS
Ministry of Environment and Urban Planning	MEUP
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 Ore Shoot	NOS
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
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
Sequencing a Batch Reactor	SBR
Silicon	Si
Silver	Ag
Sodium Cyanide	NaCN
Sodium Hydroxide	NaOH
Sodium Isobutyl Xanthate	SIBX
Sodium Metabisulphite	Na2S2O5
Sodium Metabisulphite	SMBS
South	S
South East	SE
South Ore Shoot	SOS
South Rock Dump	SRD
South West	SW
Specific Gravity	SG
Spherical	SPH
Standard Reference Material	SRM
Strontium	Sn
Sulfur	S
Sulfur Dioxide	SO2

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Sulphide	S2-
Sulphuric Acid	H2SO4
Tailings Management Facility	TMF
Tailings Storage Facility	TSF
Technical Study	TS
Tourmaline	WMT
Transport Costs and Refining Costs	TCRC
Transportable Moisture Limit	TML
Transverse Longhole Open Stoping	TLHOS
Tuprag Metal Madencilik Sanayi Ve Ticaret Limited Sirketi	Tuprag
Turkish Electricity Distribution Corporation	TEDAS
Turkish Electricity Transmission Corporation	TEIAS
Uninterrupted Power Supply	UPS
Universal Transverse Mercador	UTM
Uranium	U
Value Added Tax	VAT
Water Pollution Control	WPC
Wardell Armstrong International	WAI
West	W
Work Breakdown Structure	WBS
Waste Rock Dump	WRD
Waste Rock Sedimentation Pond	WRSP
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 Efemçukuru 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 on the Efemçukuru gold mine to provide an updated description of the geology and mineralization, mineral resources and mineral reserves, and mine and mill operations in light of the long shelf life of the existing Technical Report whose Effective Date was August 01, 2007 (Technical Report on the Efemçukuru Project, 2007).

Information and data for this report were obtained from Efemçukuru 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., Ertan Uludag, P.Geo., Paul Skayman, FAusIMM, and Imola Götz, P.Eng., whom are all are employees of Eldorado. All of the qualified persons have visited the Efemçukuru Gold mine.

1.2

Property Description and Location

The Efemçukuru mine has been an operating underground mine in commercial production since 2011 with facilities consisting of an underground crushing plant, milling and flotation plant, filtration and paste backfill plant, and water treatment plant, along with ancillary buildings.

The Efemçukuru mine site is situated within the Aegean climatic zone, which is characterized by hot and dry summers and warm and rainy winters. The mine is located in İzmir province in western Turkey, approximately 20 km from İzmir, near the village of Efemçukuru (Figure 1-1). Access to the mine site is via approximately 40 km of paved roads from the city of Izmir to the northeast and county of Menderes to the east (Figure 1-1). Alternative accesses are from the Seferihisar highway to the west and Izmir-Kavacık road to the northeast. Water is supplied from onsite water treatment sourced largely from underground dewatering and site collection ponds, the plant requires 75 m3 of water per hour.

Power from the nation grid is connected at Çamlı village with a 34.5 kV transmission line, the 15 km line was constructed by Tüprag in 2011 and transferred to the electrical authority.

Figure 1-1: Property Location Map - Çeşme Peninsula Turkey

The Efemçukuru Project land position consists of a single operating license, number 51792, with a total area of 2261.49 ha as of December 2018. According to Turkish mining law, Tüprag retains the right to explore and develop any mineral resources contained within the license area provided fees and taxes are maintained. The license was issued on April 20, 1999 and renewed on August 19, 2013 and is currently set to expire on August 19, 2033. Within the 126.5 ha operating area, forestry land makes up about 80%, treasury land makes approximately 1%. The remaining area is private land and as of May 2018 is 100% owned by Tüprag.

No environmental liabilities have been assumed with the Project.

The Project is fully permitted with no additional permits currently required. All infrastructure required to operate under the reserves disclosed in this report fall under the scope of the existing EIA and operating license.

1.3

Geology, Drilling and Sampling

The Efemçukuru gold mine, an intermediate sulfidation epithermal vein deposit, is hosted in the center of a broadly NE-SW trending horst known as the Seferihisar Horst, which regionally exposes basement rocks of the Bornova Flysch in the Menderes Massif. The flysch predominantly comprises lower greenschist facies schist with intercalations of mudstone, fine-grained sandstone, limestone and marly sandstone. Bedding dip directions of the flysch sequence across the entire Seferihisar Horst outline a broad, asymmetric NE-trending syncline.

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Local geology consists of intermediate sulfidation veins hosted by a low-grade metamorphic sequence of very fine-grained, black to dark grey shales to phyllite and schist that have been locally folded and intruded by rhyolite dikes. Mineralogy of the phyllites is fine-grained quartz, feldspar, muscovite, chlorite and rare biotite. The mineralogy of the schist is similar to the phyllite, comprising strongly deformed quartz, feldspar, chlorite and muscovite. A calc-silicate alteration, locally termed hornfels, occurs in a broadly NW-trending pattern in the center of the deposit area. The alteration commonly occurs as alternating dark green and tan-grey bands within meta-sedimentary rocks. The contact between calc-silicate alteration and phyllite is gradational. Rhyolite occurs throughout as 1 to 5 m-wide NW-striking dikes. Contacts of the rhyolite dikes with the flysch units are usually sharp,

Two major broadly NW-SE striking epithermal vein systems occur at Efemçukuru, namely Kestane Beleni and Kokarpınar, with strike extents of approximately 2 km and 4 km respectively. Both veins cut the rhyolite dikes, calc-silicate alteration and unaltered phyllite and schist. At surface, the veins are up to 5m wide and occur as multi-phase, brecciated, banded crustiform-colloform, and massive quartz-rhodochrosite veins. The individual epithermal veins within these vein systems contain multiple ore shoots with zoned mineral and metal distributions and a complex paragenesis. The 2 km long Kestane Beleni vein hosts the major gold resource and reserve at Efemçukuru and comprises four ore shoots: South Ore Shoot (SOS), Middle Ore Shoot (MOS), North Ore Shoot (NOS) and Kestane Beleni Northwest (KBNW). The Kestane Beleni vein has a distinct mineralogical zonation with the proportions of Mn‐silicate and carbonate and sulfide vein material varying across the vein system. Mn-rich vein assemblages are most abundant in the upper portions of the SOS, whereas the sulfide content of the MOS and NOS, particularly at depth, is much higher.

The Efemçukuru gold mine has seen numerous diamond drill campaigns since 1992. A total of 903 exploration and resource delineation drillholes, drilled from surface and underground locations and totaling 216,000 m, have been drilled to 2019. Infill drilling programs, designed to increase the geologic confidence in gold grade distribution and mineralization contacts just ahead of mining, generally drill 30,000 m annually from underground stations. These programs also convert indicated resources to measured resources.

All samples from the Efemçukuru gold mine drilling are assayed for gold by 30 g fire assay with an AA finish for multi-element determination using fusion digestion and inductively coupled plasma spectroscopy (ICP) analysis. Samples that returned assays greater than 10 ppm were re-assayed by fire assay with a gravimetric finish. Eldorado employs a comprehensive QA/QC program as part of the assaying procedure, involving regular insertion of Certified Reference Materials (CRMs), duplicates and blank samples. Site geologists regularly monitor the performance of CRMs, blanks and duplicates as the assay results arrive on site. In Eldorado’s opinion, the QA/QC results demonstrate that the Efemçukuru mine’s assay database is sufficiently accurate and precise for the resource estimation.

Another form of data verification is the reconciliation to production of mined portions of the resource model. Annual reconciliation records, especially since 2016, show excellent performance between the resource model and milled production. These results clearly demonstrate that the Efemçukuru data management and QAQC protocols produce highly verifiable data that form the basis of quality resource estimation at Efemçukuru mine.

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1.4

Mineral Resources Estimates

The mineral resource estimates for Efemçukuru consist of 3D block models formed on the Kestane Beleni, Kokarpinar and Batı epithermal vein systems. Creation of these models utilized a commercial mine planning software package (Geovia Gems). Currently, mining only occurs within the Kestane Beleni vein system. Gold mineralization at Efemçukuru, primarily occurring in the principal veins, can only be confirmed through assays. Domains to control grade interpolation are, by necessity, grade based. For the Efemçukuru mineralization, creation of the modeling domains used a 2.0 g/t Au grade threshold and general vein geometry. Risk posed by extreme gold grades was examined. The examination showed a risk does exist and this risk was mitigated by a series of assay gold grade caps (40 to 200 g/t). Prior to grade interpolation, the assay data were composited into 1 m fixed length composites.

Modelling consisted of grade interpolation by ordinary kriging for Kestane Belani domains and inverse distance weighting to the second power (ID) in the remainder of the zones where data was too limited to create correlograms. Nearest-neighbour (NN) grades were also interpolated for validation purposes. No grades were interpolated outside the modeling domains. The search ellipsoids were oriented preferentially to the orientation of the vein in the respective domains. 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. The second pass allowed a single hole to place a grade estimate in any block that was uninterpolated from the first pass. 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 Efemçukuru mine were classified using logic consistent with the CIM definitions referred to in the National Instrument 43-101-Standards of Disclosure for Mineral Projects (NI 43-101). The mineralization of the project satisfies sufficient criteria such that it can be classified into measured, indicated, and inferred mineral resource categories.

Efemçukuru mine mineral resources, as of December 31, 2019, are shown in Table 1-1.The Efemçukuru mineral resources are reported at a 2.5 g/t Au cutoff grade.

Table 1-1: Efemçukuru Gold Mine Mineral Resources, as of December 31, 2019

Mineral Resource Category	Resource (t x 1,000)	Grade Au (g/t)	Contained Au (oz x 1,000)
Measured	2,555	7.93	651
Indicated	1,684	6.84	370
Measured & Indicated	4,239	7.50	1,021
Inferred	4,399	6.55	927

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1.5

Mineral Reserves and Mining Methods

The Efemçukuru mine has produced 3.7 Mt of ore at an average grade of 7.6 g/t Au as of December 2019, using combination of drift and fill and longhole open stoping methods. A planning cut-off grade (COG) of 3.68 g/t Au was calculated based on Eldorado’s mineral reserves gold price of $1250/oz, the 2020 budget costs and a steady state life of mine (LOM) production profile. The 2020 budget costs are supported by 2019 actual production costs. Use of the Deswik Stope Optimizer software identified potentially mineable material in the form of mining shapes for both drift & fill (DAF) and longhole open stoping (LHOS) mining methods. Dilution was captured as internal dilution (mining shape) and planning (overbreak). The latter equaled 16%. A mining recovery factor of 96% was also implemented. Both of these factors are defended by regular reconciliation and stope closure exercises.

The mineral reserves of the Efemçukuru Project were classified using logic consistent with the CIM definitions referred to in the National Instrument 43-101-Standards of Disclosure for Mineral Projects (NI 43-101). The mineralization of the project satisfies sufficient criteria to be classified into proven and probable mineral reserves. Only measured and indicated mineral resources were converted, using appropriate modifying factors, to mineral reserves. The mineral reserves are inclusive to the mineral resources.

The mineral reserve estimate is summarized in Table 1-2 and has an effective date of December 31, 2019.

Table 1-2: Efemçukuru Mineral Reserves Effective December 31, 2019

Category	Ore (t x 1,000)	Grade Au (g/t)	Contained Au (oz x 1,000)
Proven	1,996.0	6.77	434.2
Probable	1,383.4	5.68	252.4
Proven & Probable	3,379.4	6.32	686.6

Efemçukuru mine employs small-scale underground mechanized mining methods to exploit the narrow, high- grade, subvertical mineralization. The projected mine life is 7 years at current production rate of 510,000 tpa.

The current mine layout has the following features:

●

Four declines (SOS, MOS, NOS and KBNW), each covering approximately 400 m strike extent.

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Two surface portals (south and north).

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One surface conveyor adit for conveying crushed ore to the surface crushed ore bins.

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Four primary ventilation surface exhausts (south, central, north and northwest) and one fresh air raise for NOS.

●

Link drives connect declines and serve as a secondary egress from the mine.

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The mine plan is based on the combination of drift-and-fill and longhole-open-stoping methods. Both DAF and LHOS stopes are mined concurrently from multiple production blocks to fulfil production requirements. Overall, the production blocks are mined in a top-down sequence, but stopes within a production block are mined bottom-up (overhand).

A geotechnical domain model has been developed and updated for geotechnical logging of exploration and stope definition drilling information. At Efemçukuru the rock mass has been classified by the widely used Q-System by adopting characterization logging values to determine Q input parameters. The selection of DAF and LHOS mining methods is primarily based on the orebody geometry (width and dip) and the expected ground conditions determined through geotechnical assessment. Regular geotechnical assessments indicate that the current mining method, stope sizes, and mining sequence will not change significantly change over the LOM.

The mine operates seven days a week and three shifts a day. This annual schedule is equivalent to 365 days per year of operation

1.6

Mineral Processing and Recovery Methods

The Efemçukuru operation is an underground mine with facilities consisting of an underground crushing plant, milling and flotation plant, filtration and paste backfill plant, and water treatment plant, along with ancillary buildings. The process plant produces a gold-containing bulk sulphide rich flotation concentrate. Major sulfide minerals comprise pyrite, sphalerite and galena. Small percentage of chalcopyrite is also present. Major gangue minerals consist of quartz and manganese-minerals. Gold is generally fine-grained and primarily associated with pyrite and galena. Pyrite and sphalerite generally show good liberation, with improving liberation for the fine size fractions. Galena also shows improving liberation for the fine size fractions.

Optimum primary grind size was determined to be 80% passing 63 µm. The reagents used in flotation were; sodium bisulfite (NaHS) as sulfidizing agent, copper sulfate (CuSO4.5H2O) as activator, xanthate (SIBX) as collector, S-8045 as promoter and OrePrep F-549 as frother. In most cases, gold recovery is proportional to sulfur recovery and has averaged about 93 to 94% in recent years.

Run-of-mine ore is crushed underground and transferred to two ore storage bins on surface via a conveyor. The two ore storage bins allow for blending of different ore types feeding the process plant feed to target a desirable gold/sulfur ratio and reduce contents of penalty elements for concentrate sales.

The comminution circuit consists of a semi-autogenous grinding (SAG) mill operated in closed circuit with a pebble crusher, a ball mill operated in closed circuit with hydrocyclones and a flash flotation cell.

Ball mill discharge is treated in a flash cell to recover the fast-floating liberated sulfide mineral particles and prevent overgrinding of gold containing particles. Overflow from hydrocyclones is sent to a rougher/scavenger flotation bank.

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Flotation circuit consists of a rougher/scavenger flotation bank and two parallel cleaner flotation banks. Concentrates from flash flotation cell and the first two cells of the rougher/scavenger bank are combined and upgraded in cleaner bank 1. Rougher cells 3-6 concentrate are treated in cleaner bank 2. Concentrates from cleaner banks 1 and 2 are combined and sent to the final concentrate thickener.

Underflow of the concentrate thickener is filtered and the filtered concentrate is stored in big bags for shipping. The tailings are sent to a tailings thickener. The final tails are filtered. A portion of the tailings is used in the underground paste backfill plant, and the rest is dry stacked in the tailings storage facility (TSF).

Column flotation as the third cleaner flotation stage is expected to increase concentrate quality and reduces concentrate tonnage with negligible gold recovery loss. A project was initiated in late 2019 for design and installation of column flotation cells in the circuit. Based on mass balance and simulation studies performed using plant surveys and pilot scale column flotation tests, two flotation column cells of 1.8 m (diameter) x 8 m (height) will be installed and commissioned in 2020. The two column flotation cells can be operated in parallel or in series as the third cleaner flotation stage depending on plant requirements.

1.7

Project Infrastructure

The Project infrastructure is well established for LOM purposes.

Management of the site water will use the existing ponds and an additional pond at the tailings management facility (TMF). The water treatment plant is appropriately sized to include the new facilities. The constructed areas will be sloped and ditched appropriately to tie into the existing systems.

Existing ancillary buildings will continue to be utilized such as the warehouse and administration buildings. A new change room facility will be installed for the personnel that will be working in the mill.

1.8

Environmental

Tüprag conducted baseline studies in throughout the early 2000’s prior to development. An EIA was submitted in 2005 and was approved with Environmental Positive Certificate being granted in September 2005. Since mining began in 2011, Efemçukuru mine operations have routinely collected environmental data outlined in the Environmental Management Plan (EMP) and submitted data to the relevant government agencies.

Tüprag submitted applications for revisions to the EIA and received approvals for the revisions in 2012 and 2015 to allow for larger facilities. Amendments to the Environmental Licenses were requested and granted in 2012, 2015, 2017, and 2018.

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1.9

Capital and Operating Costs

Efemçukuru is fully constructed and operating, and actual costs form the basis of future operating and sustaining cost estimates. All costs are expressed in US dollars.

Mining sustaining capital costs include mine development, paste backfill borehole development, purchase of additional equipment, equipment leasing costs and health and safety initiatives (Table 1-3).

Additional growth capital cost is included for the supply and install of a column flotation system to improve the concentrate quality. The column cells will be commissioned in 2020.

Table 1-3: Capital Cost Summary

Sustaining Capital	LOM Total (US$ x1,000)
Mine Development	7,304
Mine Infrastructure & Equipment	39,909
Process Plant	9,520
Exploration	548
Health & Safety	345
G&A	1,097
Total	58,723
Growth Capital
Process Column Flotation	1,553

The underground mine operating costs were estimated based on actual 2019 operating costs and 2020 budget estimates that allow for maintaining a steady state production profile.

The underground operating costs include all consumables (ground support, explosives, services, cement, aggregates and fuel) and equipment required to meet the development and production schedule objectives. The operating unit costs for mobile equipment and fuel consumption rates were largely obtained from historic mine data. Labour requirements were developed to support the operation and maintenance of the fleet and for the general operation of the underground mine. All these estimates are in line with manpower levels.

General and administrative costs are based on current personnel requirements and salaries. Adjustments have been made if known changes, such as increasing manpower, are required in future. General supplies are based on the current operating experience.

Process operating costs were based on current annual consumption of process reagents, major wear parts, and utilities. Budget quotations were obtained for supply of all significant consumables and utilities. Power consumption is based on 2019 operating experience.

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Unit rates representative to a steady state peak production profile are summarized in Table 1-4.

Table 1-4: Operating Cost Summary

Area	Unit Costs (US$/t)
Mining	33.28
Operating Development	5.66
Geology	4.77
Plant Maintenance	8.54
Risk Management	3.59
Processing	21.31
Administration	19.23
Finance	6.93
Purchasing	1.63
Total	104.93

1.10

Economic Analysis

An economic analysis of the Efemçukuru mine shows a 7 year mine life yielding an undiscounted cumulative cashflow of US$341.2 M and an NPV of US$278.3 M and are estimated based on the following:

●

Discount rate of 5%.

●

Development cost of US$2,154 per metre of waste development.

●

Operating cost of US$38.05 per ore tonne.

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Total gold recovery of 87.5 % after processing, smelting and refining.

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Process plant throughput of 510,000 tonnes per annum

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Gold price of US$1,400 per ounce.

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Silver price of US$18.00 per ounce.

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Transport and refining cost of US$118.48 per mined Au ounce.

●

Capital costs outlined in Table 1-3

●

Pre-tax basis

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 (Figure 1-2 and Figure 1-3).

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Figure 1-2: LOM Plan NPV Sensitivities based on Au Price Variation

Figure 1-3: LOM Plan NPV Sensitivities based on Cost Variations

The Project is most sensitive to gold price. Changes in the other factors do not significantly impact the viability of the Project.

1.11

Interpretations and Conclusions

It is concluded Efemçukuru has a solid working history and that the work completed in the technical study indicate that the mineral resource and mineral reserve estimates and Project economics are sufficiently defined to indicate that the Project remains technically and economically viable.

The qualified persons have a high degree of confidence in the contents of this technical report.

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1.12

Recommendations

The technical study outlines a solid technical and economical assessment for the Efemçukuru mining operation. It is recommended to proceed studies to further optimize and improve operations.

●

Continue with drilling programs to seek ancillary vein systems near strike and other targets in the region.

●

Study narrow vein mining techniques for thinner structures.

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Investigate ore sorting and dense media separation to minimize dilution.

●

Continue studies for process solutions to upgrade the gold concentrate.

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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 Efemçukuru 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 Efemçukuru gold mine to provide an updated description of the geology and mineralization, mineral resources and mineral reserves, and mine and mill operations in light of the long shelf life of the existing Technical Report whose Effective Date was August 01, 2007 (Technical Report on the Efemçukuru Project, 2007).

Information and data for this report were obtained from Efemçukuru 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 this technical report. In order 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 Efemçukuru 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.

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 Technical Report are David Sutherland, P.Eng., Paul Skayman, FAusIMM, Ertan Uludag, P.Geo., Imola Götz, P.Eng. and Sean McKinley, P.Geo., whom are all employees of Eldorado.

David Sutherland, Project Manager, was responsible for overall preparation of the technical study and sections related to infrastructure and environment (report sections 1, 2, 3, 4, 5, 6, 18, 20, 24, 25, 26 and 27). He most recently visited the Efemçukuru gold mine on May 18, 2018.

Imola Götz, Manager, Mine Engineering (Underground), 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). She most recently visited the Efemçukuru gold mine on September 14 to 18, 2019.

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 Efemçukuru gold mine on February 6 to 7, 2019.

Ertan Uludag, Resource Geologist, 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 11, 12 and 14). He most recently visited the Efemçukuru gold mine on June 10 to 11, 2019.

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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 Efemçukuru gold mine on November 18, 2019.

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

Currency used is US$ throughout, unless otherwise 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
Efemçukuru	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

Tüprag, a wholly owned subsidiary of Eldorado, is operating the Efemçukuru gold mine and related processing facilities on its mineral licenses near the village of Efemçukuru of Menderes District, Izmir Province, in the Aegean Region of Turkey.

The Efemçukuru underground gold mine has been in commercial production since 2011. Surface facilities consist of the processing plant, tailing filter plant, pastefill plant, tailings storage area, waste rock dump, water treatment plant, sedimentation ponds and administrative facilities.

Ore is mined using conventional mechanized cut and fill and longhole mining methods. Ore is processed to produce a flotation concentrate for sale to refiners.

4.2

Property Location

Efemçukuru is located in İzmir province in western Turkey, approximately 20 km from İzmir, near the village of Efemçukuru (Figure 4-1). The village of Efemçukuru has a population of approximately 650 people and lies 700 m southwest of the project area. Izmir, Turkey’s third largest city, is located along the Aegean Sea and has major port facilities available to service the Project. Izmir is an industrial centre with a population in excess of 4 million people.

Approximate project co-ordinates are:

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UTM 497 850 E and 4 238 425 N

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UTM Zone 35S

●

Map Sheet İZMİR L17 – İZMİR L17b3 (1:25,000 Scale)

●

Longitude E 26° 58' 30.30"

●

Latitude N 38° 17' 31.59"

Land use within the concession area falls into three categories: agricultural land (vineyards), forestry land and treasury land. Forestry land makes up about 80% of the project area (126.5 ha) and treasury land makes approximately 1%. The remaining area belongs to private land. As of May 2018, Tüprag is the owner of 100% of the private land within the operation area.

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Figure 4-1: Location Map Showing Project Location

4.3

Land tenure

The Efemçukuru Project land position (Figure 4-2) consists of a single operating licence, number 51792, with a total area of 2261.49 ha as of December 2018. 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 20, 1999 and renewed on August 19, 2013 and is currently set to expire on August 19, 2033. The duration of a 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: Efemçukuru Land Position

4.4

Royalties

Mining licences in Turkey are divided into 5 groups. The Efemçukuru licence is in group 4 which includes gold, silver, and platinum mines. Royalty rates for group 4 licences are calculated on a sliding scale implemented in 2019. Royalty rates (Table 4-1) are based on the run of mine (ROM) sales price. The ROM sales price is calculated by subtracting processing, transport, and depreciation costs from the gold and silver revenues. This amount is then multiplied by the appropriate royalty rate. The royalty rates are determined once a year by the General Directorate of Mines based on the average sales price of gold and silver quoted on the London Metal Exchange (LME). Royalty rate for Efemçukuru in 2019 was 7%.

Concentrate produced at the Efemçukuru mine is considered the product of ore processing and is eligible for the 40% reduction in the royalty rate.

4.5

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 4-2 lists key Project permits obtained to date, including the date and the governmental authority that issued them.

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Table 4-1:Royalties Calculation

Royalty	Gold (Annual average price)	Silver (Annual average price)
(%)	(US$/oz)	(US$/oz)
1	<800	<10
2	801-900	11-12
3	901-1000	13-14
4	1001-1100	15-16
5	1101-1200	17-18
6	1201-1300	19-20
7	1301-1400	21-22
8	1401-1500	23-24
9	1501-1600	25-26
10	1601-1700	27-28
11	1071-1800	29-30
12	1801-1900	31-32
13	1901-2000	33-34
14	2001-2100	35-36
15	<2101	<37

Table 4-2: Key Project Permits

Name of Permit	Issue Date	Issuer
EIA Permit	9/8/2005	Ministry of Environment and Urbanization
Energy Permitting	6/21/2010	TEDAŞ
Sewage Water Treatment Plant Design Approval	6/25/2010	Ministry of Environment And Urbanization
Private Security Permission	12/24/2010	Governor of İzmir
Provisional Acceptance Certificate of Bademler Substation	3/21/2011	TEDAŞ
Provisional Acceptance Certificate of Mine Site Electrical Installation	4/12/2011	TEDAŞ
EIA Capacity Expansion Permit	12/31/2012	Ministry of Environment and Urbanization
Workplace Opening Permit	2/27/2013	Governor of İzmir
Operation Licence	8/19/2013	Ministry of Energy and Natural Resources
Operating Permit	8/19/2013	Ministry of Energy and Natural Resources
Underground Explosive Magazine	9/18/2013	İzmir Security Directorate
Temporary Waste Storage Area Permit	8/6/2014	Ministry of Environment And Urbanization
After upgrading switchgear in Yelki Substation, new Provisional Acceptance Certificate	8/27/2014	TEDAŞ
EIA Capacity Expansion Permit	11/17/2015	Ministry of Environment and Urbanization
Forestry Permit Exploration Drilling and Drilling Access Road	5/4/2017	Ministry Forest And Water Affairs
Forestry Permit Power Line	5/4/2017	Ministry Forest And Water Affairs
Forestry Permit Operation	5/4/2017	Ministry Forest And Water Affairs
Purchase and Use of Explosive Certificate	11/8/2017	İzmir Security Directorate
Waste Water Treatment Plant Design Approval	7/25/2018	Ministry of Environment And Urbanization
Waste Rock (Uneconomical Rock Dump Area) Storage Facility	1/12/2019	Ministry of Environment And Urbanization
Environmental License	5/16/2019	Ministry of Environment And Urbanization

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SECTION ● 5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1

Access and Infrastructure

Access to the mine site is via approximately 40 km of paved roads from the city of Izmir to the northeast and county of Menderes to the east (Figure 5-1). Alternative accesses are from the Seferihisar highway to the west and Izmir-Kavacık road to the northeast. A 2 km paved mine road currently provides access to the mine property from the regional roads. The travel time by road from Izmir is 45 to 60 minutes.

Figure 5-1: Project Accessibility Climate

Power is supplied to the mine via a 10 MW power line that originates from the Urla 2TM station. There are emergency generators with 8.5 MW of capacity located at the mine site to supply power to critical units at the mill as well as water discharge and ventilation at underground operations.

The proximity of the site to Izmir, one of the largest industrial centers in Turkey, and the second largest port, is advantageous to the project for concentrate shipments and material supplies as well as a source of qualified personnel. The workforce from nearby villages represents 52%.

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The Efemçukuru mine site is situated within the Aegean climatic zone, which is characterized by hot and dry summers and warm and rainy winters. Seasonal rains occur from November to April, with annual precipitation of 600 to 800 mm, while annual evaporation may reach up to 1,500 mm. Under these climatic conditions, Kokarpinar Stream is generally intermittent and stream flow is restricted to the wet season in response to precipitation, or to short-term storm events during the summer.

Average temperatures in the region range between 30ºC in summer and -5ºC in winter with an annual average of approximately 17ºC. The study area is susceptible to orographic effects caused by the lifting of moisture-laden air from the Aegean Sea. Accordingly, the study area experiences a significant amount of rainfall variation on a monthly basis.

Long-term climatic records are not available for the Kocadere River catchment. Although a meteorological station does exist at the Efemçukuru mine site, it has only been in operation since 1998. The next closest meteorological station located at Beyler, approximately 7 km to the southwest, has a longer period of record. However, the lack of data overlap prevents correlation with the meteorological data recorded at the mine site. The next closest stations with overlapping periods of record are Balçova, roughly 14 km to the northeast; Seferihisar, approximately 15 km to the southwest and Izmir, approximately 20 km to the northeast.

Owing to the relatively long period of record, and strong correlation with data collected at the mine site, the precipitation record from the Seferihisar meteorological station (1975-2010) has been used in the evaluation of long-term precipitation conditions at the Efemçukuru mine site. Table 5-1 presents the expected seasonal variation of monthly climate data for the study area. The expected annual average precipitation is 610 mm, while precipitation extremes for wet and dry years (1:100 year return period) are respectively 1,255 mm and 383 mm.

Average annual precipitation is 750 mm due to the moderating influence of the Aegean Sea (Table 5-1). There is limited snowfall. Average high wind velocity is 30 km/h with a maximum of 50 km/h.

5.2

Physiography

The Efemçukuru Project is located at the western end of the Izmir-Ankara Suture Zone, on a major uplifted structure that extends northeast-southwest at the southwest from Izmir. Steep hills and narrow valleys characterize the project site with elevation on site ranging from 540 masl in the valley to 770 masl in the surrounding hills. The deposit crops out at Kestane Beleni hill which slope steeply to the Kokarpinar creek valley. Majority of mine facilities are also located on the same slope to Kokarpinar creek (Figure 5-2). Vegetation consists of mature pine trees with sparse undergrowth covering the hillsides. The flatter land in the valleys and upper slopes of the hills has been cultivated with grape vines.

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Table 5-1: Distribution of Annual Climate Data

Month	Temp	Precipitation		Run-Off (mm)			Potential
							Evapotranspiration
	°C	mm	% of	Rc* =	Rc =	Rc =	mm	% of
			Annual	0.4	0.55	0.8		Annual
January	8.2	108.4	17.7%	43.4	59.6	86.7	0.0	0.0%
February	8.4	84.2	13.8%	33.7	46.3	67.4	0.0	0.0%
March	10.5	68.3	11.2%	27.3	37.6	54.6	4.8	0.3%
April	14.2	43.3	7.1%	17.3	23.8	34.6	106.7	7.4%
May	9.0	23	3.8%	9.2	12.7	18.4	170.9	11.8%
June	24.0	3.6	0.6%	1.4	2.0	2.9	237.7	16.4%
July	26.6	0.4	0.1%	0.2	0.2	0.3	285.4	19.7%
August	26.1	0.6	0.1%	0.2	0.3	0.5	297.8	20.6%
September	22.1	13.9	2.3%	5.6	7.6	11.1	181.1	12.5%
October	17.5	45.9	7.5%	18.4	25.2	36.7	109	7.5%
November	12.8	87.8	14.4%	35.1	48.3	70.2	48.9	3.4%
December	9.8	132.0	21.6%	52.8	72.6	105.6	5.9	0.4%
Totals	N/A	611	100%	245	336	489	1,448	100%

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Figure 5-2: Efemçukuru Mine Facilities Layout

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SECTION ● 6 History

Efemçukuru Deposit was discovered by Tüprag in 1992, while carrying reconnaissance work in western Turkey. The area was noted as a base metal and manganese occurence on records of MTA (Turkish Mineral Exploration Institute) with its old mine workings. Surface evidence of these workings has been found in the form of shallow excavations in the main Kestane Beleni Vein structure as well as Kokarpınar Vein. After Tüprag’s discovery during exploration and early mining stages ancient workings were further identified and it was concluded that the deposit was probably mined during the Roman dynasty two thousand years ago. Later, in early 1900’s a British Company owned the exploration rights and did a limited amount of surface work on Efemçukuru deposit.

Between 1992 and 1997, Tüprag conducted extensive exploration work including a magnetic survey and mapping, soil, rock chip, and channel sampling and surface trenching, and in excess of 12,000 m of HQ drilling. The exploration work identified a high-grade vein-hosted gold system consisting of three separate ore zones along the Kestane Beleni structure known as the SOS, MOS, and NOS. The Kestane Beleni NW shoot was named as Mezarlık Tepe earlier, but later work done in the area has shown that it is the continuation and part of the same major Kestane Beleni structure. A metallurgical testwork program was completed to support a conceptual study in 1994 by Tüprag, described a 1,000 t/d underground operation using CIL ore processing.

In 1998, Micon evaluated the geological model and confirmed a measured and indicated resource of 1.87 Mt at 14.26 g/t, with an inferred resource of 660,000 tonnes at 11.99 g/t Au. Later, in 2007 Wardrop Engineering came up with a resource measured and indicated resource of 3.88 Mt at 11.2 g/t, with an inferred resource of 753,000 tonnes at 8.79 g/t Au.

Between 2006 and 2008, a 30,349 m HQ infill drilling program was undertaken along the SOS, MOS, and North Ore Shoot (NOS) to further delineate the initial identified resource. Additional diamond drilling was carried out for hydrogeological testing in the vein structure as well as the hanging wall and foot wall rocks.

Additional drilling programs from 2008 to 2010 with step-outs to deeper levels and along strike significantly increased the mineral resource estimate and provided a base for calculating a mineral resource estimate for the NOS. Drilling in 2011 and 2012 focused on a new zone along strike from the NOS, referred to as Kestane Beleni Northwest (KBNW), on down-dip extensions to the SOS and on the Kokarpinar vein, a parallel vein system located approximately 400m east of the Kestane Beleni vein. Exploration drilling programs in 2013 through 2017 tested the Kokarpinar vein over a 3 km strike length, and identified resources in several discrete shoots.

In 2014 and 2015, soil sampling was extended to cover the area west of the Kestane Beleni vein. New vein targets hosted in phyllite were identified in the Karabag and Dedebag areas as a result of this work. The veins are particularly anomalous in silver.

In 2017 approximately 19,765 m of drilling was completed from the Kestane Beleni Drift, to test the extent of the mineralization especially at deeper parts of MOS and SOS-MOS transition area, and Kokarpinar vein. In 2018, 22,867 m of exploration drilling was completed targeting to expandthe resource in the Kokarpinar vein and Bati Veins located at the footwall of the Kestane Beleni vein which was discovered later in the year. 4,944 m were also drilled to delineate a mineralized zone at Kestane Beleni NW shoot.In 2019, at Bati and Kokarpinar Veins a total of 26,084 m of exploration and 5,827 m of delineation drilling on Kokarpinar vein was completed.

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Eldorado is continuously exploring on the property while mining the Kestane Beleni Ore Body since 2011. Table 6-1 summarizes the drilling at different mineralized ore bearing structures at Efemçukuru. Figure 6-1 shows the current drilling program in progress.

Table 6-1: Summary of Drilling on the Efemçukuru Deposit

PRE-MINING
Location	Type of Drilling	Year	# of Holes	Meters
Kestane Beleni Vein	Core	1992-2011/1	299	51,052
	RC		58	5,028
Kokarpınar Vein	Core		16	4,340
	Percussion		8	393.8

POST-MINING
Location	Type of Drilling	Year	# of Holes	Meters
Kestane Beleni Vein	Core	2011/2-2018	194	48,603
Kokarpınar Vein			232	70,107
Batı Vein			87	34,781
Others			9	1,512
Total			903	215,816

Figure 6-1: Current Drilling Program Site

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Permitting for the project was initiated in 1998 and an Environmental Impact Assessment (EIA) study was completed in May 2004. The EIA was submitted to the Ministry of Environment and Urbanization in 2005, and the environmental positive certificate was issued in September of that year. Subsequent to completion of the EIA, a revision was approved in December 2012, allowing for a larger disturbance footprint and an increased mining production rate of 600,000 tonnes per annum.

The Efemçukuru mine started commercial production in June 2011, and ore production ramped up from 435 kt to 520 kt with no major changes on the operation.

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SECTION ● 7 Geological Setting and Mineralization

7.1

Regional Geology

Western Anatolia, Turkey, is host to several major porphyry and epithermal gold deposits including the Kişladağ porphyry gold mine (17 Moz), Efemçukuru intermediate sulfidation gold mine (2.5 Moz) and Ovaçik (3.9 Moz) low sulfidation gold mine (Baker, 2019). The gold-rich region is part of the Western Tethyan orogen defined by a series of Cretaceous through Cenozoic magmatic belts that have a strike length of over 3,500 km from Slovakia to Iran, and then continue to the east through the Central and Eastern Tethyan orogen. The magmatic belts in Turkey are broadly younger to the south from Cretaceous to Paleogene subduction-related arc magmatism in the Pontides, through to post-collision, extension-related Neogene magmatism in central and western Anatolia formed as a result of slab-roll back (Jolivet et al., 2015).

During slab-roll back the Hellenic trench retreat relative to the Cyprus trench diverged and consequently tore the Aegean slab, resulting in the upwelling of hot asthenosphere beneath western Anatolia between 15 and 8 Ma (e.g., Dilek and Altunkaynak, 2009; Jolivet et al., 2015).The combination of slab roll-back, tearing and continental rifting facilitated the exhumation of metamorphic core complexes such as the Menderes Massif and Cyclades (Figure 7-1; van Hinsbergen, 2010; Jolivet et al., 2015). Miocene magmatism in western Anatolia was emplaced in NE-trending extensional to transtensional continental rifts during this exhumation and associated crustal thinning (e.g., Dilek and Altunkaynak, 2009; Ersoy et al., 2010). Structurally-controlled epithermal, porphyry and base metal deposits formed along detachment faults and WNW- to NE-trending grabens and faults and their intersections during Neogene tectonic denudation in the Menderes Massif region (Figure 7-1; Gessner et al., 2017; Menant et al., 2018).

The Efemçukuru deposit is hosted in the center of a broadly NE-SW trending horst known as the Seferihisar Horst, which regionally exposes basement rocks of the Bornova Flysch in the Menderes Massif (Boucher, 2016). The flysch predominantly comprises lower greenschist facies schist with intercalations of mudstone, fine-grained sandstone, limestone and marly sandstone. Locally, lenses of ophiolitic basalt (spilite) and serpentinite, tens to hundreds of meters thick, are interbedded with limestone and marly sandstone segments. Although the Bornova Flysch within the Seferihisar Horst is non-fossiliferous, other stratigraphically similar sequences of the flysch contain abundant faunal records that provide age constraints. In these, Upper Jurassic to Lower Cretaceous pelagic limestone is overlain by ophiolitic sequences of spilite and serpentinite in an Upper Cretaceous to Lower Paleocene siliciclastic sequence (Okay and Altiner, 2007). Bedding dip directions of the flysch sequence across the entire Seferihisar Horst outline a broad, asymmetric NE-trending syncline.

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Note: Numbered deposits and main commodity in parentheses are: 1 = Efemçukuru (Au); 2 = Kalecık (Hg); 3 = Arapdağ (Au, Ag, Pb, Sb); 4 = Ovacık (Au, Ag); 5 = Küçükdere (Au); 6 = Balya (Au, Ag, Pb, Zn); 7 = Sındırgı (Au, Ag); 8 = Kişladağ (Au); 9 = Tepeoba (Cu, Mo, Au); 10 = Ayazmant (Fe, Cu); 11 = Kalkan (Fe); 12 = Pactolus (modern day Gediz) River (historic Au); 13 = Irlamaz-Manisa (Au); 14 = Sart (Au)

Figure 7-1:Simplified Tectonic Map of the Western Anatolian Extensional Province, outlining Major Structures, Geologic Units, and Locations of Mineral Deposits (from Boucher, 2016)

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7.2

Local Geology

The intermediate sulfidation veins at Efemçukuru are hosted by schist and phyllite of the Bornova Flysch (Figure 7-2). On the property the Bornova Flysch comprises a low-grade metamorphic sequence of very fine-grained, black to dark grey shales to phyllite and schist that have been locally folded and are commonly calc-silicate altered. Phyllitic rocks locally contain lenses of fine-grained sandstone, typically with an increased abundance of quartz-carbonate veinlets, and sequences of black carbonaceous silty mudstones. Mineralogy of the phyllites is fine-grained quartz, feldspar, muscovite, chlorite and rare biotite. The mineralogy of the schist is similar to the phyllite, comprising strongly deformed quartz, feldspar, chlorite and muscovite. Spilitic basalt lenses within the schist are fine- to medium- grained and strongly chlorite altered. Larger lenses or blocks of fine-grained, massive white marble also occur locally within the schists.

Figure 7-2:Geological Map of the Efemçukuru Deposit

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A calc-silicate alteration, locally termed hornfels, occurs in a broadly NW-trending pattern in the center of the deposit area. The alteration commonly occurs as alternating dark green and tan-grey bands within meta-sedimentary rocks. The banded texture is characterized by up to 10-20 cm wide alternating bands of very fine-grained quartz and fine-grained chlorite, actinolite, epidote and clinozoisite. The calc-silicate assemblage also occurs in veins and veinlets that cut small-scale folds and locally overprints rhyolite dikes. The contact between calc-silicate alteration and phyllite is gradational, and mapped as transitional hornfels, as rocks become progressively less altered and the banding texture disappears.

Rhyolite outcrops in the deposit area occur as 1 to 5 m wide NW-striking dikes. These rhyolites are creamy white in color, fine-grained with a granular texture, and include 1 to 3 mm round quartz phenocrysts in a groundmass of orthoclase, quartz, and minor plagioclase. Sulfides are rare, except adjacent to crosscutting veins or stockwork veinlets, where up to 0.5% disseminated pyrite can occur in weakly to moderately silicified rhyolite. Contacts of the rhyolite dikes with the flysch units are usually sharp with little associated visible alteration or faulting. Boucher (2016) sampled a rhyolite dike from the deposit and obtained a U-Pb zircon age of 13.17 ± 0.25 Ma which provides an upper age for the epithermal mineralization at Efemçukuru.

Polymictic breccias also form dikes particularly in the Goktepe region and in Kokarpinar NW where they cut rhyolite. Polymictic breccia is a dark green, matrix supported breccia with angular to sub-round clasts of schist, phyllite and spilite and more rarely, round to angular clasts of rhyolite and epithermal vein fragments. Petrographic analysis of the breccia reveals that the groundmass comprises aphantic quartz, interstitial chlorite and very fine-grained (< 1 mm) euhedral to subhedral pyrite and galena. Total sulfide content is typically less than 0.5 %, and these are disseminated throughout groundmass.

Two major broadly, NW-SE striking epithermal vein systems occur at Efemçukuru, namely Kestane Beleni and Kokarpınar, with strike extents of approximately 2 km and 4 km respectively (Figure 7-2). Both veins cut the rhyolite dikes, calc-silicate alteration and unaltered phyllite and schist. At surface, the veins are up to 5m wide and occur as multi-phase, brecciated, banded crustiform-colloform, and massive quartz-rhodochrosite veins. Stockwork veins are commonly associated with the major epithermal veins and occur both in the footwall and hangingwall, as well as between vein splays, in zones of at least three (3) m wide. The stockwork zones manifest as thin (< 3 cm) Fe-oxide stained quartz veinlets cutting strongly to moderately illite and smectite altered phyllites over widths. Minor (< 2%), fine-grained disseminated pyrite characterize the stockwork zone. The bulk of the epithermal veins at surface contain Mn- and Fe-oxide coated, banded quartz-rhodochrosite with pyrite-galena-sphalerite. Evidence that indicates the major veins exploited fault zones includes the common presence of lithified cataclasite proximal to the veins and clasts of cataclasite within the veins. Furthermore, the vein‐hosting fault network was likely coeval with the calc-silicate alteration event because calc-silicate (epidote) partly cements the cataclasite and the distribution of the hornfels broadly maps out the fault-vein network.

In detail, the two main veins have complex geometries with multiple shoots and splays. The Kestane Beleni vein is divided into several ore shoots: South Ore Shoot (SOS), Middle Ore Shoot (MOS), North Ore Shoot (NOS) and Kestane Beleni Northwest (KBNW; Figure 7-2). The southern portion of Kestane Beleni system, including the SOS and MOS, dips moderately to steeply (45-70°) to the eastnortheast. In the northern zones at the NOS and KBNW, the vein has a more moderate dip (55-60°) to the northnortheast (020° to 025°). In the footwall to the Kestane Beleni vein occur two similarly oriented but narrower veins called the Bati veins (Figure 7-3). The Kokarpınar vein has a more consistent northwesterly strike and dips moderately to the northeast. In the furthest northwest extent of Kokarpinar the veins display a complex horsetail geometry with numerous splays.

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Figure 7-3:Cross Section through the Bati, Kestane Beleni (MOS) and Kokarpinar Veins

The thickest vein segments typically coincide with areas where the controlling fault system has a more northerly strike and steeper dip. NW‐striking fault segments lack veins, or have thin vein material. Such patterns are consistent with an oblique east side down normal‐dextral (right lateral) shear sense on the overall fault‐vein network, with northwest striking segments accommodating a dominantly right lateral component, and north striking steeply dipping segments accommodating dominantly normal and extensional shear sense leading to preferential vein development.

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SECTION ● 8 Deposit Types

8.1

Deposit Geology

The epithermal veins at Efemçukuru contain multiple ore shoots with zoned mineral and metal distributions and a complex paragenesis. The 2 km long Kestane Beleni vein hosts the major gold resource at Efemçukuru and comprises four ore shoots: South Ore Shoot (SOS), Middle Ore Shoot (MOS), North Ore Shoot (NOS) and Kestane Beleni Northwest (KBNW) (Figure 8-1).

The SOS outcrops for 500 m and dips between 45° to 70° to the northeast. The shoot has a down dip extent of approximately 500 m. The southern half of the shoot consists mostly of a single vein whereas in the central part of the SOS, the vein splays at depth. A second larger splay develops nearer the surface, close to the mid-point of the SOS and continues for approximately 230 m to the north end of the shoot. Where the SOS consists of a single vein, its thickness generally ranges from 3 m to 5 m but can locally reach more than 10 m. Gold mineralization is not generally distributed across the whole vein, but more typically occurs as high grade (> 50 g/t x m) discrete zones within the vein that have a steep northerly dominant plunge and shallow to moderate southerly secondary plunge. Where the vein breaks into splays, the splays are generally narrower, with thicknesses of 1 to 2 m. Additionally, significant stockwork type mineralization is locally present between the vein splays where they cut hornfels. Limited amounts of stockwork mineralization occur where the vein hanging wall consists of phyllite, but are more restricted in size and continuity.

In the MOS area, the Kestane Beleni vein strikes 320° over approximately 230 m and has an average dip of 60° to 65° to the northeast. High grade gold mineralization in the MOS forms a steeply plunging shoot in the central part of the MOS and is hosted entirely within hornfels. The shoot has a narrow surface expression of ~3 m, but expands with depth to more than 20 m. A single splay diverges from the main vein in the central part of the shoot and an extensive zone of stockwork mineralization occurs between the two veins and extends into the hanging wall. The stockwork zone is best developed between 550 m and 600 m elevation with some extensions above and below these elevations, and is traceable along strike for approximately 75 m.

The NOS is separated from the MOS by an inflection that coincides with an additional 25° westward bend in the strike of the Kestane Beleni vein. Here the vein strikes approximately 300° and vein flattens to a 45°dip. The NOS outcrops intermittently for approximately 200 m along this bearing before it pinches out. Another inflection occurs between the NOS and KBNW, with the latter striking nearly east west over a 300 metre strike length before reverting to a 300° strike orientation. KBNW is typically narrower than the other three main shoots (< 3 m) and comprises two vein splays for much of its length.

The Kestane Beleni vein has a distinct mineralogical zonation with the proportions of Mn‐silicate and carbonate and sulfide vein material varying across the vein system. Mn-rich vein assemblages are most abundant in the upper portions of the SOS, whereas the sulfide content of the MOS and NOS, particularly at depth, is much higher. The increased sulfide content coincides with an increase in Pb and Zn in the MOS and NOS and the Ag/Au ratio is typically higher at depth and on the periphery of the veins.

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Figure 8-1: Long Section with Gold Distribution & Illustrating Shoots at Kestane Beleni

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The majority of the gold is fine-grained (< 30 microns) occurring as free grains in quartz and rhodonite gangue, and within and in contact with pyrite, chalcopyrite and sphalerite, and to a lesser extent in galena (Figure 8-2). Higher gold grades, however, are not directly related to sulfide abundance. Silver content of the deposit is highly variable. The average silver grade for both the MOS and SOS is low (11.9 and 8.3 g/t), however, parts of the MOS shoot contain silver values in excess of 100 g/t.

Figure 8-2:Gold deportment at Efemçukuru

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The veins have a distinct paragenesis in addition to the mineralogical and metal zonation (Figure 8-3; Boucher, 2016). The earliest hydrothermal stage produced calc-silicate alteration, veins and breccia that locally cements cataclasite in the fault-vein system, and consists of a mineral assemblage of quartz-chlorite-epidote-actinolite. The main epithermal stage commenced with vein- and breccia-fill of very fine-grained quartz and rhodonite (pyroxmangite), rhodochrosite, and minor pyrite. This stage is associated with significant brecciation of calc-silicate altered phyllite wall rock, characterized by highly variable clast sizes reaching up to a few decimeters. The breccia clasts are angular and do not show significant comminution prior to hydrothermal vein emplacement. Comb-textured quartz and the Mn-rich phases commonly occur along the edges of the brecciated clasts growing towards the vein interiors. Adularia is extremely rare and has only been observed in the earlier phases of this vein stage where it appears in thin veinlets cutting flysch wall rock breccia clasts. Veins and veinlets are characterized by colloform banded quartz, rhodonite (pyroxmangite), axinite, and rhodochrosite, with minor calcite and rare pyrite and galena. Quartz textures vary from anhedral fine grained to chalcedonic crystals in individual bands or, in the later veins of this stage, as comb-textured vug infill. The Mn-phases typically grow together in pink bands superimposed on quartz-rich bands. Calcite and rhodochrosite can also be bladed in some vugs. This stage lacks abundant sulfide, although rare disseminated euhedral pyrite (< 0.5%) is observed in the later stages.

Figure 8-3:Mineral Paragensis at Efemçukuru (from Boucher, 2016)

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The main quartz-sulfide stage consists of veins of quartz and minor rhodonite (pyroxmangite), axinite, rhodochrosite, and calcite that are disseminated and banded with very fine-grained pyrite, galena, and sphalerite. Rhodonite and rhodochrosite are less common in this stage and quartz is the dominant gangue mineral. Base metal sulfides and pyrite are more abundant with subhedral to euhedral pyrite, sphalerite, galena, and rarely chalcopyrite occuring as fine-grained disseminations in quartz-rich portions of the veins, but also as very fine-grained, cockade textured bands surrounding brecciated clasts of the previous stages of veining. Colloform banding of quartz, carbonate, and base-metal sulfides, as well as carbonate and quartz vug infill are also common. The highest gold grades commonly occur with weakly disseminated (< 2%) pyrite, galena, and sphalerite.

The last vein stage includes thin veinlets of quartz, calcite, and rarely rhodochrosite. These 1-5 mm wide veinlets are observed in all drill core intercepts where they traverse the previous paragenetic stages. Quartz is anhedral, very fine-grained, and granular with interstitial calcite.

8.2

Deposit Model

The formation of the Efemçukuru gold deposit in western Anatolia coincided with Miocene extension, magmatism and hydrothermal activity including the formation of several other significant gold-rich porphyry and epithermal deposits in the region (Baker, 2019). In detail, Efemçukuru can be classified as an intermediate sulfidation epithermal vein deposit due to its high-base metal content and the Mn-rich nature of the veins, analogous to intermediate-sulfidation epithermal veins of Mexico (Simmons et al., 2005). The dominantly NE dipping Efemçukuru veins formed within faults that had east side down normal‐dextral (right lateral) shear sense.

Classic porphyry-epithermal models commonly show epithermal deposits forming on the periphery of higher temperature and deeper magmatic-hydrothermal systems such as porphyry and/or skarn deposits (Sillitoe and Hedenquist, 2003). The spatial and temporal distribution of rhyolite, high temperature calc-silicate alteration, and intermediate sulfidation epithermal veins support a magmatic-hydrothermal origin. Furthermore, detailed carbon and oxygen isotope analysis of vein carbonates indicate a mixed meteoric and magmatic source for the hydrothermal fluids and strongly support degassing and boiling of magmatic fluids in the generation of the main epithermal veins (Boucher, 2016).

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SECTION ● 9 Exploration

9.1

Surface and Subsurface Exploration Work

Prior to modern exploration and mining, Turkish Mine Exploration Institute records documented Efemçukuru as a manganese occurrence. However, it is believed that Romans mined the deposit in ancient times. Later, in the early 1900s, a British Company owned the exploration rights and carried a limited amount of work on the Efemçukuru deposit.

Modern exploration activity at Efemçukuru began in 1992, when Tüprag geologists recognized the exploration potential of the area while conducting reconnaissance work in western Turkey. Between 1992 and 1996, Tüprag conducted ground magnetic surveys, rock chip and soil sampling, geological mapping and 6,000 metres of diamond-drilling focusing primarily on the Kestane Beleni vein. This work identified high-grade gold mineralization in three separate zones: the South Ore Shoot (SOS), the Middle Ore Shoot (MOS) and the North Ore Shoot (NOS). Later exploration identified the Kestane Beleni northwest shoot (KBNW).

Infill drilling in 1997 and 1998 provided the basis for an initial resource estimate for the SOS and MOS, and a prefeasibility study was completed in 1999. Drilling programs from 2006 to 2010 significantly increased the mineral resources along strike to the north and downdip, and resulted in the first resource estimate for the NOS. Drilling in 2011 and 2012 focused on a new zone, KBNW, along strike from the NOS. Exploration since 2010 has focused on the Kokarpinar vein located east of, and oriented subparallel to, the Kestane Beleni vein. In 2018, the Bati veins were discovered in the footwall to the Kestane Beleni vein. In total over 175,000 km of drilling have been carried out at Efemçukuru (Table 9-1 and Figure 9-1).

Table 9-1:Exploration Activity Summary for Efemçukuru Mining Area

Exploration Activity	Quantity	Length (m)	Area (km²)
Stream Silt Sampling	147	-	-
Soil Sampling	4,530	205,500
Rock Chip Sampling	5,004	-	-
Trench Sampling	868	1,820
Mapping	-	-	24
Geophysics Magnetics	-	77,410
Geophysics IP	-	70,225	-
RC Drilling	66	5,421	-
Diamond Drilling	709	172,995	-

 

Figure 9-1: Exploration drilling at Efemçukuru

9.2

Mapping

Geological mapping at 1:5,000 scale was carried out over the Kestane Beleni vein in the early stages of exploration. To date, a total area of 24 km² has been mapped at Efemçukuru, including detailed geological mapping over the entire strike length of the Kestane Beleni, Kokarpinar and Bati vein systems. In addition mapping of the underground mine developments is ongoing. The surface and underground mapping, in addition to drill hole logging combined with structural data have been used to model the vein systems in 3D and help define new mineralized targets around the mine.

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9.3

Surface Sampling and Trenching

The total of 5,004 rock chip samples have been collected across the majority of outcropping veins on the property and 868 trench samples were obtained from 20 trenches. These data have helped prioritize drill targets but have not been used in the resource estimations.

9.4

Geophysics

Nearly 150 line km of ground geophysics have been conducted on the property to assist in identifying the epithermal vein systems (Table 9-1). This includes ground magnetic and IP and gradient IP surveys.

9.5

Mine Exploration

Mine exploration from underground has helped identify new mineralized targets and extensions at Efemçukuru. Targets include the down-dip extensions of the SOS and the MOS that have been assessed through underground exploration drilling from a hangingwall development named the Kestane Beleni HW drift. In 2017-2018, 9,440 m of diamond drilling in 28 holes was completed from the drift and extended gold mineralization down to 200 m RL. In 2019 mine exploration drilling has focused on the deeper parts of the Bati vein in the footwall to Kestane Beleni.

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SECTION ● 10 Drilling

10.1

Project Drilling

Five phases of drilling were carried out between 1992 and 1997 to gather geological, geochemical and metallurgical data.

Infill and further exploration drilling from 2006 to 2008 focussed on the north ore shoot (NOS). Following the positive feasibility study for the project, Eldorado continued drilling activity on Kokarpınar Vein and KBNW area between 2009 and 2011.

Drilling in 2011 and 2012 focused on a new zone along strike from the NOS, referred to as Kestane Beleni Northwest (KBNW), on down-dip extensions to the SOS and on the Kokarpinar vein, a parallel vein system located approximately 400 m east of the Kestane Beleni vein. Exploration drilling programs in 2013 through 2017 tested the Kokarpinar vein over a 3 km strike length, and identified resources in several discrete shoots.

In 2017 approximately 19,765 m of drilling was completed from the Kestane Beleni Drift, to test the extent of the mineralization especially at deeper parts of MOS and SOS-MOS transition area, and besides Kokarpinar vein. In 2018, 22,867 m of exploration drilling was completed targeting to expandthe resource in the Kokarpinar vein and Bati Veins located at the footwall of the Kestane Beleni vein which was discovered later in the year. 4,944 m were also drilled to delineate a mineralized zone at Kestane Beleni NW shoot.In 2019, at Bati and Kokarpinar Veins a total of 26,084 m of exploration and 5,827 m of delineation drilling on Kokarpinar vein was completed.

Table 10-1 summarizes the amount of drilling that has been completed on the property. Table 10-2 presents the drilling completed in last three years. A drill hole location map of the Efemçukuru mine area is shown on Figure 10-1.

Table 10-1: Summary of Drilling on the Efemçukuru Deposit

PRE-MINING
Location	Type of Drilling	Year	# of Holes	Meters
Kestane Beleni Vein	Core	1992-2011/2	299	51,052
	RC		58	5,028
Kokarpınar Vein	Core		16	4,340
	Percussion		8	393.8

POST-MINING
Location	Type of Drilling	Year	# of Holes	Meters
Kestane Beleni Vein	Core	2011/2-2018	194	48,603
Kokarpınar Vein			232	70,107
Batı Vein			87	34,781
Others			9	1,512
Total			903	215,816

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Table 10-2: Efemçukuru Deposit Drilling between 2017 and 2019

Kestane Beleni Deposit	2017		2018		2019		Total (m)
	Meters	# Holes	Meters	# Holes	Meters	# Holes
Geotechnic	1,220	7	-	-	-	-	1,220
Delineation	15,408	65	4,944	16	-	-	20,352
Infill	34,437	440	28,413	331	29,099	360	91,949

Kokarpinar Vein	2017		2018		2019		Total (m)
	Meters	#Holes	Meters	# Holes	Meters	# Holes
Exploration	4,357	21	6,285	20	7,886	54	18,527
Delineation	-	-	-	-	5,827	31	5,827

Bati Vein	2017		2018		2019		Total (m)
	Meters	#Holes	Meters	# Holes	Meters	# Holes
Exploration	-	-	16,582	39	18,198	48	34,781

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Figure 10-1: Drill Hole Location Map at Efemçukuru Mine Area

Various drilling contractors were used at different stages of the programs. In the early days, Kennebec and Major Drilling from Quebec and Turkish contractors like IDC, Ortadogu Drilling and Spektra Jeotech Drilling companies were contracted in later years.

The core drilling programs executed by using drill rigs of skid mounted Longyear 38, D-120, D-150, Atlas Copco CS-14 and Geo1500 rigs. Reverse circulation (RC) drilling was completed with Tüprag’s Explorer rig and an IDC Mustang rig.

Exploration and delineation drilling was carried out along the Kestane Beleni Vein on profiles spaced from 20 m to 40 m apart. The down dip spacing along profiles ranges from 20 m to over 40 m. The deeper exploratory holes were drilled from 435 exploration drift to intersect the vein over 500 m below surface vein exposure.

Most core holes, in the early drilling programs, were drilled approximately perpendicular to the vein at dips ranging from -45 to -85. The inclination and direction of drilling for the 2006 and 2007 all programs was variable due to the limited number of collar location available from which to drill. Holes drilled along 195°-260° azimuths at Kestane Beleni Vein and along 200°-300° at Kokarpınar Vein. Down hole deviation surveys for holes drilled in 2007, 2008 was measured using gyroscope (1991-1997), a Reflex EZ Shot (2006-2016) and Devishot (2017-2018) survey instruments with readings taken at 25 m intervals down the hole.

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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 were then entered into the project database. Specific gravity measurements were made on mineralized intervals as well as adjacent wall rock on each side. The core was photographed before being sampled.

Core recovery in the mineralized units was very good, averaging 97% for over 92% of core intervals in the mineralized zones. The relatively small number of poorer recovery intervals should have negligible impact on the Efemçukuru mineral resource estimate.

10.2

Infill Drilling

Diamond drilling is essential for infill and delineation to increase the geologic confidence in the mining areas as well as to improve the grade control model. At Efemçukuru the gold is not visible and is not evenly distributed along the mineralized vein. Added complications are the local structural patterns controlling the gold mineralization. As such, the definition of gold mineralization is dependent on diamond drilling. Over 35,000 m of mostly BQ -size core are drilled annually by three drill rigs at a cost of about US$35/m.

Figure 10-2 shows a diamond drill rig in an underground drill station.

Figure 10-2: Diamond Drill Rig in an Underground Drill Station

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The planning and execution of the drilling pattern depends on the configuration of the ore shoot and the planned mining method. Ore extraction is mainly by drift and fill mining augmented by ~15% long hole stoping. For areas mined by drift and fill, a 10 by 10 meter staggered grid pattern is implemented. (Figure 10-3a). Narrow vein long hole stoping was implemented over the last few years. This mining type is defined by modified 10 by 10 meter grid that contains a fifth hole located at the centre of the defined square (a “five-spot” or “dice” pattern; Figure 10-3b).The drilling of a specific area is scheduled to be finished 6 to 8 months prior to mining. This scheduling is sufficient for short-term grade model updates and any needed changes to the planned development or stope sequencing prior to planned production from the stope area.

Figure 10-3:Infill Drilling Pattern a) 10 m x 10 m grid b) five dice pattern

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SECTION ● 11 Sample Preparation, Analyses and Security

11.1

Sample Method

Most of the drilling since production commenced at Efemçukuru mine comprised infill and delineation drill holes (indicated to measured programs and inferred to indicated campaigns). Only sampling for the last three years (2017 to 2019) are to be described in this section as much of any earlier sampled material is already mined.

Drill core is placed into core boxes marked with hole ID, sequence numbering and depth interval. Sample intervals, selected and marked up by the logging geologist, ranged from 0.1 m to 3.7 m. The drill core samples were either cut with a diamond rock saw (if a delineation hole) or whole core sampled (if an infill hole) at the mine’s core logging facility at Gaziemir. The sample was bagged and were sent to the nearby ALS analytical laboratory in Izmir for sample preparation. The sample preparation procedure is as follows:

●

Samples are logged into the labs tracking system.

●

The samples are crushed to 90% passing 2 mm.

●

The samples are sub-sampled by riffle splitter until about one kilogram remains.

●

The sub-sample is pulverized to 90% passing 75 microns.

11.2

Assay Method

All samples are assayed for gold by 30 g fire assay with an AA finish and for multi-element determination using fusion digestion and inductively coupled plasma spectroscopy (ICP) analysis.

Samples that returned assays greater than 10 ppm were re-assayed by fire assay with a gravimetric finish.

11.3

Quality Assurance and Quality Control (QA/QC) Program

Assay results were provided to Eldorado in electronic format and as paper certificates. The procedure included inserting either a CRM, blank and duplicate into the sample stream every 8th samples. Upon receipt of assay results, values for Certified Reference Materials (CRMs) and field blanks were tabulated and compared to the established CRM pass-fail criteria:

●

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

●

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

●

Automatic batch failure if the field blank result is over 0.5 g/t Au, if the analytical blank is over 0.05 g/t.

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

Laboratory check assays were conducted at the rate of one per batch of 20 samples, using the same QA/QC criteria as routine assays. In addition, the ALS laboratory is regularly visited to observe and check that stated procedures are being used.

11.3.1

   QA/QC

Eldorado regularly monitors the performance of CRMs, blanks and duplicates as the assay results arrive on site. The last three years of CRM, blank and duplicate data were used for this report.

Table 11-1 shows the number of samples, blanks, duplicates and CRM’s used during drill campaigns for 2017, 2018 and 2019.

Table 11-1: QAQC sample frequency, 2017 to 2019

Number of Samples	2017 –2019
	Exploration	% of Total	Mine	% of Total
# Samples	9,439	85	34,803	90
#BLANKS	417	4	1,708	4
#DUPLICATES	512	5	497	1
#CRM	699	6	1,687	4
Total Assayed Samples	11,067		38,695

11.3.2

CRM Performance

Eldorado strictly monitors the performance of the CRM samples as the assay results arrive at site. Multiple CRM samples are inserted into the sample batches in order to cover a wide range of gold grades (0.5 g/t to 35 g/t). This corresponds to three grade groupings: ones close to the resource / reserve cut-off grade range, 2 to 4 g/t; ones that hover about the average gold grade of deposit, 7 to 10 g/t; and ones that control high-grade gold mineralization (greater than 10 g/t). CRM samples utilized over the last three years are shown in Table 11-2. CRM performance is monitored by charting. Examples of CRM charts are presented in Figure 11-1.

All samples were given a “fail” QAQC flag as a default entry in the mine’s database. A “pass” flag only was allowed once the assays were shown to have passed acceptance criteria. At the data cut-off, of December 2, 2019, all samples had passed QAQC criteria.

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Table 11-2: Main CRM Samples used between 2017 and 2019

	Au g/t	Standard	# Used	Period used
CRM CODE	Mean	Deviation	CRM Samples	From	To
COS055	2.638	0.087	224	Sep-12	Dec-19
COS058	3.231	0.096	565	Oct-12	Dec-19
COS085	4.014	0.106	342	Mar-16	Sep-18
COS086	6.692	0.140	266	Mar-16	Nov-17
COS087	11.655	0.334	361	Mar-16	Dec-19
GS11B	11.040	0.280	190	Jan-18	Nov-18
GS20A	21.110	0.460	227	Feb-18	Feb-19
GS20B	20.310	0.560	596	Apr-16	Feb-18
GS3L	3.180	0.090	306	Jul-14	Jul-17
GS3Q	3.300	0.120	97	Jun-16	Jul-19
GS8C	8.660	0.230	193	Sep-16	Nov-19
GS9A	9.290	0.320	670	Jun-13	Aug-17
GS25	25.600	0.510	129	Feb-19	May-19
GS3P	3.050	0.080	82	May-19	Jul-19
GS7G	7.180	0.210	186	Jul-19	Nov-19

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Figure 11-1: CRM Charts, 2017 to 2019, Efemçukuru Mine

11.3.3

 Blank Sample Performance

Two blank sample types were used: a field blank in the form of a rock sample and a pre-pulverized sample. The field blank samples were used to check for contamination during the entire sample preparation and analytical process at the laboratory. The analytical blank sample were used to check the contamination after the sample preparation process. The analytical detection limit for gold was 0.01 g/t. The rejection threshold of field blank was chosen to equal 0.5 g/t, the chosen threshold for pulp blank was 0.05 g/t. The results are shown in Figure 11-2 and show no evidence of contamination.

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Figure 11-2: Efemçukuru Blank Data – 2017-2019 Drill Programs

11.3.4

Duplicates Performance

Eldorado implemented a program, which monitored data from regularly submitted coarse reject duplicates. The duplicate data for Efemçukuru reproduce well and are shown in a relative difference charts in Figure 11-3 and percentile rank chart in Figure 11-4. Patterns observed in the relative difference plot are symmetric about zero, suggesting no bias in the assay process. The coarse reject chart shows that almost all data greater than 1 g/t fall well within the 20% limits. Of note is the excellent replication of samples with values greater than 10 g/t. This is also shown in the percentile rank plot where, at the 90th percentile of the population, the Efemçukuru data shows 10% difference in the coarse reject data.

Figure 11-3:Relative Difference Chart, Efemçukuru 2017-2019 Duplicate Data

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Figure 11-4: Percentile Rank Plot, Efemçukuru 2017-2019 Duplicate Data

Duplicate data were plotted on a Quantile-Quantile (QQ) plot to test for any bias in the analyses. If the distribution lies on or oscillates tightly about the 1:1 line, then the sample population is unbiased. This is the pattern observed for the Efemçukuru gold data in Figure 11-5

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Figure 11-5: QQ Plot for Efemçukuru Duplicate Data

11.3.5

Specific Gravity Program

Samples taken for assay from core holes were being measured for specific gravity and tabulated by rock type. The specific gravity for non-porous samples (the most common type) was calculated using the weights of representative samples in water (W2) and in air (W1). The bulk density was calculated by:

W1

(W1 – W2)

Less-common porous samples were dried and then coated with paraffin before weighing. An allowance was made for the weight and volume of the paraffin when calculating the specific gravity.

11.4

Concluding Statement

In Eldorado’s opinion, the QA/QC results demonstrate that the Efemçukuru mine’s assay database, in particular for data since 2017, is sufficiently accurate and precise for the resource estimation.

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SECTION ● 12 Data Verification

12.1

Introduction

All collected and assayed data were imported into Efemçukuru mine’s database. An in-house interface was used for data entry. This interface uses numerous data verification and cross-checking protocols to ensure the quality of the data being entered. A “quarantine” system is in place to hold information until validated and checked. Release can only be made by the database administrator. Assay data are imported directly from the laboratory’s csv file. Drilling data such as collar surveys, down hole surveys, assays and logged lithologies are routinely cross checked against original source files. Discrepancies are rare but when observed, immediately corrected. As a result of these checks and balances, the data supporting the Efemçukuru resource work are sufficiently free of error and adequate for resource estimation.

12.2

Block Model to MillReconciliation

Another form of data verification is the reconciliation to production of mined portions of the resource model. The reconciliation process flowchart that occurs at Efemçukuru mine is shown in Figure 12-1.

Figure 12-1: Flowchart of the Reconciliation Procedure at Efemçukuru Mine

The annual reconciliation of resource model to mill feed tonnage and grade are shown since 2012 in Table 12-1. The annual reconciliation improved markedly in 2016 to within 1-2 percentage points and has maintained this excellent performance through 2019. These results clearly demonstrate that the Efemçukuru data management and QAQC protocols produce highly verifiable data that form the basis of quality resource estimation at Efemçukuru mine.

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Table 12-1: Annual Resource Model to Mill Reconciliation data

	Resource Model			Mill Feed			Differences of Resource Model (%)
Year	Tones	Au g/t	Au oz	Tones	Au g/t	Au oz	Tones	Grade	Gold
2012	369,953	9.86	117,325	352,156	9.26	104,791	-4.8%	-6.2%	-10.7%
2013	399,930	9.32	119,833	413,513	8.87	117,895	3.4%	-4.8%	-1.6%
2014	428,276	8.38	115,334	436,851	8.34	117,099	2.0%	-0.5%	1.5%
2015	454,238	7.28	106,384	454,864	7.82	114,329	0.1%	7.3%	7.5%
2016	471,178	7.36	111,508	476,529	7.40	113,398	1.1%	0.6%	1.7%
2017	471,680	6.94	105,238	481,648	7.01	108,594	2.1%	1.1%	3.2%
2018	498,366	6.86	109,903	499,120	6.77	108,620	0.2%	-1.3%	-1.2%
2019	511,175	7.12	117,018	521,033	7.03	117,818	1.9%	-1.2%	0.7%
Total	3,604,796	7.79	902,544	3,635,715	7.72	902,544	0.9%	-0.9%	0.0%

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SECTION ● 13 Mineral Processing and Metallurgical Testwork

13.1

Introduction

The Efemçukuru concentrator has been successfully processing ore from Efemçukuru mine since commissioning in mid-2011. The original process flowsheet was designed to produce gold doré from the gravity concentrate and from the gold-containing sulfide minerals flotation concentrate. However, production of gravity concentrate has been considerably lower than what was estimated in the feasibility study (expected 30 % recovery) and thus, the process flowsheet was revised in 2018 to produce only flotation concentrate for sale. This had no effect on overall recovery because gravity concentrate was recovered from the flotation concentrate.

Mineral processing and metallurgical tests of ore samples taken during production period are discussed in this section. Table 13-1 shows the reports reviewed in order to derive the updated process design criteria, describe the process plant and production figures.

Table 13-1: Reports Reviewed

Report Title	Author	Date
Technical Report on the Efemçukuru Project, NI 43-101 Report	Wardrop	August 2007
Tüprag Madencilik Efemçukuru Gold Mine, Turkey Metallurgical Testwork	Wardell Armstrong International	March 2016
Efemçukuru Gold Ore	Petrolab, Mineralogical Report AM2767	February 2018
Performance Evaluation and Optimization of Efemçukuru Concentrator	Hacettepe Mineral Technologies (HMT)	August 2018.
Pilot Column Flotation Cell Tests and Selection of Column Flotation Cells for Efemçukuru Flotation Plant	Hacettepe Mineral Technologies (HMT)	January 2019
Plant Archive: Plant data recorded from 2011 to end of 2018	TÜPRAG	2019

13.2

Efemçukuru Gold Ore

Metallurgical tests have been performed using variability ore samples (8 samples) and composite samples representing south ore shoot (SOS) and middle ore shoot (MOS) during feasibility study. Process flowsheet was designed based on this testwork program. Another metallurgical testwork program was undertaken by Wardell Armstrong International (WAI) in 2016 to update the information about metallurgical characteristics and flotation behaviour of main ore types from SOS, north ore shoot (NOS) and MOS.

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13.2.1

Head Assay

Efemçukuru is an underground mine and the ore is currently mined from three ore shoots, namely, SOS, MOS and NOS. The run-of-mine (ROM) ores are crushed in a jaw crusher underground and stored in bins on the surface. The plant feed is prepared as a blend of these ores, and their distribution in the blend is determined according to head assays. Typical head assays of these ore samples are given in Table 13-2. Head assay of plant feed sample taken during plant survey in February 2018 by HMT is also given in the table for comparison. This sample represents a snapshot of the feed during three-hour plant survey.

During original design stage, expected mill feed grades were around 10.0 g/t gold and 17.8 g/t silver. Head grades of typical high-grade ores from SOS, NOS, MOS and one plant feed in February 2018 are presented in Table 13-2. These samples are variability materials representing high-grade part of the ore body.

Table 13-2: Head Assay of typical High-Grade Ores

Element	SOS*	NOS*	MOS1*	MOS2*	Plant Feed**
Au (g/t)	17.5	23.5	7.0	15.2	5.1
Ag (g/t)	28.6	36.4	13.4	17.2	18.9
Cu (%)	0.09	0.14	0.03	0.04	0.04
Pb (%)	0.21	1.32	0.25	0.31	0.64
Zn (%)	0.36	1.48	0.28	0.77	0.39
Fe (%)	3.53	7.89	2.26	5.25	4.63
STOTAL (%)	2.17	6.65	1.13	4.88	2.77
S2- (%)	1.46	0.07	1.06	4.83	-
CTOTAL (%)	1.55	1.38	2.41	-	-

*WAI Report; **HMT report

13.2.2

Mineralogical Examination

Typical mineralogical analysis of the plant feed is given in Table 13-3. Major sulfide minerals identified in all the samples are pyrite, sphalerite and galena. Small percentage of chalcopyrite is also present. Major gangue minerals are quartz, manganese-minerals with minor feldspar, calcium-carbonates, siderite, mica-group minerals and epidote group minerals. There are trace amounts of several other phases.

Gold is generally fine-grained and primarily associated with pyrite and galena, some association with manganese-carbonates and iron oxides.

Pyrite and sphalerite generally show good liberation, with improving liberation for the fine size fractions. Galena also shows improving liberation for the fine size fractions.

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Table 13-3:Modal Mineralogy of One Plant Feed Sample in February 2018

Minerals	%
Manganese-Silicate	9.89
Manganese-Carbonate	9.85
Sphalerite	1.60
Galena	0.47
Chalcopyrite	0.03
Quartz	38.29
Pyrite	11.18
Feldspar	7.40
Calcium-carbonates	4.84
Siderite	4.33
Mica	3.97
Epidote	3.49
Ultramafics	1.79
Iron Oxides	1.62
Fluorite	0.93
Others	0.30

13.2.3

Comminution Characteristics

Comminution characteristics of the ore have been measured by various laboratories at different periods. Original design values and the results of latest tests performed by Wardell Armstrong International (WAI) are given in Table 13-4. All the tests reflect that the ore is relatively hard, abrasive and resistant to impact breakage.

Table 13-4: Comminution Characteristics of Efemçukuru Ore

Sample Name	BWi (kW.h/t)	Ai	Axb	ta	SCSE* (kW.h/t)	S.G. (t/m3)
Plant Feed in February 2018	20.4	-	43.4	0.35	9.9	2.89
SOS Ore	18.1	0.567	43.1	0.53	10.1	2.95
NOS Ore	19.4	0.636	34.4	0.21	11.6	3.11
Original Design Values	20.7	0.713	46.4	0.51	-	2.90

*SAG circuit specific energy

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13.3

Flotation

Flotation testwork was performed on four samples (SOS, NOS, MOS1, MOS2) by WAI. Optimum primary grind size was determined to be 80% passing 63 µm. The reagents used in flotation were; sodium bisulfite (NaHO3) as sulfidizing agent, copper sulfate (CuSO4.5H2O) as activator, xanthate (SIBX) as collector, S-8045 as promoter and OrePrep F-549 as frother.

After rougher and open-circuit cleaner tests, locked-cycle tests (LCT) were performed for each sample under optimum flotation conditions. Table 13-5 shows concentrate mass pull, concentrate grade and recovery values of final concentrates from LCT tests. Gold grade of the concentrates ranged between 106 g/t and 220 g/t, and gold recovery between 86.5% and 94.5%. Despite the NOS containing 6.58% sulfide sulfur content, the final concentrate contained 119 g/t Au with 93.6% gold recovery. For SOS ore sample, gold recovery (86.5%) is lower, despite total sulfur recovery being almost 95%. Mineralogical analysis has identified gold as being predominantly associated with pyrite, and in most cases, gold recovery is proportional to sulfur recovery. The situation with SOS ore samples is perhaps indicative of some minor gold association with non-sulfide gangue, particularly manganese-minerals.

Table 13-5: Concentrate Mass Pull, Grade and Recovery of Final Concentrates from Locked Cycle Tests

			Grade			Recovery (%)
Sample		Weight (%)	Au (g/t)	Ag (g/t)	ST (%)	Au	Ag	ST
SOS	Concentrate	6.3	220	305	34.3	86.5	76.1	94.7
	Feed		16.1	25.3	2.29
NOS	Concentrate	16.3	119	179	38.8	93.6	95.1	98.2
	Feed		20.6	30.6	6.42
MOS1	Concentrate	4.4	185	246	24.7	94.5	89.9	97.3
	Feed		8.6	12.0	1.11
MOS2	Concentrate	14.0	106	106	33.9	93.7	93.2	96.4
	Feed		15.8	15.9	4.90

13.4

Gravity Concentration

A series of gravity concentration tests were performed by HMT in 2018 using pulp samples taken from SAG Mill discharge and ball mill discharge during a plant survey in February 2018. The pulp samples were screened through a 500-µm aperture screen and the undersize material was classified further in a hydrocyclone to remove the -38 µm size fraction as slime material. The cyclone underflow was again classified in a teetered bed separator (TBS) into two size/density streams. Both streams (underflow and overflow streams) were concentrated separately by using a shaking table to produce coarse and fine gravity concentrates.

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The results show that a gravity concentrate (combined coarse and fine concentrates) from SAG Mill discharge contained 67 g/t Au with 33% recovery, and from Ball Mill discharge contained 107 g/t Au with 65% (stage) recovery. These gravity concentrates are not suitable for direct smelting, but very compatible with the bulk sulfide flotation concentrate. This is mainly due to fine grain size of gold particles associated with sulfide minerals, such as pyrite. Therefore, the gravity concentration circuit was decommissioned in 2017.

13.5

Performance Analysis of the Flotation Circuit

A comprehensive plant survey was performed in the comminution and flotation circuits of Efemçukuru flotation plant in February 2018. Cell by cell profile sampling and hydrodynamic measurements (gas hold-up, superficial air velocity and froth depth) were conducted in the flotation circuit. The measurements show that flotation performance of all of the flotation cells in the rougher and cleaner sections are generally in an expected operating range.

Mass balance of the flotation circuit was performed on a size-by-size assay basis. Mineralogical analyses, including gold deportment, of feed, tail and concentrates were also performed. Mass balance calculations show that the flotation circuit produces a concentrate assaying 50 g/t Au, 160 g/t Ag and 28% S at recoveries of 95%, 96% and 97%, respectively. The sulfur grade is quite low for a bulk sulfide concentrate due to presence of significant amount of non-sulfide gangue minerals at -25 µm particle size. The majority of these particles are recovered by entrainment mechanism.

Batch scale diagnostic flotation tests conducted during the plant survey show that gold grade of the concentrate can be increased from approximately 50 g/t to 80 g/t by further cleaner flotation at high froth depth. It must be noted that rejection of non-sulfide gangue particles increases not only grades of gold and silver, but also lead and zinc which are the main penalty elements in the concentrate. The plant aims to keep the combined lead and zinc grade below 20%.

13.5.1

Pilot Scale Column Flotation Tests

A pilot scale column flotation unit was installed in the flotation plant in 2018 for upgrading the final concentrate by removal of entrained gangue mineral particles. The pilot unit has dimensions of 508 mm diameter x 4,000 mm length and two sparger systems; SparJet and cavitation tube. The pilot tests were performed using the final concentrate and rougher cell 1-2 concentrate. Effects of various operational parameters (froth depth, froth wash water rate, air flow rate, sparger system, feed flow rate i.e. residence time, frother addition) were investigated to determine the optimum operating conditions.

The results show that upgrading the final concentrate by using the column flotation cell as the third cleaner flotation stage is the most beneficial option. The optimum operating conditions for the column cell were determined as; froth depth: 30 cm, wash water: 12 L/min, air flow rate: 8 m3/h, cavitation tube sparger system at 70% pump speed, residence time: 30 minutes. Continuous column flotation tests were conducted for 33 days (on day shift, two sampling campaigns per shift) applying the optimum operating parameters to validate the findings from optimization studies.

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Mass balance of each sampling campaign were conducted by using JKSimFloat software to determine flowrates and metal recoveries. The feed grade to the column cells (i.e. the final concentrate of the flotation circuit) ranges from 35 g/t to 120 g/t Au, and from 13% to 35% S. Recoveries of both gold and sulfur are over 95% independent of feed grade. Sulfur grade of the column concentrate is generally over 35% S, independent of the feed grade. Gold grade of the concentrate, however, depends on feed grade and increases linearly with gold grade of the feed to the column flotation cell. Gold grade of the column concentrate ranges from about 60 g/t to 150 g/t depending on feed grade. Table 13-6 shows average mass pull, grade and recoveries obtained from the continuous column tests. The concentrate can be produced assaying approximately 98 g/t Au, 37% S with 98% recovery for both elements. Approximately 30% of the mass is rejected.

Table 13-6: Average Mass Pull, Grade and Recoveries of the Continuous Column Flotation Tests

		Grade					Recovery
	Mass Pull (%)	Au  g/t	Ag  g/t	S %	Pb %	Zn %	Au %	Ag %	S %	Pb %	Zn %
Concentrate	68.6	97.6	212	37.0	7.94	8.48	98.0	97.5	98.2	97.5	97.6
Feed	100	66.4	148	25.5	5.54	5.93	100	100	100	100	100

13.6

Recommendations for Future Testwork

13.6.1

Variability Tests

Tüprag has continued to complete variability testing as new areas have been developed. This includes:

●

Complete chemical analysis

●

Mineralogical analysis including gold deportment

●

Comminution tests (SMC, Bond ball mill work index, Abrasion Index)

●

Cleaner flotation tests

With minor variations this variability work has continued to indicate that the designed process is suitable for the treatment of these ores. As we continue to develop and access the NOS and eventually Kestane Beleni North-West, base metals increase in grade and may warrant increased attention to blending or a plant adjustment at some point in the future.

13.6.2

Improving Concentrate Quality

Final concentrate contains significant amount of non-sulfide gangue minerals, which are mostly in liberated form and recovered by entrainment. It is therefore expected that column flotation cells will improve the grade of the concentrate by controlling froth drainage.

It must be noted that grades of lead and zinc can be higher after rejection of non-sulfide gangue minerals. Combined Pb and Zn grade higher than 20% could cause marketing issues. Mineralogical analysis of various ore samples and final concentrates have shown that galena and sphalerite are the source of base metals in the concentrate. Further testwork is recommended to investigate the opportunities of reducing Pb and Zn content in the final concentrate.

Recently, some buyers have demonstrated an interest in material containing high Pb/Zn and are prepared to pay for the contained lead. This market will be more fully developed as we complete the installation of the column flotation cells.

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SECTION ● 14 Mineral Resource Estimates

14.1

Introduction

The mineral resource estimates for Efemçukuru consist of 3D block models formed on the Kestane Beleni, Kokarpınar and Batı epithermal vein systems. Creation of these models utilized a commercial mine planning software package (Geovia Gems). Currently, mining only occurs within the Kestane Beleni vein system. The project limits and block model properties are presented in Table 14-1.

 Table 14-1: Project Limits and Block Model Properties

Number of Blocks (#)
Columns	820
Rows	372
Levels	144
Origin
X	496,636
Y	4,237,916
Z	765
Block Size (m)
Column Size	4
Row Size	4
Level Size	5

14.2

Mineralization Models

Gold mineralization at Efemçukuru primarily occurs in the principal veins. Within these veins, the gold distribution can be irregularly distributed, either located along the footwall or hanging wall vein margins, within the central portions or combinations of all three. This distribution can only be confirmed through assays. Domains to control grade interpolation are, by necessity, grade based.

The modeling domains used a 2.0 g/t Au grade threshold and general vein geometry for their construction. The domains also honored a minimum 2.0 m rule for mineralization thickness. The generated shapes were checked in plan and section to ensure conformance to assay data and the vein models.

The domains were further divided into zones or shoots according to structural and spatial considerations. (Figure 14-1). The shoots of the Kestane Beleni vein system are South Ore Shoot (SOS), Middle Ore Shoot (MOS), North Ore Shoot (NOS) and Kestane Beleni Northwest shoot (KBNW). A stockwork mineralization zone occurs in the MOS and comprises a separate modeling domain. The Kokarpinar vein system consists simply of a South domain (KPS) and Middle domain (KPM).

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These domain solids coded the drill hole data and block model cells ahead of grade interpolation.

Figure 14-1: Modeling domains for the Kestane Beleni, Kokarpinar and Bati vein systems

14.2.1

Historical Mine Openings or Voids

Evidence of historical mining, by the presence of numerous small underground openings and small surface rock piles, occurs predominantly in the upper parts of the SOS (Figure 14-2). These openings or voids, first identified by the diamond drill campaigns, became exposed by mining. This allowed better location definition through surveying tools such as cavity-monitoring systems. The block model captured the void volumes to ensure proper accounting in tonnage estimates and for operational safety purposes. The volume of known voids was 36,074 m3.

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Figure 14-2: Modeled voids in the Kestane Beleni Vein System

14.3

Data Analysis

The Efemçukuru resource database, as of December 01, 2019, comprised 33, 018 samples in 4,103 drill holes Most of the sample lengths are between 0.5 and 1.5 meters, and average 0.92 m. This supported using a down hole composite length of 1 m.

The composited data were interrogated, by domain using descriptive statistics, histograms, box plots, and cumulative distribution function (CDF) plots. Summaries of the statistical properties for uncapped data occur in Table 14-2.

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Table 14-2: Efemçukuru Statistics for 1 m Uncapped Composite Au Data (g/t)

	Kestane Beleni
	SOS	MOS	NOS	KBNW	Kokarpınar	Batı
	Au	Au	Au	Au	Au	Au
# Samples	15,041	12,349	4,641	268	682	148
Min	0.01	0.007	0.009	0.007	0.002	0.007
Max	570	309	519	174	705	260
Mean	8.13	7.1	9.99	7.61	9.33	10.5
SD	16.1	13.81	19.06	18.9	39.89	29.41
Variance	259	191	363	357	1,591	865
CV	1.98	1.95	1.91	2.48	4.28	2.8
Q25	2.07	0.54	2.81	0.69	0.9	1.59
Median	4.07	2.42	5.39	2.55	2.5	3.66
Q75	8.91	7.62	10.76	6.3	5.73	6.7
Kurtosis	301.3	80.4	188.2	46.8	162.9	42.6
Skewness	12.9	6.5	10.5	6.2	11.4	6.1

14.3.1

Evaluation of Extreme Grades

Extreme grades were examined for gold, mainly by histogram and CDF plots. The examination showed a risk does exist with respect to extreme gold grades at Efemçukuru. Assay gold grades were capped to 100 g/t (SOS and NOS), 200 g/t (MOS) and 40 g/t (KBNW, Kokarpınar and Bati) prior to compositing. The capped composite data statistics are shown in Table 14-3.

Table 14-3: Efemçukuru Statistics for 1 m Capped Composite Au Data (g/t)

	Kestane Beleni
	SOS	MOS	NOS	KBNW	Kokarpınar	Batı
	Au Cap	Au Cap	Au Cap	Au Cap	Au Cap	Au Cap
# Samples	15,041	12,349	4,641	268	682	148
Min	0.01	0.007	0.009	0.007	0.002	0.007
Max	100	200	100	40	40	40
Mean	7.82	7.05	9.49	5.93	5.37	6.71
SD	11.67	13.21	13.54	9.13	8.5	9.51
Variance	136	175	183	83	72	91
CV	1.49	1.87	1.43	1.54	1.58	1.42
Q25	2.07	0.54	2.81	0.69	0.9	1.59
Median	4.07	2.42	5.39	2.55	2.5	3.66
Q75	8.91	7.62	10.76	6.3	5.73	6.7
Capped Au g/t	100	200	100	40	40	40
# Capped Samples	59	4	38	9	22	7
Kurtosis	23.6	28.6	28.5	76.8	8.8	2.9
Skewness	4.2	4.1	4.2	7.3	3	1.9

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The coefficient of variation (CV), defined as the ratio of the standard deviation to the mean, is a useful metric for comparing the effect of capped grades on the distributions. Average CV values for capped and uncapped composite data in the main modeling domains are plotted in Figure 14-3. The assay cap grade strategy succeeds in moderating the effects of extreme grades in each of the modeling domains, especially in the Kokarpınar vein system.

Figure 14-3: Average Coefficient of Variation for Capped and Uncapped Data by Domain

14.3.2

Histograms and Cumulative Distribution Function (CDF) Plots

Cumulative frequency or cumulative distribution function (CDF) diagrams demonstrate the relationship between the cumulative frequency (expressed as a percentile or probability) and grade on a logarithmic scale. These are useful for characterizing grade distributions and identifying multiple populations within a data set.

The analyses show asymmetrical positive skewed trends, typical patterns for epithermal precious metal systems. Additionally, multi-modal populations are present, again a common observation for these mineralization types whose veining clearly show episodic pulses of emplacement. CDF plots, here as cumulative probability graphs, for the main zones are shown in Figure 14-4 and Figure 14-5.

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Figure 14-4: Cumulative Probability Plots, Kestane Beleni Vein System

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Figure 14-5: Cumulative Probability Plots, Kokarpinar and Bati Systems

14.3.3

Estimation Domains

The data analysis and geologic interpretation of the gold mineralization at Efemçukuru supports the use of a grade-based shell to define the mineralized portions of the vein systems. As described above, grade shells were constructed in the veins using a threshold grade of about 2 g/t Au. There are several small splays at the hanging wall of main vein. The spatial relationships between the main vein and small splay veins do not necessitate the use of separate domains for grade interpolation. The MOS stockwork zone with its marked grade contrast and mineralization style, was treated as a separate estimation domain. Grades in these domains are estimated with a hard boundary logic, that is, only composites within a domain would be used to interpolate grade into blocks defined by that domain.

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14.4

Variography

Variography, a part of data analysis, is the study of the spatial variability of an attribute. Correlograms, rather than the traditional variograms, were used on the Efemçukuru data because of their lower sensitivity to outliers and their normalization to the variance of the data for a given lag.

Correlograms were calculated for gold in the in the main domains (SOS, MOS, NOS). Correlogram model parameters and orientation data of rotated correlogram axes are shown in Table 14-4.

Table 14-4: Correlogram Parameters for the main Kestane Beleni domains

	SOS	MOS	NOS
Nugget (Co)	0.12	0.18	0.204
First Structure (C1) Model	SPH	SPH	SPH
Proportion of C1	0.707	0.561	0.534
Range of anisotropy X	2.5	4.7	3.9
Range of anisotropy Y	2.1	2.2	1.7
Range of anisotropy Z	5.7	3.1	6.5
Rotation Z	19	-2	-51
Rotation X	-34	-17	34
Rotation Z	29	-40	36
Second Structure (C2) Model	SPH	SPH	SPH
Proportion ofC2	0.172	0.259	0.262
Range of anisotropy X	8.2	38.6	35.3
Range of anisotropy Y	48.8	22.6	7.2
Range of anisotropy Z	74.9	61.2	77.6
Rotation Z	-8	67	-51
Rotation X	54	-24	34
Rotation Z	32	-95	36

 Notes: Models are Spherical. The Anisotropy Rotation is ZXZ RRR hand rule. Positive rotation around the X axis is from Y towards Z, around Y axis is from Z towards X, and around the Z axis is from X toward Y.

Overall observations from the generated correlograms are:

●

For the Kestane Beleni domains that experienced more closely spaced, infill drilling campaigns (conversion of Indicated to Measured resources) the models show distinct small- ranged first structures and a smaller nugget.

●

Gold in the SOS domain displays dominantly NW to SE trending, moderate NE-dipping and SE-plunging structures. The SOS gold nugget effect is 12% of the total variation.

●

MOS domain distribution displays a NW to SE-trending, moderately steep NE-plunging structure. The nugget effect for the MOS distribution is somewhat higher at 18% of the total variation.

●

The gold distribution in NOS displays a WNW to ESE-trending, moderately steep NE- plunging structure. The nugget effect for NOS distribution, the highest of the three main domains, comprises 20.4% of the total variation.

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14.5

Model Set-up

The model was set-up using Geovia GEMS software. The block size for the Efemçukuru model was selected based on mining selectivity considerations (underground mining). The block size mirrors the minimum mining unit for Drift and Fill mining method: 4 m east x 4 m west x 5 m high.

  The capping limits were applied to the assay data prior to compositing. The assays were composited into 1 m fixed length down-hole composites. The compositing honoured the estimation domain by breaking the composites on the domain code values. The compositing process was reviewed and found to have performed as expected.

Various coding was done on the block model in preparation for grade interpolation. The block model was coded according to domains and sub-domains for usage of different search ellipsoids. Once mining in an area is deemed to be complete, construction of 3D depletion shapes allows proper remaining ore percentage to be calculated ahead of remaining tonnage tabulation.

14.6

Estimation

Efemçukuru estimation plans, or sets of parameters used for estimating blocks, were designed using a philosophy of restricting the number of samples for local estimation. This was found to be an effective method of reducing smoothing and producing estimates that match the Discrete Gaussian change-of-support model and ultimately the actual recovered grade-tonnage distributions. While local predictions based on the small number of samples were uncertain, this method produced reliable estimates of the recovered tonnage and grade over the entire deposit. The global grade-tonnage curves from the estimations are accurate predictors of the actual grade-tonnage curves.

Modelling consisted of grade interpolation by ordinary kriging (KG) for SOS, MOS and NOS domains and inverse distance weighting to the second power (ID) in the remainder of the zones due to their insufficient data to create correlograms. Nearest-neighbour (NN) grades were also interpolated for validation purposes. No grades were interpolated outside the modeling domains.

The search ellipsoids were oriented preferentially to the orientation of the vein in the respective domains. 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. SOS, MOS and NOS model blocks received a minimum of 3 to 5 and maximum of 2 to 4 composites from a single drill hole (for the two-hole minimum pass). Maximum composite limit equaled 16. The remaining domains received a minimum of 3 and maximum of 2 composites from a single drill hole whereas the total maximum composite limit ranged from 6 to 8. The minimum and maximum number of composites were adjusted to incorporate an appropriate amount of grade smoothing (see 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 Kokarpinar and Bati vein systems. The restricted distance generally ranged from 30 to 40 m meaning that beyond this distance from a model block center, composites exceeding the outlier values are not used in estimation. The threshold grades generally ranged from 25 to 40 g/t Au.

Bulk density values were estimated into the Kestane Beleni blocks by inverse distance weighting to the fourth power using a minimum of 2 and a maximum of 8 composites. Averaged measured density values, equaling 2.8, were simply assigned to Kokarpinar and Bati model blocks.

14.6.1

Validation

14.6.1.1

Visual Inspection

Eldorado completed a detailed visual validation of the resource model. Models were checked for proper coding of drill hole intervals and block model cells, in both section and plan. Coding was found to be properly done. Grade interpolation was examined relative to drill hole composite values by inspecting sections and plans. The checks showed good agreement between drill hole composite values and model cell values. The hard boundaries between grade shells 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. Examples of representative sections containing block model grades and point view drill hole composites are shown on Figure 14-6 to Figure 14-11.

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Figure 14-6: SOS showing modeled Gold Grades and composited Drill Hole Samples (g/t)

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Figure 14-7: MOS showing modeled Gold Grades and composited Drill Hole Samples (g/t)

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Figure 14-8: NOS showing modeled Gold Grades and composited Drill Hole Samples (g/t)

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Figure 14-9: KBNW showing modeled Gold Grades and composited Drill Hole Samples (g/t)

Figure 14-10: Kokarpinar Domains (KPS on left; KPM on right) showing modeled Gold Grades and composited Drill Hole Damples (g/t)

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Figure 14-11: Bati veins (#1 on left; #2 on right) showing modeled Gold Grades and composited Drill Hole Samples (g/t)

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14.6.1.2

Model Checks for Bias

Eldorado checked the block model estimates for global bias by comparing the average metal grades (with no cut-off) from the OK and ID models with means from nearest-neighbour estimates. The nearest-neighbour estimator declusters the data and produces a theoretically unbiased estimate of the average value when no cut-off grade is imposed. It is a good basis for checking the performance of different estimation methods. Results summarized in Table 14-5 show no global bias in the estimates.

Table 14-5: Global Model Mean Grade Gold Values (g/t) by Domain

	ZONE	OK or ID Estimate	NN Estimate	Difference
Kestane Beleni	SOS	5.81	5.66	-2.51%
	MOS	5.36	5.35	-0.22%
	NOS	5.96	6.00	0.67%
	KBNW	4.62	4.59	-0.62%
Kokarpinar	KPS	2.63	2.76	4.73%
	KPM	5.91	6.04	2.10%
	KPMSP1	3.36	3.39	0.86%
	KPMSP4	4.72	4.66	-1.33%
BATI	BATI 1	5.88	6.09	3.51%
	BATI 2	8.14	8.37	2.81%

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Eldorado also checked for local trends in the grade estimates (grade slice or swath checks). This was done by plotting the mean values from the nearest-neighbour estimate versus the OK or ID modeled results for elevation benches, eastings and northings in 10 m swaths. Examples of the three major domains (SOS, MOS and NOS) are shown in Figure 14-12 to Figure 14-14. The model estimate should be smoother than the nearest-neighbour estimate, thus the nearest-neighbour estimate should fluctuate around the model estimate on the plots. The observed trends behave as predicted and show no significant trends in the estimates in models.

Figure 14-12: Model trend plot showing 10m binned averages along elevations for kriged (OK) and nearest neighbour (NN) gold grade estimates, SOS domain

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Figure 14-13: Model trend plot showing 10m binned averages along elevations for kriged (OK) and nearest neighbour (NN) gold grade estimates, MOS domain

Figure 14-14: Model trend plot showing 10m binned averages along elevations for kriged (OK) and nearest neighbour (NN) gold grade estimates, NOS domain

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14.6.1.3

Model Check for Change-of-Support

An independent check on the smoothing in the estimates was made using the Discrete Gaussian or Hermitian polynominal 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:

●

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 4 m x 4 m x 5 m block grades for gold in the SOS, MOS and NOS are shown in Figure 14-15 to Figure 14-17. These 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 likely range of mining grade cut-off values (around 3.0 g/t Au).

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Figure 14-15: HERCO plot, SOS domain

Figure 14-16: HERCO plot, MOS domain

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Figure 14-17: HERCO plot, NOS domain

14.7

 Mineral Resource Classification and Summary

The mineral resources of the Efemçukuru Project were classified using logic consistent with the CIM definitions referred to in the National Instrument 43-101-Standards of Disclosure for Mineral Projects (NI 43-101). The mineralization of the project satisfies sufficient criteria to be classified into measured, indicated, and inferred mineral resource categories.

Inspection of the model and drill hole data on plans and sections combined with spatial statistical work showed good geologic and grade continuity in areas where near active mining and were tested by underground infill drilling that resulted in an approximate sample spacing of about 10 m. When taken together with all observed factors, blocks covered by this data spacing at SOS, MOS and NOS shoots may be classified as measured mineral resource. To assist in selecting eligible model blocks, a three-hole rule was used, where blocks containing an estimate resulting from four or more samples from different holes were tagged. These blocks were examined in longitudinal section where polygons were digitally drawn around contiguous areas of appropriately tagged blocks. These shapes were subsequently used to classify blocks as measured mineral resources.

The indicated mineral resource category is supported by the present drilling grid over most of the remaining part of the deposits. The drill spacing is at a nominal 45 m on and between sections. Geologic and grade continuity is demonstrated by the inspection of the model and drill hole data in plans and sections over various zones, combined with spatial statistical work. Considering these factors, blocks covered by this data spacing may be classified as indicated mineral resource. A two-hole rule was used by limiting potential blocks to those interpolated by the first pass. As in the measured resources, eligible model blocks were looked at in longitudinal section and polygons digitally drawn around contiguous areas of appropriately tagged blocks. These polygons were used to classify blocks not already assigned as measured resources as indicated mineral resources.

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All remaining model blocks containing a gold grade estimate were assigned as inferred mineral resources. The classified mineral resources are shown at the Figure 14-18.

Figure 14-18: Mineral Resources Classification, Efemçukuru Mine

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Efemçukuru mine mineral resources, as of December 31, 2019, are shown in Table 14-6. The Efemçukuru mineral resources are reported at a 2.5 g/t Au cutoff grade.

Table 14-6: Efemçukuru Mine Mineral Resources, as of December 31, 2019

		Resource	Grade Au	Contained Au
Deposit / Shoot	Mineral Resource Category	( t x1,000)	(g/t)	(oz x1,000)
South Ore Shoot (SOS) Kestane Beleni	Measured	1,465	7.29	344
	Indicated	244	4.75	37
	Measured & Indicated	1,709	6.93	381
	Inferred	213	4.09	28
Middle Ore Shoot (MOS) Kestane Beleni	Measured	788	8.41	213
	Indicated	681	5.62	123
	Measured & Indicated	1,469	7.12	336
	Inferred	263	3.78	32
North Ore Shoot (NOS) Kestane Beleni	Measured	295	9.84	93
	Indicated	510	8.57	140
	Measured & Indicated	804	9.03	234
	Inferred	378	4.30	52
KBNW Kestane Beleni	Measured	-	-	-
	Indicated	250	8.69	70
	Measured & Indicated	250	8.69	70
	Inferred	447	6.54	94
Kokarpınar	Measured	-	-	-
	Indicated	-	-	-
	Measured & Indicated	-	-	-
	Inferred	2,301	6.64	491
Batı	Measured	-	-	-
	Indicated	-	-	-
	Measured & Indicated	-	-	-
	Inferred	797	8.94	229
Stockpile	Measured	7	6.15	1
Total Mineral Resources	Measured	2,555	7.93	651
	Indicated	1,684	6.84	370
	Measured & Indicated	4,239	7.50	1,021
	Inferred	4,399	6.55	927

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SECTION ● 15          Mineral Reserve Estimates

The Efemçukuru operation commenced commercial production in 2011 and as of December 2019 the mine has produced 3.7 Mt of ore at an average grade of 7.6 g/t Au using combination of Drift and Fill (DAF) and Longhole Open Stoping methods (LHOS).

Diluted and recovered mineral reserves (as of December 31, 2019) are estimated as 4.3 Mt at 5.0 g/t Au, containing 687 Koz’s Au at a cut-off grade of 3.68 g/t Au.

15.1

Cut-off Grades

A range of cut-off grades (COGs) have been calculated to identify ore, incremental ore, and marginal ore.

A planning cut-off grade (COG) of 3.68 g/t Au was calculated based on the 2020 budget costs and a steady state life of mine (LOM) production profile. The 2020 budget costs are supported by 2019 actual production costs. The reference point of the COG is at the entrance point to the processing facility. As such, it refers to the diluted and recovered material reporting to the processing facility.

15.2

Planning Cut-off Grade

The planning COG is defined as the grade of the material, which will generate revenue from the sale of the finished product that is equal to the cost of mining, processing, G&A, sustaining capital costs, treatment and refining of contained metal(s), inclusive of applicable royalties. The COG determined is 3.68 g/t.

The planning COG has been calculated at $1,250/oz Au price to ensure a consistently positive cashflow and an acceptable return on investment under the present and future gold price environments. All cost assumptions are representative of Efemçukuru’s actual costs and are in line with budget costs.

The planning COG is used to begin identifying the potentially mineable material in the mine design process, and is the main input parameter in the stope optimizations process. Efemçukuru is using the Deswik Stope Optimizer software to identify potentially mineable material in the form of mining shapes for both drift & fill (DAF) and longhole open stoping (LHOS) mining methods.

The planning COG is calculated with the following formula:

COG	=	Mining Cost + Processing Cost + G&A Cost + Sustaining Cost
		(Metal Price – TCRC Cost – Royalties) x Process Recovery

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Note the following regarding calculation of the planning COG:

●

Sustaining capital costs include sustaining development capital costs relating to footwall drifts and haulage drives, underground infrastructure, ventilation raises and escape ways.

●

Tailings dam capacity is deemed to be adequate for the duration of the current life of mine, and therefore there are no cost allowances for tailings dam expansions.

●

Silver credits have been used to offset royalties.

●

The process recovery factor used is reflective of the mill recovery factor and payability factor achieved after refining and selling the concentrate produced on site.

●

A single planning cut-off for both LHOS and D&F mining methods were applied for simplicity given the very small differences in cut-offs and the accuracy of the inputs for the cut-off grade calculations.

15.3

Incremental Cut-off Grade

The Incremental cut-off grade (ICOG) is defined as the grade of the material which will generate revenue from the sale of the finished product that is equal to the downstream variable costs of mining, processing, G&A, treatment and refining of contained metal(s) inclusive of applicable royalties. The ICOG is calculated with the following formula:

ICOG	=	Variable Mining Cost + Variable Processing Cost + Variable G&A Cost
		Process Recovery x (Metal Price – TCRC Cost – Royalties)

There are circumstances when mineralized material below the stated planning COG may be added to the production stream by using an incremental cut-off. For example, when additional stopes with grades slightly below planning COG are identified between two high grade stopes, or at the extremities of the mining horizon and do not require additional development. The decision to include incremental material in the mining profile is carefully analyzed and approved by senior engineers and/or managers. The ICOG determined is 1.93 g/t.

Note the following regarding calculation of the planning ICOG:

●

The cost components in the numerator refer to variable costs only. Fixed costs are assumed to be fully covered by mining areas with average diluted grades above the stated planning COG.

●

Transport costs and refining costs (TCRC) contain both the variable and fixed components.

●

Material identified using the ICOG must not displace the mining material above the planning COG. There must be capacity available within the entire system to accept the ICOG for it to be evaluated with only the variable costs.

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15.4

Marginal Cut-off Grade

There are circumstances where mineralized material is already broken (e.g. development material or stockpiles) and is below the stated planning or incremental COG, but still may be deemed economical as most of the costs associated with it have already been spent (sunk costs). In these instances, mining costs are removed from the cost structure.

The marginal cut-off grade (MCOG) is defined as the grade which will generate revenue from the sale of the finished product that is equal to the variable portion of re-handling (or transporting to the processing plant), processing, G&A and final refining of contained metal(s), applied to material that has already been mined, or is stockpiled.

MCOG	=	Mine Re-handling Cost + Variable Processing Cost + Variable G&A Cost
		Process Recovery x (Metal Price – TCRC Cost – Royalties)

Mine re-handling cost can include either transporting cost of the broken development material from underground to the surface processing plant or loading and transporting material from a surface stockpile to the processing plant.

If development waste is to be sent to the surface waste dump, the ore handling cost is the difference between the cost of the material sent to the process and the costs of the material sent to the surface waste dump.

The low-grade MCOG of 1.29 g/t was applied to mineralized development material that must be mined to access higher grade areas. Ideally, all low-grade material would be stockpiled and fed to the processing stream when there is spare processing capacity.

A summary of the cut-off grade calculation inputs are in Table 15-1.

Table 15-1: Cut-off Grade Input and Calculation Summary

Input Values based on actual Metrics and Budget
Cut-off Grade Cost Inputs	Units	Total Cost	Variable Cost	Fixed Cost
Mining Cost	$/t	38.05	22.83	15.22
Process Cost	$/t	29.85	19.4	10.45
General & Administrative	$/t	31.38	4.71	26.67
Sustaining Capital	$/t	8.98	N/A	N/A
Transport &Refinery Cost*	$/t	21.1	N/A	N/A
Gold Price	$/oz	1,250
Metallurgical Recovery (%) **	87.50%
Plant Recovery Factor	93.60%
Payability Factor	94.50%
Settlement Adjustment Factor	1.00%
Outputs
Cut-off Grades Calculated***
Planning Cut-off Grade (COG)	3.68 g/t
Incremental Cut-off Grade (ICOG)	1.93 g/t
Marginal Cut-off Grade (MCOG)	1.29 g/t

* Treatment Cost and Refinery Cost are expressed as a cost per tonne mined.

** = (Plant Recovery Factor x Playability Factor) x (1-Settlement Adjustment)

*** Grade of the diluted & recovered mined ore at the plant entrance point

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15.5

Dilution

15.5.1

Dilution and Ore Losses

Mineable Shape Optimizer (MSO) was used to generate the LOM plan stope shapes according to stope design parameters. MSO gives a detailed estimate of the mining inventory that contains the following dilution types (also refer Figure 15-1):

●

Internal dilution – dilution encapsulated as part of normal stope design (mining shape).

●

Planning dilution – dilution external to mining shape which is attributable to a certain degree of overbreak.

A recovery factor is also applied to allow for incomplete recovery of the design stope. Recovery is affected by the following issues:

●

Incomplete clean-out of the stope during remote bogging.

●

Underbreak during blasting.

●

Complications due to excessive stope failures resulting in abandoning a stope because of excessive dilution.

●

Low sill pillar recoveries.

Figure 15-1:Schematic Showing Dilution Encapsulated by MSO Mining Shape

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The blended planned mining dilution and mining recovery factors are prorated averages between DAF and LHOS metrics and are provided in Table 15-2.

Table 15-2: Planned Mining Dilution and Mining Recovery Factors

Blended Dilution and Mining Recovery Factors
Planned Mining Dilution	16 %
Mining Recovery	96%

15.6

Mineral Reserves Statement

The mineral reserves of the Efemçukuru Project were classified using logic consistent with the CIM definitions referred to in the National Instrument 43-101-Standards of Disclosure for Mineral Projects (NI 43-101). The mineralization of the project satisfies sufficient criteria to be classified into proven and probable mineral reserves. Only measured and indicated mineral resources were converted, using appropriate modifying factors, to mineral reserves. The mineral reserves are inclusive to the mineral resources.

The mineral reserve estimate is summarized in Table 15-3 and has an effective date of December 31, 2019.

Table 15-3: Efemçukuru Mineral Reserves as of December 31, 2019

Category	Tonnes (Kt)	Grade (g/t Au)	Contained Au (Koz)
Proven	1,996.0	6.77	434.2
Probable	1,383.4	5.68	252.4
Proven +Probable	3,379.4	6.32	686.6

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SECTION ● 16          Mining Methods

16.1

Introduction

Efemçukuru mine employs small-scale underground mechanised mining methods to exploit the narrow, high- grade, subvertical mineralization. The present mine plan is based entirely on the proven and probable reserves of 3.4 million fully diluted and recovered tonnes presented in Section 15. The projected mine life is 7 years at current production rate of 510,000 tpa.

Efemçukuru is an epithermal gold deposit comprising two major veins, Kestane Beleni and Kokarpinar, with Kestane Beleni containing the bulk of the ore. Both veins strike north-westerly (320°–340°), dip 60°E to 70°E, and can be traced on surface for strike lengths of over a kilometre. This study describes mining of the Kestane Beleni vein’s four ore bodies or shoots (south, middle, north and northwest), which differ slightly in strike and dip orientation, but the vein and the fault zone is continuous between them (Figure 16-1).

Figure 16-1: Isometric View Showing Mine Areas and Vein Wireframe

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The mine design has been developed to allow flexible access to all of the shoots. Four spiral footwall ramps at each orebody provide access for moving men, equipment, and supplies underground. Advantages of the four-ramp system include increased stope availability, more robust ventilation with increased equipment and labour productivity. All declines are interconnected by link drives, which serve as a secondary egress from the mine, spaced vertically 80–100 m apart.

Ore is truck hauled to a central ore pass system above the underground crusher before being conveyed to surface via a 380 m belt conveyor. The orepass system provides 3,500 tonne surge capacity for underground production with a further 5,000 tonne capacity in bins on surface. Waste rock is hauled to surface via the South 672 Portal or North 656 Portal.

Drift & Fill (DAF) is the primary stoping method used for widths between 2 m and 8 m. This method allows selective recovery of ore within the orebody. LHOS is used where the ground conditions are amenable to large stable spans.

Paste backfill is used as a “free standing” structure to control stability of walls, dilution, and safety for the LHOS. In the DAF stopes, paste backfill is used as the working floor. The paste plant is located near the North 656 Portal.

16.2

Mining Methods

The present mine plan is based on the combination of the following mining methods:

●

Drift-and-fill (DAF)

●

Longhole-open-stoping (LHOS), with two different subversions of this mining method:

o

Longitudinal longhole-open-stoping (LLHOS)

o

Transverse longhole-open-stoping (TLHOS)

The mining method allocations are based on the following:

Drift-and-fill:

●

Shallow, weathered areas close to surface where the HW is not stable enough for large spans.

●

NOS fold hinge where ground conditions are expected to be less favourable.

●

Adjacent to existing DAF mining where LHOS is not practical.

A schematic for a typical DAF production block showing crosscut layout and sublevels is provided in Figure 16-2.

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Figure 16-2: Schematic Cross Section Showing DAF Production Block Layout

Longitudinal longhole-open-stoping

●

Good ground conditions amenable to large stable spans.

●

Deeper areas of the mine.

●

Narrow sections of orebody where DAF results in excessive dilution.

Transverse longhole-open-stoping method is applied to recover the MOS sill pillars on the 555 Level and 595 Level.

A schematic for a typical LHOS production blocks are shown in Figure 16-3.

The current sublevel of 20 m vertical spacing is optimized for DAF mining. The 20-m sublevel is also an appropriate sublevel interval for LHOS given that dilution mitigation is a primary focus for the mine.

Current operations have also established the division of the orebody into independent production blocks of 80 m in vertical extent, with each production block consisting of four sublevels where possible. Production blocks are shown in Figure 16-4.

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Both DAF and LHOS stopes are mined concurrently from multiple production blocks to fulfil production requirements. Overall, the production blocks are mined in a top-down sequence, but stopes within a production block are mined bottom-up (overhand).

Figure 16-3:Schematic Long and Cross Sections Showing LHOS Production Block Layout

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Figure 16-4: View Looking South-West Showing Production Blocks

16.3

Underground Mine Desıgn

The current mine layout has the following features:

●

Four declines (SOS, MOS, NOS and KBNW), each covering approximately 400 m strike extent.

●

Two surface portals (south and north).

●

One surface conveyor adit for conveying crushed ore to the surface crushed ore bins.

●

Four primary ventilation surface exhausts (south, central, north and northwest) and one fresh air raise for NOS.

●

Link drives connect declines and serve as a secondary egress from the mine.

16.3.1

Development Dimensions

Development dimensions are provided in Table 16-1 and are based on current practices and equipment sizes.

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Table 16-1:Development Dimensions

Development	Width (m)	Height (m)	Gradient
Diamond Drill Access	4.5	4.5	1:20 to1:50
Diamond Drill Cuddy	6.0	5.0	1:20
Decline or Incline	4.5	4.5	1:6.7 (15%)
Footwall Drive	4.5	4.5	1:20 to 1:50
Level Access (Decline to FWDRV)	4.5	4.5	1:20 to 1:50
Ore Drive	4.5	5.0	1:50
Stockpile	4.5	4.5	1:20
Sump	5.0	4.5	-1:5
Vent Access	4.5	4.5	1:20 to 1:50
Vent Rise	4.0	4.0	NA
Crosscut	4.5	4.5	-1:6 to 1:6

16.3.2

 Development Stand-off Distances

The following development stand-off distances are used in the LOM plan:

●

Footwall-drive: 50 m lateral standoff (40 m true distance) from stopes/orebody.

●

Decline: 20 m lateral standoff from the footwall drives.

●

Vent raises: 40 m minimum distance from stopes.

●

Infrastructure: 60 m from stopes.

Efemçukuru easily maintains adequate standoff distances for access infrastructure due to the use of footwall drives and DAF stopes requiring 50 m long crosscuts.

16.4

Geotechnical Assessment

16.4.1

Geotechnical Conditions

General rock types and geotechnical and mining conditions at Efemçukuru include:

●

The current reported resources are contained within the Kestane Beleni vein structures known as SOS, MOS, NOS and KBNW.

●

All veins outcrop at surface and a minimum 20m crown pillar is being maintained between the surface and the underground mining.

●

Mineralization consists of moderate dipping (60° to 70°) stacks of various quartz veins that strike north-south.

●

Phyllites and Hornfels are the primary host rocks (silica and calcosilicate altered phyllites are referred to as Hornfels).

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Mineralization is bound by planar fault zones along footwall and hangingwall contacts; there are also minor faults sub parallel to the veins.

●

Inter-veins waste usually carries low background Au grades.

●

There are no in-situ stress measurements but the direction of principal stress is estimated to be horizontal and sub-perpendicular to the strike of veins.

●

The mine is generally considered dry but affected by high seasonal ground water inflows and seepages primarily associated with historic workings and structural features (faults, shears and dykes).

●

Ground conditions in MOS are generally worse than SOS and NOS with the latter having the strongest ground conditions.

●

The KBNW is relatively shallow dipping (50 degrees) and narrow with indications for relatively poor ground conditions amenable for DAF mining method.

16.4.2

Rock Mass Quality

The geotechnical observation of geologic units encountered at Efemçukuru are as follows (Table 16-2):

●

RQD analysis from exploration and delineation drilling indicates that the Hornfels is the strongest unit followed by the veins and phyllite.

●

The rock strength data indicates that the strongest intact rock unit is the mineralised veins followed by Hornfels, with the Phyllite being the weakest of the three rock units.

Table 16-2: Rock Mass Quality

Rock Type	Quality
Hornfels	Competent
Stockworks Hornfels	Competent
Veins	Competent but blocky and jointed
Phyllite	Generally competent
Stockworks Phyllite	Generally competent
Fault zones	Very weak soil like material with clay alteration

16.4.3

Geotechnical Domaining

A geotechnical domain model has been developed and updated for geotechnical logging of exploration and stope definition drilling information. Major fault structures are also mapped and updated in the geotechnical domain model. The model contains RQD, Q’, and Q for rock mass classification.

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16.4.4

Rock Mass Classification

At Efemçukuru the rock mass has been classified by the widely used Q-System (Barton et al., 1974) by adopting characterization logging values to determine Q input parameters. The Q-system was developed to measure tunnelling conditions and support requirements. The rock quality (Q) is a function of six parameters:

Q = (RQDJn) x (JrJa) x (Jw/SRF)

Where:

●

RQD = rock quality designation. RQD is defined as the percentage of intact, sound pieces of core, longer than 100 mm in the total length of core measured.

●

Jn = joint set number. This parameter refers to the number of joint sets identified in the rock mass. The RQD/Jn parameter is a representation of the block size in the rock mass.

●

Ja = joint alteration number. This parameter refers to the type and thickness of infill on the joint.

●

Jr = joint roughness number. This parameter refers to the surface roughness of the joint. The Jr/Ja parameter is a representation of the inter-block shear strength of the rock mass.

●

Jw = joint water reduction factor. Jw is a measure of water pressure, which has an adverse effect on the shear strength of joints by reducing the normal effective stress.

●

SRF = stress reduction factor. SRF is based on one of the following parameters: a loosening load in an excavation through shear zones; rock stress in competent rock; or squeezing loads in plastic incompetent rock. The Jw/SRF parameter indicates the conditions of active stress around an excavation.

The last two parameters (Jw and SRF) are related to the active stress conditions around an excavation. If the rock mass alone is to be characterized, these parameters are removed from the classification system to calculate the modified rock quality (Q’). Q’ is based on rock mass conditions and is the basis for determining N’ from the Mathews Stability Graph Method to assess stope stability (Hutchinson and Diederichs, 1996). Therefore, Q’ is calculated as:

Q’= (RQDJn) x (JrJa)

Where any statements about Q’ rock mass classification ratings are made, it is assumed that:

Jw / SRF = 1

The analysis indicates that the Q’ values vary between 4 and 40 for the vein and main host rocks indicating fair to very good rock quality. Fault zones have a rock quality index of less than 0.3, indicating very poor to extremely poor rock quality.

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16.4.5

Stope Stability Assessment

It is a common practice to classify rock mass quality and assess ground conditions using the Norwegian Geotechnical Institute (NGI) Q system. The geotechnical information is gathered from the underground mapping and from the stope definition drilling. The stope stability is assessed utilizing site-specific stability graph method developed at the mine.

The Mathews stability number modifies the Q' value by three factors to consider the magnitude of the induced stresses in the stope surface, the relative orientation of the major joint set to the surface and the angle of the stope surface.

The modified Mathews stability number is defined as:

N= Q’ x A x B x C

Where:

●

N’ is the Mathews stability number.

●

Q' is the modified Q value.

●

A is the rock stress factor.

●

B is the joint orientation factor.

●

C is the gravity adjustment factor.

The selection of DAF and LHOS mining methods is primarily based on the orebody geometry (width and dip) and the expected ground conditions determined through geotechnical assessment.

Because of insufficient geotechnical data from exploration boreholes for KBNW, mining method in this area has been conservatively based on the DAF mining method. There may be opportunity to use LHOS mining method once sufficient geotechnical information became available from ongoing exploration and planned stope definition drilling prior to mining.

Since start of LHOS operations in 2015, there has been considerable numbers of stope performance data available for the development of a site-specific empirical stability graph for mine design and planning purposes at Efemçukuru (Figure 16-5).

The geotechnical assessment indicates that the current mining method, stope sizes, and mining sequence will not change significantly change over the LOM.

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Figure 16-5: Back Analysis of Stope Performance

16.4.6

Sill Pillar Recoveries

Sill pillars are temporarily left in place and serve to increase the number of production areas in operation at the same time. These pillars are designed to 20 m vertical height, similar to the regular long-hole stope height, and placed every 80 m vertical interval. The sill pillars are recovered, using blind up holes, as the mining progresses upwards to the sill level.

16.4.7

Geotechnical Mine Design Guidelines

Adopted stable stoping spans and mining sequence include:

●

A minimum crown pillar size of 20 m (below surface).

●

Maximum sublevel spacing of 20 m (floor to floor).

●

Maximum stope length of 20 m along and across strike.

●

Continuous retreat transverse blind up-hole stoping in MOS, and SOS sill pillars.

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Minimum one stope pillar separation (of 5 m waste) between two lenses in case of DAF mining.

●

Minimum separation of 60 m between major developments and between stoping and major infrastructures.

●

Preferred sequence for mining the stacked lenses is from the hangingwall towards the footwall.

●

Filling of all stope voids with backfill and tight filling of stopes where multiple lifts of stopes is mined.

●

Pouring 0.5 m thick mesh reinforced 30 MPa grade concrete at the base of sill pillar stopes until the quality of paste fill is improved to allow safe mining directly beneath paste fill.

16.4.8

Ground support

Typical ground support used at Efemçukuru consists of:

●

Fibrecrete – 30 MPa (50 mm or 75 mm thick).

●

Steel mesh (6 mm diameter and 100 mm gauge).

●

Primary bolting – 2.4 m long splitsets.

●

Cable bolts – 6 m, single strand, plain 25-tonne capacity in intersections and along hangingwall contacts in longhole stopes and 9 m long cable bolts in sill pillars.

●

Multiple combinations of the above.

16.5

Backfill

Current mining methods require systematic filling of all stope voids for stability and to enable extraction of adjacent ore. Efemçukuru is currently using a combination of backfill types:

●

Paste fill made from filtered tailings.

●

Development waste rock is being used as unconsolidated rockfill.

●

Cemented rockfill is used where required.

The primary backfill method is paste fill. The current fill recipe is to mix paste fill at yield stress of 100 Pa to 500 Pa (or conical slump of 190 mm to 240 mm) at solids density between 65% and 69% by weight. The current cement content is 7% or 9% by dry solids weight depending on the stope and required duty. The bulk density of the paste fill is 1.7 t/m3.

The pastefill reticulation system has been extended to all SOS, MOS and NOS mining areas, with the exception of KBNW, which is currently being developed. A combination of 4-inch and 5-inch schedule 80 and schedule 40 steel pipelines are used. Paste is delivered underground using an 80 bar capacity Putzmeister positive displacement pump, pumping through a Schedule 40 and 80 steel pipelines and a network of surface and internal boreholes.

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16.6

Mining Rate

Efemçukuru has a sustainable ore production rate of 510,000 tpa. The mine plan has a duration of approximately 7 years.

16.7

Mine Development and Production Assumptions

The development and production assumptions have been prepared on the basis of demonstrated performance, at Efemçukuru. The following assumptions apply to the schedule:

●

Lateral development: 300 m/month mine total

●

Lateral development: 60 m/month for single face advance

●

Development advance rate: 3.5 m/round

●

Longhole stoping rate: 5,000 t/month

16.8

Mining Sequence

Each mine area is divided into vertical production blocks 80 m high (four sublevels).

Within each production block, stopes are mined bottom-up (overhand sequence), with each sublevel mined and filled in its entirety before the overlying sublevel can commence.

For DAF, the intra-production block sequence is such that each cut (lift) must be completed and filled before the crosscut is stripped and the next cut can commence. A schematic outlining the basic DAF sequence is provided in Figure 16-6.

For LHOS, the intra-production block sequence commences at the lowest sublevel at the furthest extents of the ore drive(s) from the crosscut access. Stopes are mined and filled sequentially, retreating back to the crosscut access. When a sublevel is finished in its entirety, stoping on the next sublevel can commence. A schematic outlining the basic LHOS sequence is provided in Figure 16-7.

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Figure 16-6:Schematic Showing DAF Mining Sequence for a Typical Production Block

Figure 16-7:Schematic Showing LHOS Mining Sequence for a Typical Production Block

16.9

Mobile Equipment Fleet

The current equipment fleet is provided in Table 16-3. The equipment levels are generally consistent with the current fleet numbers throughout the LOM plan.

Table 16-3: Current Equipment Fleet

Equipment	Capacity	Number
Twin Boom Jumbo Atlas Copco H282,S2	Twin boom	3
Trucks Sandvik TH430 (or equivalent)	30 t	5
LHD Sandvik LH 307	6.7 t (3.2 m3)	2
LHD Sandvik LH 410	10.0 t (4.7 m3)	2
Cablebolter Sandvik DS421	-	1
Normet Charmec 1605B	-	1
Normet Spraymec	-	2
Normet Transmixer	4.5m3	2

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16.10

Mine Schedule

The mine operates seven days a week and three shifts a day. This annual schedule is equivalent to 365 days per year of operation.

16.10.1

Development Schedule

Since 2010, more than 30,000 m has been developed as of December 2019. The current existing underground mine development and the LOM plan for development is shown in Figure 16-8.

Note: View - looking to the East

Figure 16-8: Current and LOM Plan Mine Development

Development planned for the LOM is summarized in Table 16-4.

Table 16-4: LOM Plan Capital Development

Mine Development	Unit	2020	2021	2022	2023	2024	2025	2026	Total
Operating Dev.
Lateral	m	1,860	1,672	1,766	1,599	1,394	363	88	8,742
Vertical	m	0	42	21	44	22	0	0	129
Sustaining Capital Dev.
Lateral	m	602	755	671	828	143	0	0	3,000
Vertical	m	38	116	101	100	36	0	0	391
Total Lateral	m	2,462	2,428	2,437	2,427	1,537	363	88	11,742
Total Vertical	m	38	158	122	144	58	0	0	520

16.10.2

Production Schedule

Efemçukuru plans to mine 513 Kt of ore at an average grade of 6.8 g/t Au in 2020. The LOM production plan is summarized in Table 16-5.

Table 16-5: LOM Plan Production Schedule

Ore Classification	Unit	2020	2021	2022	2023	2024	2025	2026	Total
Proven	Tonnes (Kt)	378	347	310	308	241	235	177	1,996
	Au g/t	7.45	7.58	7.13	6.36	5.57	5.90	6.55	6.77
	Ounces (Koz)	90	85	71	63	43	45	37	434
Probable	Tonnes (Kt)	136	164	203	205	273	277	124	1,383
	Au g/t	4.93	5.08	4.82	5.30	6.16	6.92	5.44	5.68
	Ounces (Koz)	22	27	32	35	54	62	22	252
Total Ore Production	Tonnes (Kt)	514	511	513	513	515	512	302	3,379
	Au g/t	6.78	6.78	6.22	5.94	5.89	6.45	6.09	6.32
	Ounces (Koz)	112	111	102	98	97	106	59	686

16.11

Ventilation

The primary ventilation layout relies on declines serving as fresh air intakes and longhole raises interconnected between sublevels serving as an exhaust.

Each mine area has a decline coupled with an exhaust raise system to provide primary ventilation. Fresh air flows down the declines and onto each level via the level access and the footwall drives. Secondary ventilation using axial flow fans and vent duct takes primary airflow from the footwall drives and directs it into the ore drives and stopes. The used air returns to the footwall drives under positive pressure and exits the level via the exhaust rises connected to the footwall drives on each level.

The exhaust raise system comprises longhole rise extensions between levels. Each longhole rise extension is approximately 20 m long and has a cross sectional area of approximately 16 m2. Axial flow primary vent fans are installed in underground bulkheads within the exhaust.

The main components of the LOM plan primary ventilation layout are shown in Figure 16-9.

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Figure 16-9:Primary Ventilation Components

16.11.1

Fresh Air Demand

Fresh air demand for the current operations and LOM plan requirements, equipment quantities and power demands are summarized in Table 16-6. The determination of air quantities is based on 0.06 m3/s per kW of power, which is in line with international norms.

Table 16-6:Fresh Air Requirement for Diesel Equipment

Equipment	No of Units	Diesel Engine Power (kW)	Utilization (%)	Required Airflow (m3/s)
Production Fleet
Jumbo - Atlas Copco 282 - S2	4	120	20	6
Longhole Drill Rig - Atlas Copco Simba H1254	1	65	20	1
LHD - Sandvik LH410	3	220	75	30
LHD - Sandvik LH307	2	150	65	12
Truck - Sandvik TH 30SX	2	240	80	23
Truck - Sandvik TH 430	4	315	80	60
Support Fleet
Service Vehicles - Normet	12	110	80	63
Forklift - Manitou	4	101	30	7
Backhoe- Komatsu WB93R-5	2	74	40	4
Light Vehicle	10	120	50	36
Utility Vehicle- Normet RBO 4x4	5	97	60	17
Allowance for Leakage	10%			26
Total				285

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The current mine intake airflow is 320 m3/s, which is considered to be sufficient to adequately ventilate the workings at the current and planned LOM production levels.

The mine has eight Zitron 90 kW axial main fans installed; two per each mining area (Figure 16.9). For the primary fans in the existing mining areas, no major changes are anticipated over the LOM

16.11.2

Ventilation Design Parameters

The ventilation system design is being modelled using Ventsim Mine Ventilation Simulation Software (Ventsim). This software allows input parameters including resistance, k-factor (friction factor), length, cross-sectional area of openings, perimeter, and fixed quantities (volume) of air. The k- factors used are average standards for various types of drifts, raises, and openings. Underground ventilation control requires several sets of ventilation control doors, regulators, and auxiliary fans (of various kW) to direct air quantities to the workings.

16.12

Mining Services

16.12.1

Dewatering

The current mine dewatering system at Efemçukuru is a “clear water” system that consists of primary infrastructure and secondary infrastructure. The primary infrastructure includes large fixed pumps that handles all water feeding into and out of the sedimentation sumps (settling ponds). The secondary infrastructure includes smaller face pumps that handles dirty water before it reaches the sedimentation sumps.

Pumps stations have been designed at approximately 80 m level spacing. Keeping the pump stations within this limit will avoid the need for high pressure pipes and reduce the pump sizes required.

At present the mine experiences a ground inflow of approximately 78 L/s. During periods of high rainfall, however, volumes of water from the underground can be as high as 135 L/s.

The current dewatering capacity is 140 L/s, therefore a project to increase the dewatering capacity has been initialized in 2019 and will be completed by Q2 2020. By adding four new pumps, two each at levels 295 L and 475 L, the capacity will increase to 200 L/s.

In this project, a pumping station will be located at the base of the MOS mining area on 295 level with all water collected from SOS, NOS and KBNW delivered to the 295-main pump station under gravity through pipelines installed along the connecting drives. The clean water from MOS 295 pump station will be pumped to MOS 475 pump station and from there to the WRSP on surface.

16.12.2

Power Supply

The incoming power supply which serves Efemçukuru operations is a 34.5 kV, 50 Hz overhead pole line from the National Grid to the substation at site.

This power line originates from the substation located in the district of Urla, which has the required capacity and is located approximately 20 km to the east of the plant site (Figure 16-10).

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Figure 16-10:Power Line

16.12.3

Compressed Air

Compressed air is only utilised as instrument air (vortex of electrical boxes), at the workshop and for blowing out the main pump sumps. Compressed air is not required for drilling equipment because all drilling equipment has on board compressors.

16.12.4

Explosives Storage and Handling

The explosives and detonator magazine with 25,000 kg capacity is located on level 650, where blasting agents and accessories are stored in two separate rooms. ANFO and bulk emulsion are used for underground production and development.

16.12.5

Delivery of Supplies and Personnel Transportation

Flatbed diesel-powered utility vehicles move supplies including drill parts, explosives, and other consumables from surface to underground work areas. Diesel-powered enclosed personnel carriers transport the crews. Supervisors, engineers, geologists, surveyors, mechanics, and electricians share smaller diesel-powered vehicles.

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16.12.6

Communications

A leaky-feeder radio system provides the primary communication underground. Supervisors and mobile maintenance crews utilize hand held radios. The leaky-feeder radio system is linked to the surface PABX system.

All main ramps and level accesses are covered with fibre optic cables and Wi-Fi hotspots. Wi-Fi telephones are used by supervisors. All underground personnel are tracked with a personnel tracking software.

16.12.7

Maintenance

Preventive maintenance encompasses all activities that prolong the life of equipment and reduce premature failures. Maintenance personnel underground perform preventative and corrective maintenance work including adjustments, lubrication, and refuelling.

All major repair and maintenance on mining equipment including drills, loaders, and trucks is performed in underground workshop at 555 level.

16.12.8

Fuel Storage and Distribution

Diesel fuel is delivered to the mine-site by road tanker and stored in fuel tanks buried on surface, compliant with local Turkish regulations. The storage fuel tank is installed on a concrete pad with concrete berms to prevent contamination in the event of a spillage. All bulk lubricants for operations are stored in the warehouse.

Mine trucks hauling waste rock are refuelled on surface. A lube-fuel truck fuels LHD units, drills, and other underground diesel equipment not reporting to the surface each shift.

16.12.9

Fire Protection, Safety, and Mine Rescue

The North, South and Middle ramps are fresh air escape routes. A total of eight refuge stations are located underground; one on the KBNW, three on the North Ramp and two on the middle ramp and two on the South Ramp. Six of the refuge stations are portable and two of them are permanent. The refuge stations provide a self-contained atmosphere, provide oxygen at controlled rates, and remove carbon dioxide from the air.

Mine rescue equipment and facilities are located at the mine site. Two mine rescue teams are trained with the necessary firefighting and rescue skills. Detailed ventilation plans are updated regularly for the mine rescue teams.

Fire extinguishers are located at key infrastructure locations and at strategic points along each underground sub level. All underground miners are trained in basic safety, first aid, and underground mine survival techniques. A stench gas system is used to warn all employees in case there is an emergency underground.

Fire suppression systems are fitted to all mobile equipment.

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SECTION ● 17          Recovery Methods

17.1

Introduction

Run-of-mine (ROM) ore is crushed underground and then transferred to two ore storage bins on surface via a conveyor. The two ore storage bins allow for blending of different ore types feeding the process plant feed, target a desirable gold/sulfur ratio and reduce contents of penalty elements for concentrate sales. The process flowsheet for the recovery of gold was a combination of gravity concentration and flotation. The cleaner flotation concentrate was subjected to further upgrade by gravity concentration to produce a smeltable grade product. This gravity concentrate was dried and smelted at Efemçukuru to produce gold doré. The remaining flotation concentrate after gravity concentration was thickened, filtered and stored in big bags for shipping. The original process design criteria (PDC) specified gold recovery to gravity concentrate at 30%. However, actual gravity gold recovery from plant operation was only around 1% of total gold production and not economically viable, and hence the gravity concentration circuit was decommissioned in 2018.

A simplified process flow diagram of the current circuit is provided in Figure 17-1. The comminution circuit consists of a semi-autogenous grinding (SAG) mill operated in closed circuit with a pebble crusher, a ball mill operated in closed circuit with hydrocyclones and a flash flotation cell. Entire ball mill discharge is treated in a flash cell to recover the fast-floating liberated sulfide mineral particles and prevent overgrinding of gold containing particles. Overflow from hydrocyclones is sent to a rougher/scavenger flotation bank. Flotation circuit consists of a rougher/scavenger flotation bank and two parallel cleaner flotation banks. Concentrates from flash flotation cell and the first two cells of the rougher/scavenger bank are combined and upgraded in cleaner bank 1. Rougher cells 3-6 concentrate are treated in cleaner bank 2. Concentrates from cleaner banks 1 and 2 are combined and sent to the final concentrate thickener. Underflow of the concentrate thickener is filtered and the filtered concentrate is stored in big bags for shipping. Rougher/scavenger flotation tailings are passed through a Knelson gravity concentrator whose concentrate is further upgraded in a Salter gravity concentrator. The tailings after Knelson gravity concentrator are sent to a tailings thickener. Salter concentrate is circulated to the flash flotation cell while Salter tailings are recycled to the cyclone feed pumpbox. The final tails are filtered. A portion of the tailings is used in the underground paste backfill plant, and the rest is dry stacked in the tailings storage facility (TSF).

Column flotation cells will be installed for upgrading the final concentrate. Figure 17-2 shows location of the column flotation cells which are going to be installed and commissioned in July 2020. The tail of the column flotation cells is circulated to cleaner bank 1 or to another suitable location.

17.2

Process Performance

A summary of performance data of the process plant is presented in Table 17-1, using the data recorded between 2013 and 2018. Original process design values are also included for comparison. Current annual throughput and plant availability are higher than that of design values.

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Figure 17-1: Simplified Process Flow Diagram of Efemçukuru Process Plant

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Figure 17-2:Simplified Process Flow Diagram of Efemçukuru Process Plant with Column Flotation Cells

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Table 17-1: Process Design Criteria of Efemçukuru Process Plant and Actual Plant Performance Data

	Unit	Design	Actual Operation (2013-2018)
Availability/Utilization
Annual ore throughput	dry t	401,500	460,421
Daily (calendar) ore throughput	dry t	1,100	1,238
Crusher plant availability	%	68.5	80.5
Crusher plant utilization	%	-	37.4
Crushing processing rate	dry t/h	67	140
Grinding and flotation availability	%	90.0	96.8
Grinding and flotation utilization	%	-	94.7
Grinding and flotation processing rate	dry t/h	50.9	55.5
Ore Properties
Head grade – Gold	g/t	10.0	7.7
Head grade – Silver	g/t	17.8	17.3
Total sulfur	%	2.73	2.96
Bond abrasion index	g	0.670	0.567
Bond crushing work index	kW.h/t	17.0	23.4
Bond ball mill work index	kW.h/t	17.2	20.4
JK drop weight S.G.		2.91	2.89
JK drop weight Axb		46.4	43.4
JK drop weight ta		0.51	0.35
SCSE	kW.h/t	-	9.9
Gold recovery	%	92.0	93.7
Gold grade in the concentrate	g/t	-	69
Silver grade in the concentrate	g/t	-	150
Annual concentrate tonnage	dry t	33,000	47,226
Moisture content of concentrate filter cake	%	8.0	11.0

Operating crushing and ball millwork indices of the plant feed are significantly higher than the design values, which means that the current plant feed is harder than the ore samples used in comminution tests during design stage. Moisture content of the concentrate filter cake is 11.3%, which is 3% higher than the design value. Other parameters are similar to the design values.

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17.3

Process Flowsheet and Equipment

17.3.1

Crushing

The crushing circuit is located in the underground mine. The ore is fed to the jaw crusher from two ROM bins, originally by chain feeder but more recently by gravity. There are gates over the ROM bins to minimize oversized rocks being fed to the crusher. A scalping grizzly removes any fine material of less than 80 mm prior to crushing. The crusher produces a product of 80% passing 90 ~ 100 mm.

Certain sections of the crushing circuit have been revised to improve operational efficiency since the beginning of the operation. In 2014, the chain feeders were removed and isolation gates were installed. The design value of scalping grizzly aperture size was 150 mm. In 2015, the scalping grizzly aperture size was reduced first to 100 mm and then to 80 mm. The scalping magnet was replaced with a stronger unit to reduce tramp material.

The crushing circuit operates at a nominal tonnage of 140 tph with a circuit availability of 80% and a utilization of 37%. The crushed material is transferred to two bins of 4,200 tonne capacity. Material can be directed to either of the two bins by the operator to improve plant blending ability. Figure 17-3 shows average availability and utilization figures of the crusher between 2013 and 2019.

Figure 17-3: Crusher availability and utilization between 2013 and 2019

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17.3.2

Grinding and Classification

The grinding circuit consists of a semi-autogenous mill (SAG), a ball mill (BM) and a pebble crusher (SABC). Fresh mill feed particle size is 80% passing 90 ~ 100 mm. Particle size of cyclone overflow is 80% passing 53 µm. The 5.5 m x 3.0 m grate discharge SAG mill is installed with a 1,500 kW variable speed motor. The SAG mill draws typically an average of 745 kW, corresponding to a motor load of about 50%. Grinding ball top size is 125 mm and the ball charge ranges between 10% and 15%. Particle size of SAG mill discharge is approximately 80% passing 600 µm.

The pebble crusher is a short head cone crusher equipped with 132 kW motor. Pebbles of 100% passing 70 mm are crushed to 80% passing 10~12 mm. The pebble crusher processes approximately 15% of the fresh mill feed in the form of pebbles.

The 3.6 m x 6.0 m overflow ball mill (BM) is installed with a 1,500 kW variable speed motor. The mill draws typically an average of 891 kW, corresponding to a motor load of 59%. The mill operates using grinding ball top size of 40 mm at typically 27% ball charge. Grinding media consumption is approximately 1.04 kg/t in the SAG mill and 1.37 kg/t in the ball mill.

The entire ball mill discharge is treated in a flash flotation cell of 10 m3 volume. Flash flotation tailings and SAG mill discharge are collected in a cyclone feed pumpbox, together with cleaner tail and the tailing from the Knelson/Salter combination and classified in a cyclone cluster. The cyclone cluster was revised in April 2016, by replacing three (2 operating, 1 standby) 381 mm hydrocyclones with eight (5 operating, 3 standby) 250 mm hydrocyclones with 49 mm apex and 90 mm vortex. The new cyclone cluster gives better control of particle size in the overflow. Particle size of cyclone overflow is approximately 80% passing 54 µm.

17.3.3

Flotation

Gold in Efemçukuru ore is mainly associated with sulfide minerals, predominantly pyrite, galena, sphalerite and trace amounts of chalcopyrite. Hence, the process plant is based on bulk sulfide mineral flotation and produces a gold concentrate containing bulk sulfide minerals. The flotation circuit consists of rougher/scavenger flotation for recovering gold-bearing sulfide minerals in the cyclone overflow, flash flotation for recovering liberated gold-bearing sulfide minerals in the ball mill discharge and cleaner flotation for upgrading rougher/scavenger concentrate and flash concentrate to produce the final gold concentrate. Type of flotation reagents, their dosages and addition points are given in Table 17-2.

Activator (copper sulfate) and sulfidizer (sodium bisulfide) are introduced into SAG mill discharge, i.e. to the cyclone feed pumpbox. Copper sulfate is used as an activator for sphalerite flotation and pyrite. The sodium bisulfide is mainly used for surface cleaning of slightly oxidized sulfide minerals, particularly pyrite in this case. Collector and promoter are added at two points, ball mill discharge (flash feed) and conditioning tank, where cyclone overflow is conditioned with flotation reagents. Sodium isobutyl xanthate (SIBX) is of moderate strength and works as the collector in bulk sulfide mineral flotation. In addition to SIBX, Aero 8045, which is a dithiophosphate formulation promoter, is used to improve flotation of gold bearing sulfides. Both reagents are added to the ball mill discharge (prior to flash flotation) and to the conditioning tank before rougher/scavenger flotation.

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Table 17-2: Reagent Types, Addition Points and Dosages in Efemçukuru Flotation Plant

Reagent	Addition Point	Dosage (g/t) of fresh ore
Activator (CuSO4.5H2O)	SAG Mill Discharge	34
Sulfidizer (NaHS)	SAG Mill Discharge	15
Collector (SIBX)	Conditioning Tank	15
	Ball Mill Discharge	16
Promoter (Aero 8045)	Conditioning Tank	33
	Ball Mill Discharge	27
Frother (OrePrep F-549)	Conditioning Tank	3
	Ball Mill Discharge	5

Characteristics of the flotation cells at Efemçukuru plant are given in Table 17-3. There is one flash cell of 10 m3 volume, six 20 m3 tank cells (rougher/scavenger flotation) and six 5 m3 tank cells (cleaner flotation) in the circuit. A sanding problem was previously observed with the rougher/scavenger cells, but the issue was corrected with modifications performed in 2017-2018.

Table 17-3:Characteristics of the Flotation Cells at Efemçukuru Flotation Plant

	Flash Flotation	Rougher/Scavenger Flotation	Cleaner Flotation
Flotation Cells	Dorr-Oliver 10 m3	Wemco 20 m3	Dorr-Oliver 5 m3
Cell Type	DO10m³ Flash	Wemco 20RT	DO5RT
Number of Cells	1	6	6
Motor rpm	990	888	975
Rotor rpm	280	196	312
Effective Surface Area (m2)	0.9	4.7	2.4
Launder Lip Length (m)	5.0	14.6	12.0
Impeller Diameter (mm)	500	762	400
Aspect Ratio	Cone Top and Bottom	0.71:1	1:1

Ball mill discharge combined with the tailing gravity concentrate is treated in the flash flotation cell to recover the liberated gold bearing particles and minimize overgrinding. Tailings of the flash flotation are circulated to the cyclone feed pumpbox. The cyclone overflow is treated in the rougher/scavenger flotation bank. Concentrate of the first two cells is combined with the flash concentrate and upgraded in flotation cleaner bank 1. The rougher/scavenger concentrate of the remaining cells (cell 3 to 6) is treated in cleaner bank 2. Concentrates of the two cleaner circuits are combined and sent to the final concentrate thickener. Cleaner tailings are circulated back to cyclone feed pumpbox. Tailings of the rougher scavenger bank are further treated in a gravity separation circuit, which consists of one Knelson centrifugal concentrator and one Salter separator to minimize gold loss from the final tailing. Total residence times for flash flotation cell, rougher/scavenger bank, cleaner bank 1 and cleaner bank 2 are approximately 5, 49, 20 and 15 minutes, respectively.

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Column flotation as the third cleaner flotation stage increases concentrate quality and reduces concentrate tonnage with negligible gold recovery loss. Hence, a project was initiated in late 2019 for design and installation of column flotation cells in the circuit. Based on mass balance and simulation studies performed using plant surveys and pilot scale column flotation tests, two flotation column cells of 1.8 m (diameter) x 8 m (height) size with a distributor as the feeder to the column cells. The two column flotation cells can be operated in parallel or in series as the third cleaner flotation stage depending on plant requirements. The cells will be installed and commissioning in Q3 of 2020.

17.3.4

Dewatering

The final concentrate is dewatered in a 5-m diameter high rate thickener. An anionic type flocculant is used at 8 g/t dosage (based on equivalent fresh mill feed tonnage) to improve settling rate of the solid particles. The concentrate slurry is first treated in a de-aeration tank prior to the thickener to reduce air entrainment and prevent generation of froth at the surface of the thickener. The concentrate thickener produces an underflow of 55% w/w pulp density at approximately 5.0~6.5 tph capacity. The underflow is transferred to a surge tank of 58 m3 volume, which provides approximately 11 hours of residence time. The water recovered from thickener overflow is sent to the process water tank. Solid content of the overflow water was significantly reduced after the modifications (installing de-aeration tank, water sprays and revision of feed line) made in 2017-2018.

Concentrate thickener underflow is further de-watered in a 1,500 mm x 1,500 mm filter press with 42 plates. A filter cake of 11% moisture content is produced and shipped in big bags (approximately 1.5 tonne per bag).

Flotation tailings are fed to a 13 m diameter high rate thickener, achieving a thickener underflow solids content of 55% w/w. An anionic flocculant is used at 10 g/t (based on equivalent fresh mill feed tonnage) for flocculation of the tailings particles. Thickener overflow is circulated to the flotation plant as process water. Thickener underflow is sent by a positive displacement diaphragm pump to a surge tank of 246 m3 in the tailingfilter plant, with 8.4 hours of residence time. Part of the tailing material is used for making paste for backfilling in the underground mine, and the remaining tailing is filtered and then dry stacked in tailings storage facility (TSF). Hence, three pressure filters of 1,500 mm x 1,500 mm size with 60 plates in each filter are used to produce dry stacked tailings of 16% moisture content.

Distribution of tailings between TSF and paste plant (for underground backfill) are illustrated on an annual basis in Figure 17-4. Approximately 48% of the tailing is used as backfill material in the underground mine.

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Figure 17-4:Distribution of Flotation Tailings to TSF and Paste Plant

17.3.5

Paste Backfill

In the paste backfill mode, thickened flotation tailings were pumped from the surge tank as described above to the tailing filters by a dedicated centrifugal slurry pump. The centrifugal pump was replaced with a positive displacement pump in 2019. The filters discharge directly onto the filter discharge conveyor and subsequently to the backfill delivery conveyor. This conveyor is reversible and can discharge either onto a stockpile or the paste backfill mixer feed conveyor.

The paste backfill mixer feed conveyor transfers the filtered tailings to the paste mixer. Paste components are directed to the feed chute of the paste mixer along with cement dosed at a proportion of the measured dry weight of the tailing solid. The cement addition is about 2.8 t/h for 35 dry t/h capacity and makes approximately 8% of the backfill material. The mixed paste discharges into a backfill paste hopper and from here the paste is pumped to the underground reticulation system.

Performance of paste plant has been improved with some modifications made since 2016. The paste mixer was replaced in 2016 with a larger unit with 132 kW shaft motor and four 7.5 kW high-speed mixers to improve mixing of paste material. With this modification, quality of the paste was improved and shut down time for cleaning the mixer was significantly reduced. In 2017, in order to improve paste quality and minimize dust problem, a vortex mixer was installed and the cement feeder line was revised.

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17.3.6

Process Water

Process water is supplied from the water treatment plant and from the two thickener overflows. The excess water is sent to the water treatment plant. Water treatment plant supplies clean water to the reagent preparation, pump gland water, thickener spray water and other facilities, which require high quality water. Process water is used in grinding, cyclone classification, flotation launder spray, flocculant dilution and general hose down requirements. Sulfide mineral content of Efemçukuru ore is relatively low and does not contain a significant amount of soluble mineral components affecting water quality. The flotation reagent scheme is also simple and does not have a major impact on process water quality. Therefore, process water is used without treatment and has no negative impact on flotation performance.

17.3.7

Operational Performance

Figure 17-5 shows availability and utilization of the flotation plant from 2013 to end of 2019. After commissioning and ramp-up periods, operational performance has increased to the expected levels in 2012. Availability and utilization of the flotation plant in recent years are 97% and 96%, respectively.

Figure 17-5:Availability and utilization of flotation plant between 2013 and 2019

17.4

Production

Summary of the production data from 2013 to the end of 2019 are given in Table 17-4. Mill throughput has increased sharply after ramp-up period and continued to increase gradually throughout the mine life to over 520 kt per annum.

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Table 17-4: Summary of Production Data between 2013 and 2019 in Efemçukuru Flotation Plant

	Unit	2013	2014	2015	2016	2017	2018	2019
Milled ore tonnes	Kt (dry)	414	437	455	477	482	499	521
Feed gold grade	g/t	8.87	8.34	7.82	7.40	7.01	6.77	7.03
Feed silver grade	g/t	13.9	13.3	20.8	14.9	20.9	19.8	19.3
Concentrate tonnage	Kt (dry)	40.3	44.7	44.2	46.8	54.2	59.5	60.5
Concentrate gold grade	g/t	83	75	75	69	59	54	57
Concentrate silver grade	g/t	136	120	176	139	174	156	155
Recovered gold in concentrate	Koz	107	108	106	104	102	103	111
Recovered silver in concentrate	Koz	177	170	218	211	300	298	301
Gold recovery	%	93.3	93.3	93.7	93.0	94.6	94.6	94.0

Gold grade of plant feed is about 8.0 g/t in the first four years of operation, but has decreased gradually to 6.7 g/t in 2018. Unlike gold grade, silver grade increased from approximately 13 g/t to 20 g/t in the last two years of operation.

Gold grade of the concentrate ranges between 75 g/t and 87 g/t during 2011-2015 period, and decreases gradually to approximately 54 g/t in the following years. Gold grade in the concentrate is influenced by the lower feed grade, and more correctly by the lower gold/sulphur ratio during recent years. Silver grade in the concentrate increases from about 120 g/t to 160 g/t in parallel with its increased head grade in the flotation feed.

Annual average gold and silver production and gold recoveries are given in Table 17-4. Gold recovery increases from 87% in the ramp-up period up to nearly 95% in 2018. This increase is very much related to better understanding flotation behaviour andalso improved mill availability/utilization.

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SECTION ● 18          Project Infrastructure

Efemçukuru gold mine consists of an underground mine with a process plant and ancillary facilities on surface situated southwest of Izmir in an area easily accessible by road. The site has local access to a large seaport in Izmir, approximately a 60 km drive from site, and an international airport in Menderes, approximately a 30 km drive from site. The surface infrastructure at the Efemçukuru site to support the mining and processing operations include the following:

●

Site access roads

●

Plant site roads

●

Water treatment, supply and distribution

●

Sewage collection and disposal

●

Diesel fuel storage

●

Power supply and distribution

●

Ancillary facilities

18.1

Site Access and Local Roads

The access to site is from Izmir through regional paved and gravel roads. The roads are narrow and winding with some isolated steep grades; however, the roads are paved and in good condition and easily passable by commercial trucks.

18.2

Site Access Road

The site is accessed from the regional roads. A 7.5 km long bypass road around Çatalca Village was completed in 2013. Since then the village road has not been used to access mine site.

The mine access road is a 2.7 km paved road from the region network to the gate house at the plant site. The road is private access and includes a weighbridge (Figure 18-1).

During operations, flotation concentration is bagged on site and a convoy of 40 tonne trucks haul the concentrate to a storage facility in Sarnıç, 25 km from site, on a daily basis, then the concentrate is transported to the port when the sale is recognized. A pilot vehicle leads the convoy and provides warning for oncoming vehicles; the convoys leave the plant at the same time each day so locals are aware.

18.3

Site Layout

Efemçukuru mine site consists of the concentrator process plant, ancillary buildings, tailings filtration/backfill plant at the north portal; the filtered tailings storage facility; development rock dump, and waste rock sedimentation pond (WRSP) both in the valley below the south portal. The site layout was designed to limit the disturbed footprint and the amount of tree removal and to blend in with the surroundings. A range fence primarily for control of domestic and wild animals surrounds the entire site. The site excavations have balanced cut and fill quantities to limit aggregates and rock from off the site sources.

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Figure 18-1: Mine Main Access Road

The plant site is located on the west side of the Kokarpinar valley at an elevation of 605 masl. It consists of the ore storage bins, concentrator building and ancillary facilities. The ore bins are fed from the underground crusher by a belt conveyor which daylights above the concentrator on a small pad with an access road to allow service vehicles. The site has been designed to limit the disturbed footprint by terracing the facilities into the topography to avoid a large excavations. The location of the plant site lends itself to the utilization of the existing forestry roads for access. The plant site drains to the catchment pond located at the north of the site. Water from the catchment pond is pumped to the water treatment plant.

Site roads were designed to follow the alignment of the existing exploration roads where possible in order to minimize site disturbance and removal of trees. The roads allow access from the plant to the filtration plant and the north portal; the development rock dump, the south portal, and filtered tailings storage. A haul road allows access from the filtration plant to the filtered tailings storage area. Internal roads are sealed for dust control. Figure 18-2 shows a general view of the mine site.

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Figure 18-2: General View of the Mine Site

18.4

Surface Tailings and Development Rock Management

Mining waste within the scope of the Efemçukuru gold mine project consists of waste rock and process tailings; tailings are dry stacked or used as cemented paste-fill. The waste rock dump (WRD) and tailings storage facility (TSF) are lined facilities and built pursuant to the provisions of the Mining Waste Regulation.

Within the borders of operations area, there are two operating TSFs and one operating rock dump. In addition, there is another TSF and another rock dump permitted for future use but not currently required to handle the reserves stated in this report.

The three dry tailings storage facility (TSF) areas are referred as the central TSF, south TSF (both operating) and north (future expansion). The two waste rock dump areas are referred to as central WRD (operating) and north WRD (future expansion). The 2015 EIA Report subject to the EIA-Positive decision presents three WRD areas, each being central, south and north, respectively; however, the central and south WRD areas were merged to establish the Central WRD area. The surface areas of the storage facilities are given in Table 18-1. The total capacity of the Central, South and North Dry TSFs is sufficient for approximately 2.4 Mm3 of dry tailings, whereas the total capacity of the Central and North WRD storage areas is sufficient for approximately 1.6 Mm3 of waste rock. The mining area has storage capacity in excess of the waste rock and tailings amounts to be stored on the surface throughout the mine life.

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Table 18-1: Surface Areas of the Dry tailings and WRD Storage Areas at Efemçukuru Gold Mine

Project Units	Area  (ha)
Central Dry tailings Storage Area	6.64
South Dry tailings Storage Area	5.3
North Dry tailings Storage Area	3.20
Central WRD Storage Area	8.2
North WRD Storage Area	6.28

By the end of the mine’s economic life cycle, an approximate 3.2 Mt of waste rock will be stored at the mining site. While waste placement continues at the central WRD storage area, rehabilitation of approximately 2 ha has been accomplished in the recent years in compliance with the rehabilitation requirements.

By the end of the economic life of the mine, approximately 3.87 Mt of dry tailings will be stored in central and south TSFs. Additionally, mining waste facility license for dry tailings storage facility located at the central valley were obtained from the General Directorate of Environmental Management under the Ministry of Environment and Urbanization (MEU). Construction works in the south TFS were completed in 2019 and approved by the (MEU). Construction works are not currently planned at the North TSF areas.

The tailings storage design includes the following elements:

●

Storage of dry stacked filtered tailings in a facility incorporating compacted tailings structural shells with 3H:1V (18°) outer slopes on the downstream and upstream sides of the facility to provide structural stability for the pile.

●

Underdrain and base liner system comprised of a fully-lined base, a central rock drain, and a toe drainage blanket to collect seepage from the tailings pile.

●

Engineered closure cover system comprised of a synthetic cover over the tailings, overlain by a 1 m-thick store and release soil cover system.

The development rock storage facility design includes the following design elements:

●

Overall 3H:1V slope contoured to promote drainage, reduce erosion and to provide long-term stability.

●

Underdrain and grouted/sealed foundation system below the liner to promote drainage of any collected water and limiting water infiltration into the underlying bedrock.

●

Drain system over liner to collect mine rock pile seepage.

●

Closure cover system comprised of a synthetic cover over the mine rock, overlain by a 1 m-thick store and release soil cover system.

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The detailed design for the north TSF and north rock dump storage facilities will be completed during a subsequent stage of the Project if required. This design will incorporate information from the planned geotechnical site investigations in the area of the storage facilities, along with other relevant project information as it becomes available. The detailed design will also include detailed staging plans for waste disposal for performing concurrent reclamation during the mining operations, as well as coverage of other design factors.

Figure 18-3: View of Tailings Area

18.5

Fire/Fresh Water Supply Storage and Distribution

The plant site requires 75m³ of water per hour, which is provided from mine dewatering, site collection and recycled water from the process. Mine water is pumped from underground to the waste rock sedimentation pond (WRSP) located below the development rock dump. The WRSP stores mine water (dewatering) and contact water collected by the diversion ditches around the perimeter of the project site. Mine and contact water stored in the WRSP is treated and used in the process with the excess water treated, tested, and released back to the environment or stored in the east pond for use in the process when dewatering flows do not meet process requirements.

Figure 18-4 shows the WRSP (forefront of photo). The South 676 Portal is located on the north side hill above the valley (right side of photo). Filtered tailings storage is located further up the valley (not shown in photo). Figure 18-5 shows the east pond.

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Figure 18-4: WRSP Sedimentation Pond

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Figure 18-5: East Pond

The water treatment plant located on the southeast of the plant site supplies fresh-treated water to a fire and fresh water storage tank located on the hill above the process plant, which is then distributed by gravity to the process plant. The fresh water tank serves double duty as storage of both fire and fresh water. Fire and fresh water reservoirs inside the tank are separated with the use of a standpipe inside the tank to draw off the fresh water off the top of the fresh water standpipe assuring storage for firewater is maintained.

Figure 18-6 shows a view of the water treatment plant.

The firewater distribution system consists of a dedicated buried firewater main and hydrant system for the plant site and ancillary buildings. Hose cabinets are placed within the process plant and ancillary facilities, supplemented by portable fire extinguishers in all facilities. Hose stations located at 50 m intervals and automatic sprinklers over the drive protect underground conveyor. Ancillary buildings are equipped with automatic wet sprinkler systems throughout.

The hypochlorinator and two potable water storage tanks are located at the mill site. Each potable water tank has a capacity of 60m3. Buried piping distributes potable water to the ancillary facilities. The potable water is suitable for general use in the facilities but not for consumption. A tanker truck supplies drinking water to a storage tank located at the process plant.

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Emergency showers and eyewash stations are situated throughout the process building.

Figure 18-6: Water Treatment Plant

18.6

Sewage Collection and Treatment

The number of personnel working on site is nearly 450, sewage and domestic wastewater resulting from the personnel and domestic ancillary facilities are treated in the wastewater treatment plant at a rate of approximately 47 m3 per day.

The treatment plant starts with a basket grating removing coarse solid waste such as sand and stones protecting, subsequent equipment. The wastewater then enters into the second stage compensation pond and is ventilated for biological treatment with a fixed flow rate. The biological reactor operates with a sequencing batch reactor (SBR) and the blower operating for 20 hours of aeration to ensure biological treatment. After aeration, the blower is stopped to be prepared for the next stage which is sedimentation. After 2 hours, the treated water is pumped to the treated water chamber for 2 hours. The biologically treated water is taken from the treated water chamber by a pump and then passes through first the sand filter, then the cartridge filtration and lastly the active carbon column and taken to the discharge chamber.

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18.7

Waste Disposal

There is a waste management system currently being used at the project site. It has a surface area of approximately 1.73 ha surface area and allows temporary storage of domestic wastes, packaging wastes, electronic wastes, scrap metals, waste oils generated on the site (kitchen wastes), as well as limited amount of hazardous wastes (used batteries, cartridges of photocopier, fax, printer, containers contaminated with chemicals, etc.). The area ensures collection of different types of wastes are segregated in different containers or areas to ensure waste is properly recycled or disposed of off-site and the area is isolated from other infrastructure and waste is protected from atmospheric conditions before their disposal. The area helps facilitates the collection of wastes, recycling and recovering of material effectively and safely, offering environmental and economic benefits.

All domestic wastes are collected regularly by the municipality from the covered waste containers placed in designated areas around the site. Colour coded containers segregate different wastes and are covered to protect from wind and rain

18.8

Power Supply and Electrical Distribution

18.8.1

General

The electrical system has been sized to take into account the process loads, water treatment plant loads, mining loads, and the ancillary loads, such as the workshop/warehouse mine dry/canteen and administration building (Table 18-2).

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Table 18-2: Estimated Load List – Efemçukuru

	Total	Connected	Standby	Operating	Annual
	Rating	Load	Load	Load	Operating
Area	(kW)	(kW)	(kW)	(kW)	(MWh/a)
Process Plant Area
C1 – Primary Crushing	170	170	0	143	957
D0 – Crushed Ore Storage & Reclaim	45	45	0	38	315
E0 – Air Supply & Distribution	225	115	110	95	725
E1 – Grinding & Classification	2,920	2,750	170	1,850	15,588
E2 – Gravity Concentration	41	39	2	33	270
E3 – Pebble Crushing	147	147	0	130	1,082
E4 – Flotation	455	405	50	273	2,300
E5 – Concentrate Dewatering & Loadout	181	116	65	100	805
E6 – Reagents	75	47	28	28	250
E7 – Gold Room	172	168	4	0	0
F2 – Tailings Filtrations & Paste	1800	1500	300	700	5,300
K4 – Water Supply & Distribution	305	171	134	120	450
Total Process Plant	6,536	5,673	863	3,510	28,042
Mine Area
B1 – Mining Equipment	6,235	4,875	1360	2,768	22,593
J3 – Truck Shop	94	94	0	55	458
Total Mine	6,329	4,969	1360	2,823	23,051
Site Area
G1 – Water Treatment	1155	735	420	350	2798
F1 – Tailings Thickening	252	171	81	82	690
K3 – Fresh Water	64	34	30	32	247
J0-Ancillary Facility	420	260	160	140	1130
Total Site	1891	1200	691	604	4,865
Total Efemçukuru	14,756	11,842	2,914	6,937	55,958

18.8.2

Power Supply

The existing power line is 34.5 kV, 50 Hz overhead pole line from the National Grid to the mine substation. The power line originates from the substation located at Çamlı village in the district of Urla, approximately 15 km to the west of the plant site. The substation has sufficient capacity to meet the mines current and future needs.

There is a second substation available also near Çamlı village if required (Figure 18-7).

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Figure 18-7: Power Line

18.8.3

Site Power Distribution

The site layout shows the location of the incoming power line, main substation, and site power distribution. The following description of the plant site electrical distribution system is in accordance with the site electrical single line diagrams.

The incoming 34.5 kV overhead power line from the national grid has terminated at the main substation located at the north of the plant site. The main substation consists of a main disconnect, metering facilities, main transformer (34.5/6.6 kV,10 MVA) and 6 x 6.6 kV feeder positions in a walk-in type outdoor rated enclosure. Distribution from the main substation location will include the following feeders:

●

6.6 kV power line extending north to provide power to filter plant area (electrical room – ER4) and the north adit area (ER5) which supply underground mining equipment and services

●

6.6 kV power line extending east to provide power to the concentrator building substation loads including:

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■

6.6 kV grinding area motor loads in the concentrator building (ER3)

■

6.6 kV feeder cable to serve the low voltage loads in the concentrator building (ER3) and buildings adjacent to the concentrator building such as the administration, canteen, lab, and gatehouse buildings

■

6.6 kV feeder cable to serve the low voltage loads in the Mine Dry, Truck Shop and Warehouse

■

6.6 kV power line extending east to provide power to the Water Treatment Plant substation loads including:

–

6.6 kV feeder cable to serve the low voltage loads in Water Treatment, Sewage Treatment, Project Department and Contractor Area

–

6.6 kV overhead power line extending east to west to provide power to Air Quality station and Surface work office building

●

6.6 kV feeder cable installed along the primary crusher conveyor structure to underground distribution which serves the primary crusher (ER1)

●

6.6 kV power line to provide power to the south portal and underground equipment(ER8)

●

6.6 kV power line to provide power to HV Compensation Panel

These rooms contain area low voltage motor control, control system cabinets, HVAC, lighting, and provision for power correction equipment where required. The electrical rooms are installed on concrete supports where appropriate and adjacent to structures where there are concentrations of electrical equipment needing requiring power and control.

Underground 400 V cables exit from ER3 location to provide power to the power distribution centers and motor control centers of the ancillary buildings. Motor control centers are complete with motor starters, contactors, disconnect switches, transformers, panels, circuit breakers, and fuses.

Stand-by generators are located around site located with enough power to supply for operation of emergency lighting and power essential drives in the event of a power outages in site.

18.8.4

Electrical Equipment and Materials

All electrical equipment are rated for a minimum elevation of 750 masl and an ambient temperature range of 0°C to 40°C and is certified by Conformité Européenne (CE).

18.8.5

General Power and Lighting

Power outlets consist of 16 A, 220 V, 1 Phase, 50 Hz plug in receptacles for small tools, and 63 A, 400 V, 3 Phase, 50 Hz receptacles for welders and other heavy equipment. The lighting fixtures types are:

●

High pressure sodium (HPS) fixtures, sized as required for lighting of the mill areas

●

Fluorescent and LED fixtures for office and electrical rooms

●

LED flood light fixtures mounted on the buildings will be supplied for yard lighting

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18.8.6

Power and Control Cables

Distribution cables are aluminium-armoured polyvinyl chloride (PVC) jacketed, cross-linked polyethylene insulated conductors.

Cables run from the electrical room to the electrical equipment and devices, mounted on cable tray/racking throughout the mill building and direct buried between buildings, unless otherwise noted.

18.9

Communications

A communications network is existing with satellite technology for voice, fax, and Internet service.

18.10

Ancillary Facilities

The ancillary facilities were designed utilizing concrete panels and blockwork as far as practical to maximize the use of locally available materials and methods and blend in with the local architecture. The following is a general description of the ancillary facilities included on the Efemçukuru site.

18.10.1

Process Buildings

The process buildings including the concentrator building and filtration plant building are structural steel buildings with concrete panel siding. The reagent area, bag storage area, electrical, and mechanical rooms and offices are located in a two-storey structure annexed to the main process building.

The concentrator building has a 20 tonne overhead crane servicing the grinding and flotation areas, and a 7.5 tonne overhead crane servicing the reagent and concentrate storage areas. Figure 18-8 shows the process plant and crushed ore bins.

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Figure 18-8: Process Plant and Crushed Ore Bins

The filtration plant is located beside the North 656 Portal and on the side of the hill to utilize the topography.

Figure 18-9 is a view of the tailing filter/backfill plant and north portal located in the higher up the valley.

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Figure 18-9: Filtration Plant and North Portal

18.10.2

Laboratory

A laboratory does not exist on site. Laboratory service is provided by a contractor in Izmir.

18.10.3

Workshop and Warehouse

The workshop and warehouse is a pre-fabricated concrete/blockwork building. The building has been designed to provide facilities for maintenance and repair.

The workshop and warehouse include two indoor truck bays, an outdoor wash bay, machine shop, welding shop, and electrical/instrumentation work area. Maintenance and planning personnel have offices located on the second floor.

The warehouse has 144 m2 of indoor storage area and an outdoor fenced secure storage is included

18.10.4

Administration Building

The administration building is approximately 500 m2, including space for finance, accounting and administration personnel. The general manager also has an office in this building.

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18.10.4.1

Mine Dry and Canteen

The mine dry is a three-storey building approximately 1,600 m2 in area. The mine dry is equipped with lockers and baskets for 250 miners. The building includes offices for the mine manager, mine superintendent, geology manager and superintendent, technical office superintendent, engineering and geology, shift supervisors, and safety engineer as well as a lamp room, kitchen, guest rooms and meeting rooms.

Figure 18-10 is a view of mine building beside south portal. The old mine building was located close to the mine site gateway. Construction of the new mine building was completed in 2018.

Figure 18-10: Mine Building beside South Portal

The canteen is approximately 540 m2 with a kitchen to provide hot meals.

18.10.4.2

Gatehouse

The main gatehouse is located at the access to the plant site. It includes a reception area, a training room and space for safety and security personnel.

18.10.4.3

Accommodation and Transportation

The personnel from Izmir and the surrounding communities are transported to the site by 20- seat buses.

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SECTION ● 19 Market Studies and Contracts

19.1

Gold Market

The original business plan for the Efemçukuru concentrate was to transport it in isobags to the Kisladag site for further treatment in a concentrate treatment plant (separate from the Kisladag operation) to produce final doré. Early in the operation’s life concentrate’s was also sold to refiners; this proved to be much more economically viable. Since 2012 all concentrate has been sold to traders or refiners. Except a few spot sales to South Korea and Bulgaria the main destination of the concentrate has been Chinese smelters. Contracts with traders have had similar conditions and the market has been stable. In recent years contract terms have become more stringent in terms of concentrate quality and process improvements continue to adjust the quality to have a salable product at the best market rates and contract terms.

19.2

Gold Price

Gold price movement shown in Figure 19-1 represents our sales prices since 2017. The average gold (Au) prices realized in 2018 and 2019 were US$1,267 and US $1,431.

Figure 19-1: Gold Price Movement

19.3

Contracts

Sales over the life of mine have been primarily with one customer, which covered most of the annual production with best terms. Due to changing contract terms in 2019, sourcing of additional customers allowed the operation to establish the best alternatives and price realization for the concentrate sale. Major shipments to the newer customers began in March 2019. Expansion of concentrate storage facilities in early 2019 allowed more storage capacity permitting more flexibility in shipments.

Contracts are generally agreed with treatment charges, refining charges and penalties based on the quality of the concentrate and the amounts of any excess deleterious materials.

Recent contract terms reflect market realities due to tighter penalties for deleterious elements and higher transportation and refining costs. These newer terms have been incorporated in the mineral reserves determination and economic evaluation.

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SECTION ● 20 Environmental Studies, Permitting and Social or Community Impact

20.1

Introduction

Eldorado Gold Corp. through its 100% owned Turkish subsidiary Tüprag metal Madencilik has been operating the Efemçukuru gold mine and ancillary processing facilities since 2011 on its mineral licences near Efemçukuru village of Menderes District, Izmir Province, in the Aegean Region of Turkey. The current reserve base of the operation is 3.38 Mt of ore, at a grade of 6.32 g/t Au. The total area of the project site is approximately 126.5 ha and is enclosed by a fence.

Based on current reserves there are seven (7) more years of expected operations at a processing rate of 520,000 tonnes of ore per year. The underground mining methods include mechanized cut-and-fill and longhole stoping. Mine development rock is excavated and stored on surface in engineered facilities, which are designed to accommodate approximately 1.46 Mm3 with 0.33 Mm3 of capacity remaining. The beneficiation process produces a flotation concentrate. The flotation concentrate is transported from the site by trucks for sale to refiners. Tailings from ore processing are dewatered and ware returned underground as backfill (48%) or stored on surface in engineered facilities; the tailings facility is designed to store approximately 2.56 Mm3 with a remaining capacity of 1.45 Mm3.

Eldorado’s comprehensive EIA report, received an Environmental Positive Certificate in September 2005. This report determined that the proposed mine development will affect physical components of the existing environment at the site. The report also proposed an Environmental Management Plan (EMP) to prevent or mitigate any impacts of the Project.

An inspection committee regularly visits Efemçukuru site. The purpose of this committee is to check if the present mining operations are carried out within the applicable laws, regulations and EIA commitments. The committee, which has been visiting regularly since September 2007, is composed of experts from various government agencies. During the inspections, duplicate samples are collected from monitoring boreholes and surface water, analyzed in impartial and accredited laboratories, and the analysis results are regularly evaluated. In addition, every six months, soil samples are collected from 17 separate locations around the site and are analyzed and regularly evaluated by a soil commission. The inspections by public institutions and organizations are conducted before mining and during the operation, period as well. It is expected that they will continue during the closing period.

20.2

Air Quality

Efemçukuru Project is subject to the "Regulation Amending Industrial Air Pollution Control Regulation" published in the Official Gazette dated December 20, 2104 and numbered 27277 pursuant to the relevant provisions of Turkey’s Environmental Law no. 2872.

Air quality on the site and in the surrounding rural area may be affected by project activity during construction and operation of the mine.

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The measured values within the operation and the plant's impact area are monitored online together with the meteorological values by the environment department. The relevant data are reported regularly to the ministry of environment.

20.2.1

On Site Air Quality Measurement

Emissions that would result from project activity and their effects on local air quality were evaluated by modelling studies prior to construction. Ground level concentrations of NO2, CO, HC, SO2, and Pb in particulate matter and total particulate matter were estimated and these values were compared with limits given in the Turkish Air Pollution Control Regulation, 2004 (APCR). All predicted levels of these contaminants are monitored regularly and are below the APCR limits. Additional studies showed that dust concentrations measured at 3 m from the crushing, office, and filtered tailings disposal sources would be below limits set in the Turkish Industrial Originated Air Pollution Control Regulations, 2004. Particulate matter is measured regularly at construction areas, along truck routes, and near settlements. Mitigation measures, such as the use of dust filters, road paving, and water spraying on haul roads, at the crushers, and conveyors is undertaken to ensure compliance with APCR limits. After mine closure and restoration of the site, no residual sources of particulate or gaseous emissions will exist.

20.2.2

Off Site Air Quality measurement

Dust Measurements are made in Efemçukuru and Kavacık villages by using the high volume dust collector, BAM-1020, manufactured according to the measurement standards of the Ministry of Environment and Urbanization.

The high volume dust collector, BAM-1020, manufactured according to the measurement standards of the Ministry of Environment and Urbanization, has been installed in the location determined as a result of the survey for the air quality observation station location, carried out around the Health Protection Strip together with the relevant government agencies.

There are three air quality-monitoring stations: one in the health protection strip, one in Efemçukuru village, and one in Kavacık village.

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Figure 20-1: Air Quality-Monitoring Station

20.3

Water Quality

The water component of the local environment at the site includes groundwater and surface runoff.

Within the scope of the monitoring program implemented in the construction and initial operation phases, the surface and ground water quality and quantity are monitored at the sampling stations in a manner to allow identification of any possible impacts related to the project activities. Monthly water quality monitoring continues by on-site measurement, sampling and analysis methods to ensure assessment of the efficiency of the environmental measures taken during the activities in the mine closing phase.

Since 1998, sampling work have been carried out to identify the existing situation of the surface water quality. Surface water quality sampling works are carried out by Eylul University departments including Geological Engineering and Environmental Engineering Department, Microbiology and Wastewater Laboratories. Site measurements are carried out at monitoring locations, and the water samples are taken and sent to independent laboratories by university members.

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The measurements of the surface water quality (physical, chemical and biological parameters are carried out monthly at the monitoring stations along Kokarpinar stream and their compliance with Water Pollution Control Regulations is reported.

In order to monitor the underground water quality, underground water sampling are performed at monitoring boreholes approved and certified by the State Hydraulic Works at the mine site and surrounding area.

Figure 20-2: Water Quality Monitoring

The parameters analyzed in ground water samples during the water quality tests are:

●

Temperature

●

Electrical conductivity

●

Dissolved oxygen and pH

●

Total suspended solid matter (SSM) and total soluble solid matter (TSSM)

●

Carbonate (CO3), bi-carbonate (HCO3), hydrogen sulphide (HS), nitrite nitrogen (NO2-N), ammonium nitrogen (NH4-N), nitrate nitrogen (NO3-N), sulphate (SO4),

●

Phosphorus (P), sodium (Na), calcium (Ca), magnesium (Mg), chloride (Cl), fluoride (F), aluminum (Al), arsenic (As), barium (Ba), cadmium (Cd), cobalt (Co), Copper (Cu), Iron (Fe), mercury (Hg), manganese (Mn), nickel (Ni), lead (Pb), selenium (Se), zinc (Zn), chromium (Cr), and chromium (Cr+6)

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20.3.1

Surface Water

Baseline surface water quality was well documented in the EIA. Surface water contained insignificant traces of fertilizer and pesticides but fecal coliform and biological oxygen demand levels were above regulatory norms and indicated pollution from agricultural runoff. No industrial sources of water pollution exist in the area.

With the mine development, several changes in water quantity and quality occur. Runoff from precipitation and on-site springs are collected and channelled around the engineered tailings and uneconomic rock disposal sites and diverted to a settling pond for use as process water. Excess water is treated and discharged to the Kokarpinar Stream.

Domestic wastewater emanating from the site is treated by a rotating biological contactor and tile field and complying with Turkish Water Pollution Control, 1987 (WPC) regulations. This source will cease to exist post-closure.

20.3.2

Contact Water

The filtered tailings were tested and are predicted to be non-acid generating and therefore contact and underdrain water will not generate ARD post-closure.

Contact and underdrain water from the development rock dump area may contain levels of metals non-compliant with WPC regulations. It will continue to be monitored throughout the life of the project and treated similarly to mine drainage water before being used as process water or discharged as surface drainage after treatment. The uneconomic rock dump runoff has the most potential to be of concern post-closure. However, the expected volume is small and the neutralizing potential of 80% of the blended development rock will be sufficient to neutralize any ARD generated within the dump over the long term.

The volume of potentially non-compliant drainage is expected to be small. This expectation is based on the following information and assumptions, derived from the EIA and humidity cell test results subsequently provided by Eldorado:

●

Average annual precipitation at Efemçukuru: 877.0 mm

●

Average annual evaporation: 435.1 mm

●

Net annual average precipitation: 441.9 mm

●

Surface area of development rock dump: 22,700 m2

●

Volume of precipitation on dump: 0.4419 m x 22,700 m2 = 10,031 m3/a

●

Amount of development rock in dump at closure: approximately 540,000 tonnes

●

Potentially acid generating (PAG) development rock comprises 18% or: 0.18 x 560,000 = 100,800 tonnes

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Given the diversion of surface runoff and on-site spring discharges away from the dump, cell-by-cell macro-encapsulation of the PAG rock and establishment of a substantial vegetated soil cap, as proposed, conservatively assumes that at most only 25% of precipitation infiltrates the development rock dump soil cover. This would equal 0.25 x 10,031 = 2,507 m3/a. Since only 20% of the development rock is PAG, the volume of potentially non-compliant contact water would be about (0.2 x 2,507) = 502 m3/a, or 1,375 L/d, on average

Acid-base-accounting, short-term leachability tests, and long-term humidity cell tests of PAG development rock indicated that the leachate is weakly acidic and that the neutralization potential of 80% of the development rock greatly exceeds the acid generating potential of the remaining 20%. It is suggested that selectively stacking the 20% of PAG development rock over and in contact with the non-PAG rock will utilize the available excess NP and reduce the dissolution of a hazardous material (As) in the residual underflow or seepage from the dump.

20.3.3

Underground Water

Underground mine workings receive groundwater inflows and seepage. Some of this inflow may contact acid generating rock and consequently contain dissolved metals.

This water is pumped to the surface and monitored for contaminants. If it does not meet regulatory criteria or process water requirements, it will be directed to a settling pond and treated before being used as process water. The pond operates with a minimum 0.6 m freeboard and is sized to contain an additional 1-in-100 year 24-hour storm event.

20.3.4

Discharge Water Quality in Industrial and Domestic Wastewater Treatment Plant

The basic principle of the water management established within the mine site has been designed in a manner so that the water in contact with the mine site is collected, treated and used in the process. Therefore, the basic principle is to use the available water and, to treat it according to the existing regulation standards where discharge is necessary.

For this reason, sedimentation ponds have been created to collect and control water for internal water management in the mine. When the water collected in these ponds surpasses the process requirement, they can be discharged after being treated in the treatment plant. The treatment plant effluent quality, however, will comply with the existing regulation standards. The quality of the discharge water from the domestic wastewater treatment plant established within the mine site also complies with the provisions of the existing regulation.

Efemçukuru has an active industrial wastewater treatment plant with a design flow of 12,240 m3/day and an integrated continuous wastewater monitoring system (CWMS). The Ministry of Environment approved this plant in July 25, 2018. Thanks to the CWMS, the effluent of the treatment plant is monitored online by the Ministry of Environment.

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In addition, the domestic wastewater generated by the Eldorado personnel is treated separately in the domestic wastewater treatment plant, which has a design flow of 47 m3/day and was approved by the Provincial Directorate in June 9, 2010.

20.3.5

Water Supply

Water needed for the project area is obtained from the following sources.

●

Water seepage out of rock and waste storage, mineral processing and other operation areas (contaminated water)

●

Water from the seeps in locations other than the operation area (non-contaminated water)

●

Seepage water from underground (contaminated water)

●

Water generated by the moisture content of the aggregate ore (contaminated water)

Process water is supplied by water filtered from the catchment ponds, by dewatering of wastes, dewatering of flotation concentrate filter press and process water recovered in other areas of the plant. In addition, part of the contaminated seepage water from the rock and waste storage, mineral processing and other operation areas including the underground are also used in the process after being treated in the treatment plant. Excess water is discharged into Kokarpınar Creek after treatment. The non-contaminated seepage water and precipitation from outside the operation area are directed via ditches away from the operations into nature water channels.

The treated water is used as utility water only including lavatory flush tanks and washbasin taps. The water is treated in the existing water treatment plant, by reverse osmosis plant and chlorination.

Water bottles are used to supply drinking water for all Eldorado personnel.

20.4

Land Use

Land use within the concession area falls into three categories: agricultural land (vineyards), forestry land and treasury land. Forestry land makes up about 80% of the project area (126.5 ha) and treasury land makes approximately 1%. The remaining area can be classified as private land. As of May 2018, Tüprag is the owner of 100% of the private land within the operation area.

Progressive reclamation of the waste storage areas are undertaken. As the waste storage areas are filled on a cell-by-cell basis, soil cover from the stored reserve is applied and a vegetation cover using native species is planted. At closure, only the last waste cells, the demolished building sites and roadways will remain to be re-contoured, covered with the stored topsoil and seeded, or replanted with native species.

The success of re-vegetation will be monitored throughout the life of the Project and a limited post-closure period, allowing for supplementary seeding or planting if required. Adherence to the proposed land use management plan will ensure that the mix of plant species on the site will be restored and soil erosion prevented.

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Reverse and restoration of the site to its current uses will depend on socio-economic conditions prevailing after closure.

20.5

Flora and Fauna

While the project will displace some species through human presence and activity, including the generation of noise and habitat removal, these species will be able to return to the restored site after closure. The small area of disturbance; covering about a third of the licensed 73 ha area and the relatively short duration of the entire Project, will likely lead to natural recovery of the plant and animal populations at the site.

20.5.1

Flora

Most of the 218 identified plant species are widely distributed; only 15 are endemic to Turkey. One plant species was identified as under threat, but it is not endemic to Turkey.

The baseline vegetation studies for the EIA were carried out in September 2003 and March, April, and May of 2004, representing all seasons, and were well documented. This information has been used to design the pattern for the re-vegetation of the Project site during operations and after closure. Adjacent areas, undisturbed by the project, also provide a guide of local plant species for the natural re-population of the plant communities.

20.5.2

Fauna

Efemçukuru Project site is not classified as a protected habitat area under Turkish or international law. Sixty-three animal species occurring on site are listed under some degree of protection on selected Turkish and international conservation guidelines.

Wildlife species (mammals, birds, reptiles, and amphibians) were identified and enumerated during four visual surveys of the three habitat types on the site: forest and degraded forest, agricultural land, and fluvial (streamside) habitat. The survey identified 8 mammals, 62 birds, 7 reptiles, and 1 amphibian. None are endemic (restricted) to the local area or severely threatened by habitat loss.

20.6

Approvals and Permits

The process of obtaining the necessary permits for a mining operation in Turkey is similar to the European Union EIA Directive process.

Table 20-1 lists key Project permits obtained prior and during mining to date, including the date and the governmental authority that issued them.

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Table 20-1: Permits Obtained Prior to Mining and During Mining at Efemçukuru

Name of Permit	Issue Date	Issuer
Mining License	1999-04-20	Ministry of Energy and Natural Resources
Mining License	2013-08-19	Ministry of Energy and Natural Resources
EIA Permit	2005-09-08	Ministry of Environment and Urbanization
EIA Capacity Expansion Permit	2012-12-31	Ministry of Environment and Urbanization
EIA Capacity Expansion Permit	2015-11-17	Ministry of Environment and Urbanization
Trial Permit	2011-06-01	Governor of İzmir
Workplace Opening Permit	2012-05-24	Governor of İzmir
Workplace Opening Permit	2013-02-27	Governor of İzmir
Workplace Opening Permit	2015-07-13	Governor of İzmir
Workplace Opening Permit	2016-03-18	Governor of İzmir
Workplace Opening Permit	2017-07-01	Governor of İzmir
Pre-Environmental Permit	2011-10-28	Ministry of Environment And Urbanization
Environmental License	2012-07-12	Ministry of Environment And Urbanization
Environmental License	2015-01-30	Ministry of Environment And Urbanization
Environmental License	2017-01-19	Ministry of Environment And Urbanization
Environmental License	2018-04-27	Ministry of Environment And Urbanization
Forestry Permit Exploration Drilling	2006-07-12	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling Access Road	2006-07-12	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling	2007-04-17	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling Access Road	2007-12-31	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling	2007-12-31	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling Access Road	2011-04-11	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling	2011-04-11	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling Access Road	2013-03-14	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling	2013-03-14	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling Access Road	2017-05-04	Ministry Forest And Water Affairs
Forestry Permit Exploration Drilling	2017-05-04	Ministry Forest And Water Affairs
Forestry Permit Power Line	2010-12-22	Ministry Forest And Water Affairs
Forestry Permit Power Line	2017-05-04	Ministry Forest And Water Affairs
Forestry Permit Operation	2008-02-25	Ministry Forest And Water Affairs
Forestry Permit Operation	2010-07-28	Ministry Forest And Water Affairs
Forestry Permit Operation	2012-04-30	Ministry Forest And Water Affairs
Forestry Permit Operation	2017-05-04	Ministry Forest And Water Affairs
Purchase and Use of Explosive Certificate	2008-12-02	İzmir Security Directorate
Purchase and Use of Explosive Certificate	2009-12-31	İzmir Security Directorate
Purchase and Use of Explosive Certificate	2011-12-01	İzmir Security Directorate
Purchase and Use of Explosive Certificate	2014-11-18	İzmir Security Directorate
Purchase and Use of Explosive Certificate	2017-11-08	İzmir Security Directorate
Underground Explosive Magazine	2013-09-18	İzmir Security Directorate
Energy Permitting	2010-06-21	TEDAŞ
Provisional Acceptance Certificate of Yelki Substation	2011-12-26	TEDAŞ
After upgrading switchgear in Yelki Substation, new Provisional Acceptance Certificate	2014-08-27	TEDAŞ
Provisional Acceptance Certificate of Mine Site Electrical Installation	2011-04-12	TEDAŞ
Provisional Acceptance Certificate of Bademler Substation	2011-03-21	TEDAŞ

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20.7

Conclusions

Efemçukuru mine is in commercial production since Q3, 2011. The following conclusions were drawn from this review:

●

An inspection committee regularly visits Efemçukuru mine site since 2007. The committee checks if the present mining operations are executed within the applicable laws, regulations and EIA commitments. Efemçukuru mine of Tüprag Metal Madencilik San. ve Tic. A.Ş. is subject to the "Regulation Amending Industrial Air Pollution Control Regulation" published in the Official Gazette dated December 20, 2104 and numbered 27277 pursuant to the relevant provisions of the Environmental Law no. 2872.

●

Water and mine waste water quality was adequately addressed and mitigated during the project design. Within the scope of the monitoring program implemented in the construction and operation phases, the surface and ground water quality and quantity are monitored in the sampling stations to allow identification of any possible impacts related to the project activities.

Given the adequate restoration of wildlife habitat including re-vegetation with the species now present and removal of the 750 m culvert on Kokarpinar Creek, natural repopulation by displaced species will take place and post-closure liability for terrain and wildlife and habitat restoration is unlikely.

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SECTION ● 21 Capital and Operating Costs

21.1

Capital Cost estimate

Capital expenditures for mining, processing and infrastructure were completed in 2011, prior to the start of production. All capital expenditures prior of December 31, 2019 are considered sunk costs and therefore not included in the economic evaluation. The mine is fully constructed and operating, and actual costs form the basis of future operating and sustaining cost estimates.

21.1.1

Mine Capital Cost

Table 21-1 is a summary of the sustaining and growth capital costs. The mining sustaining capital costs include mine development, paste backfill borehole development, purchase of additional equipment, equipment leasing costs and health and safety initiatives. All costs are expressed in US dollars.

In addition to the sustaining costs listed, the economic analysis presented in Section 22 also includes additional capital cost classified as growth capital. A column flotation system will be installed in 2020 in the process plant for $1.55M. The column flotation system will improve the quality of the concentrate produced.

Table 21-1: Mine Capital Cost Summary by Year (in US$ x 1,000)

Sustaining Capital	2020	2021	2022	2023	2024	2025	2026	Total
Mine Development	1,379	1,878	1,663	1,999	384	0	0	7,304
Mine Infrastructure & Equipment	8,303	7,848	8,030	8,179	7,550	0	0	39,909
Process Plant	2,890	3,230	1,700	850	850	0	0	9,520
Exploration	548	0	0	0	0	0	0	548
Health & Safety	220	40	40	45,000	0	0	0	345
G&A	813	65	93	60	65	0	0	1,097
Total	14,153	13,060	11,526	11,133	8,851	0	0	58,723
Growth Capital	2020	2021	2022	2023	2024	2025	2026	Total
Process Column Flotation	1,553	0	0	0	0	0	0	1,553

21.2

Operating Cost Estimate

21.2.1

 Cost Estimate Basis

The underground mine operating costs were estimated based on actual 2019 operating costs and 2020 budget estimates that allow for maintaining a steady state production profile.

The underground operating costs include all consumables (ground support, explosives, services, cement, aggregates and fuel) and equipment required to meet the development and production schedule objectives. The operating unit costs for mobile equipment and fuel consumption rates were largely obtained from historic mine data. Labour requirements were developed to support the operation and maintenance of the fleet and for the general operation of the underground mine. All these estimates are in line with manpower levels.

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General and administrative costs were estimated based on a projected personnel list with salaries indicative of local standards, and annual allowances for general supplies.

Process operating costs were based on estimated annual consumption of process reagents, major wear parts, and utilities. Budget quotations were obtained for supply of all significant consumables and utilities. Power consumption was calculated based on electrical load lists.

Unit rates representative of a steady state peak production profile are summarized in Table 21-2.

Table 21-2: Unit costs

Area	Unit Costs (US$/t)
Mining	33.28
Operating Development	5.66
Geology	4.77
Plant Maintenance	8.54
Risk Management	3.59
Processing	21.31
Administration	19.23
Finance	6.93
Purchasing	1.63
Total	104.93

21.2.2

Operating Cost Estimation

Annual variable operating costs were estimated by using the 2019 actual unit variable costs and 2020 budget costs to capture cashflows for the steady state 510 K tpa production rate. (Table 21-3).

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Table 21-3: Efemçukuru Operating Costs by Year (in US$ x 1,000)

	Units	2020	2021	2022	2023	2024	2025	2026	Total
TONNES & GRADE - TOTAL
Total Ore	Kt	514	511	513	513	515	512	302	3,379
Ore Grade AU	g/t	6.78	6.78	6.22	5.94	5.89	6.45	6.09	6.32
Ore Grade AG	g/t	17.01	15.64	12.16	12.75	15.29	18.38	18.14	15.47
Mined Ounces AU	Koz	112	111	102	98	97	106	59	687
Mined Ounces AG	Koz	281	257	201	210	253	303	176	1,681
VARIABLE COSTS
Mining	USD	10,261	10,203	10,242	10,244	10,277	10,231	6,028	67,486
Operating Dev	USD	2,404	2,216	2,309	2,124	1,830	470	114	11,467
Geology	USD	1,471	1,462	1,468	1,468	1,473	1,466	864	9,672
Plant Maintenance	USD	2,852	2,836	2,846	2,847	2,856	2,843	1,675	18,755
Risk Management	USD	277	275	276	276	277	276	163	1,820
Processing	USD	7,117	7,077	7,104	7,105	7,128	7,096	4,181	46,807
Admin	USD	1,482	1,474	1,479	1,479	1,484	1,478	871	9,746
Finance	USD	534	531	533	534	535	533	314	3,515
Purchasing	USD	125	125	125	125	126	125	74	825
TOTAL VARIABLE COSTS	USD	26,523	26,198	26,383	26,202	25,987	24,518	14,282	170,094
FIXED COSTS
Mining	USD	6,841	6,802	6,828	6,829	6,851	6,821	4,018	44,991
Operating Dev	USD	1,603	1,477	1,540	1,416	1,220	313	76	7,645
Geology	USD	980	975	979	979	982	978	576	6,448
Plant Maintenance	USD	1,536	1,527	1,533	1,533	1,538	1,531	902	10,099
Risk Management	USD	1,568	1,559	1,565	1,565	1,570	1,563	921	10,311
Processing	USD	3,832	3,810	3,825	3,826	3,838	3,821	2,251	25,204
Admin	USD	8,398	8,350	8,382	8,383	8,411	8,373	4,933	55,229
Finance	USD	3,029	3,012	3,023	3,024	3,033	3,020	1,779	19,919
Purchasing	USD	711	707	710	710	712	709	418	4,676
TOTAL FIXED COSTS	USD	28,497	28,219	28,383	28,265	28,156	27,128	15,874	184,523
TOTAL OPERATING COST	USD	55,020	54,417	54,766	54,467	54,143	51,647	30,156	354,617
Royalty	USD/t milled	107.08	106.51	106.78	106.18	105.21	100.81	99.91	104.93
	USD/oz sold	543.87	542.31	595.67	618.34	614.74	535.82	561.31	572.22
	USD	4,809	4,782	4,400	4,206	4,182	4,562	2,539	29,479
TOTAL CASH COST	USD	59,828	59,199	59,167	58,673	58,325	56,209	32,695	384,096
TONNES & GRADE - TOTAL	USD/t milled	116.43	115.87	115.36	114.38	113.33	109.71	108.32	113.66
	USD/oz sold	591.40	589.97	643.53	666.09	662.22	583.16	608.57	619.79

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SECTION ● 22 Economic Analysis

22.1

Basis

An undiscounted cumulative cashflow of US$341.2 M and an NPV of US$278.3 M are estimated based on the following:

●

Discount rate of 5%

●

Development cost of US$2,154 per metre of waste development

●

Operating cost of US$38.05 per ore tonne

●

Total gold recovery of 87.5 % after processing, smelting and refining

●

Gold price of US$1,400 per ounce

●

Silver price of US$18.00 per ounce

●

Transport and refining cost of US$118.48 per mined Au ounce. All costs are derived from the 2019 actual and 2020 budget numbers

●

Pre-tax basis

22.2

Cashflow

The scheduled cashflow and cumulative NPV are shown in Figure 22-1 and Table 22-1.

Figure 22-1: LOM Plan Cashflow and Cumulative NPV

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Table 22-1: Cashflow Summary ((in US$ x 1,000)

	Unit	2020	2021	2022	2023	2024	2025	2026	Total
TONNES & GRADE - TOTAL
Total Ore	Kt	514	511	513	513	515	512	302	3,379
Total Ore Grade AU	g/t	6.78	6.78	6.22	5.94	5.89	6.45	6.09	6.32
Total Ore Gold	Koz	112	111	102	98	97	106	59	687
Total Ore Grade AG	g/t	17.01	15.64	12.16	12.75	15.29	18.38	18.14	15.47
Total Ore Silver	Koz	281	257	201	210	253	303	176	1,681
Operating Development
Lateral	m	1,860	1,672	1,766	1,599	1,394	363	88	8,742
Vertical	m	0	42	21	44	22	0	0	129
Sustaining Capital Dev.
Lateral	m	602	755	671	828	143	0	0	3,000
Vertical	m	38	116	101	100	36	0	0	391
COSTS
Operating Cost	$	-19,553	-19,442	-19,516	-19,520	-19,583	-19,496	-11,486	-128,596
Operating Development	$	-4,007	-3,694	-3,849	-3,540	-3,050	-783	-190	-19,112
Processing Cost	$	-15,336	-15,250	-15,308	-15,311	-15,360	-15,292	-9,009	-100,865
G&A Cost	$	-16,124	-16,032	-16,093	-16,097	-16,149	-16,076	-9,471	-106,043
Sustaining Dev Cost	$	-1,379	-1,878	-1,663	-1,999	-386	0	0	-7,304
Sustaining Cost	$	-12,774	-11,183	-9,863	-9,135	-8,465	0	0	-51,419
Total AISC	$	-69,172	-67,478	-66,292	-65,601	-62,994	-51,647	-30,156	-413,340
AISC per tonne	$/t	-134.62	-132.07	-129.26	-127.88	-122.41	-100.81	-99.91	-122.31
Growth Capital	$	-1,553	0	0	0	0	0	0	-1,553
Total All-in Cost	$	-70,726	-67,478	-66,292	-65,601	-62,994	-51,647	-30,156	-414,893
AIC per Tonne	$/t	-137.64	-132.07	-129.26	-127.88	-122.41	-100.81	-99.91	-122.77
REVENUE
Head AU ounces	Koz	112	111	102	98	97	106	59	687
Head AG ounces	Koz	281	257	201	210	253	303	176	1,681
Total Recovery percent	%	87.5%	87.5%	87.5%	87.5%	87.5%	87.5%	87.5%	87.5%
Payable AU oz's	Koz	98	97	90	86	85	93	52	601
Payable AG oz's	Koz	246	225	175	184	221	265	154	1,471
Revenue AU	$	137,204	136,435	125,559	120,007	119,319	130,173	72,444	841,139
Revenue AG	$	4,426	4,046	3,158	3,313	3,986	4,769	2,772	26,470
Total Recvenue	$	141,630	140,480	128,717	123,320	123,305	134,942	75,215	867,610
Transport & Refining Cost	$	-13,270	-13,195	-12,144	-11,607	-11,540	-12,590	-7,006	-81,352
Royalty	$	-4,809	-4,782	-4,400	-4,206	-4,182	-4,562	-2,539	-29,479
Cashflow	$	52,826	55,026	45,881	41,907	44,589	66,143	35,514	341,885
Discounted (5%) Cashflow	$	49,098	48,707	38,678	33,646	34,095	48,168	25,862	278,254
Cumulative Cashflow	$	52,826	107,852	153,732	195,639	240,228	306,372	341,885	341,885
Cumulative NPV	$	49,098	97,805	136,483	170,129	204,224	252,392	278,254	278,254

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22.3

Sensitivities

Sensitivities have been carried out to test the robustness of the project. The sensitivity of the project NPV to the major variables is provided in Figure 22-2 and Figure 22-3.

The project is most sensitive to gold price. A 40% reduction in the gold price results in zero NPV.

The next most-sensitive project variable is the operating cost. An increase of 40% in the operating costs alone does not significantly impact the viability of the project, with a resultant NPV of approximately US$235 M.

Figure 22-2: LOM Plan NPV Sensitivities based on Au Price Variation

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Figure 22-3: LOM Plan NPV Sensitivities based on Cost Variations

22.4

Opportunities and Risks

The following opportunities were identified:

●

The conversion of new and the delineation of additional resources at Batı Vein and Kokarpinar may enable concurrent mining and increased levels of mine production and/or extended mine life.

●

Ore sorting (optical or conventional) to improve grade, decrease waste treatment and allow increased milling capacity and reduced mining costs which may allow reduced mining marginal grade resource due to reduced COG.

●

Improvements in payability and metallurgical recoveries through contract negotiations and continued improvement in ore treatment has the potential to further improve the project economics.

●

KBNW mining zone is planned to be extracted by DAF mining method, but there may be an opportunity to utilise the more cost effective LHOS mining method pending further geotechnical testing.

●

There is an opportunity for reducing cement costs by utilising alternative binders such as low heat cement. Further binder screening tests will be required to identify and select optimum binder that works with Efemçukuru tailings. Use of deep cone thickeners or paste thickener technology has the potential to improve paste plant throughput and reduce operating costs of paste fill and tailings filtration.

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The following risks were identified:

●

Groundwater inflow might be higher than anticipated as there is lack of sufficient orebody and hydrogeological knowledge in KBNW. This risk will be mitigated by installing excess pumping capacity.

●

Localized poorer ground conditions may result in increased dilution, reduced mining recoveries and productivities. Detailed geotechnical assessment will be an ongoing requirement for accurate inputs and mining parameters into the mine plan and design.

●

Sill pillar recovery using longitudinal blind-up hole and transverse retreat LHOS has been successful with minimal dilution and ore loss. However, with increased extraction and shrinking pillars, potential exists for poorer sill pillar recovery.

●

Increased mining depth and pillar stress as the mine matures has the potential to increase mining costs. Geotechnical instrumentation and Map 3D modelling has been implemented to enable optimal ground support and allow selection of optimal extraction sequence to avoid increased mining stresses.

●

Stope dilution will become increasingly important as the orebody narrows with depth and will require greater geotechnical input to stope design and planning and execution of proper mining practices.

●

Maintaining the backfill quality, placement rate, and tight filling of stope voids will continue to have a significant impact on the mine’s ability to maintain the production rate, minimize dilution and maximize gold recovery.

22.5

 Conclusions and Recommendations

The LOM plan shows that Efemçukuru has a production life of almost 7 years and can sustain a production rate of 510,000 tpa until the end of 2025, based on the current mineral reserves.

The mine has a strong long-term potential for increased mine life through ongoing exploration activities. In this regard, the strategic focus will continue to be aimed at exploration and infill drilling particularly for Kokarpinar and Batı Vein resources.

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SECTION ● 23 Adjacent Properties

There are no properties adjacent to the Efemçukuru Project site, nor properties in the local region. The closest active operating gold mine is located in Ovacik of Izmir province some 100 km north of Efemçukuru. Eldorado also owns and operates the Kişladağ gold mine, located 200 km east of Efemçukuru through its wholly owned Turkish subsidiary, Tüprag Metal Madencilik Sanayi Ve Ticaret Limited Sirketi (Tüprag).

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SECTION ● 24 Other Relevant Data and Information

24.1

Life of asset strategy

The mine planning team at Efemçukuru generated an extended mine design which included in addition to the proven and probable material additional inferred material.

Applying the same mine design criteria as it is described in Section 15, the MSO software was used to generate additional stope shapes from the inferred resources.

Using the same assumed mining dilution, mining recovery, and cost structure to determine cutoff grades, over 750,000 tonnes at 4.20 g/t of potentially mineable inferred inventory has been identified.

The extended LOM design is provided in Figure 24-1.

Figure 24-1: Extended LOM Design

The strategic mine plan has the potential to extend the mine life by 17 months, based on current levels of production. Based on this strategic mine plan, Eldorado proposes to extend the existing decline to access the deeper levels of the present mining zones. This strategy would provide a 20% upside to the proven and probable mining inventory.

The extended LOM production schedule is provided in Table 24-1.

Additional potential exists in near mine exploration targets, like Kokarpinar and Bati, which could benefit from Efemçukuru’s existing infrastructure.

Table 24-1: Extended LOM Production Schedule

	Units	2020	2021	2022	2023	2024	2025	2026	2027	Total
Extended Production Profile	Kt	514	511	513	513	515	512	512	559	4,148
	Au g/t	6.78	6.78	6.22	5.94	5.89	6.45	5.42	4.09	5.92
	Au Koz	112	111	102	98	97	106,	89	73	790

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SECTION ● 25 Interpretation and Conclusions

Efemçukuru mine has had a solid history of operational performance since production commenced in 2012. The mine has achieved and surpassed many metrics described in the former NI 43-101 technical report (2007). Production tonnages and gold produced have matched and now exceed the plans outlined in the 2007 report. There is no reason to believe that the mine will not continue to perform as well in the future as it has during the past 8 years of operations.

The geology of the Efemçukuru deposit is well understood. The deposit is an example of an intermediate sulfidation epithermal vein deposit. Diamond drill holes continue to be the principal source of geologic and grade data for the Efemçukuru mine. That data is well managed and controlled by a robust QAQC program and database management system. These systems demonstrate that the Efemçukuru data are sufficiently accurate and precise for resource estimation.

The mineral resource estimate used industry-accepted methods and were classified as measured, indicated and inferred mineral resources using logic consistent with the CIM definitions referred to in NI 43-101. The current measured and Indicated mineral resources remain at healthy levels due to exploration and conversion success. The inferred mineral resources have increased nearly fivefold from the levels described in the 2007 report again illustrating continued exploration success over the years.

The mineral reserve estimate used industry-accepted methods and were classified as proven and probable mineral reserves using logic consistent with the CIM definitions referred to in NI 43-101. The cut-off grade was calculated from first principles and honor current and projected costs and mining factors. The current mineral reserves define almost 7 years of mine life, which is at least five (5) more than the life estimated in the 2007 report.

Annual resource model to mill reconciliation has shown consistent excellent agreement between predicted (resource), planned (reserves) and processed tonnage and grades, especially since 2016. This outcome demonstrates the steadfast high quality level of both mineral resource and reserve estimates.

The metallurgy and metal recovery values are well understood at Efemçukuru mine. The process plant has evolved to mill in excess of 510,000 tpa with minimal capital outlay. Currently the process plant produces a gold bearing bulk sulphide concentrate for sale to third party smelters. Continual improvement initiatives led to the planned implementation of column flotation to improve concentrate gold grade and overall quality while offering savings in less total concentrate tonnes being produced and transported.

The surface waste rock dump and dry stack tailings facilities continue to occupy quite small “footprints” while operating at high level of standard levels. Water is tightly managed and controlled with any return to the environment passing through a world-class water treatment facility.

The initial capital expenditure has been long paid back and the mine continues and will continue to generate positive cash flow for Eldorado.

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Efemçukuru Gold Mine, Turkey Technical Report

SECTION ● 26          Recommendations

26.1

Geology - Exploration

●

Efemçukuru mine should continue with aggressive efforts to assess ancillary vein systems for new gold mineralization. This should consist of both surface and underground based programs.

●

Near mine district options should also be reviewed.

26.2

Mining – planning and operational

●

Ongoing evaluation into narrow vein mining methods needs to continue as future ore become notably narrower.

●

Invest in more accurate long hole drilling equipment to minimize dilution and ore loss in ever increasing recovery of sill pillars as the mine matures.

●

Evaluate the use of technological advances, such as mechanical rock excavation units based on the continuous mining technology. The continuous mining technology allows for rapid advancement of development headings and selective extraction of irregular narrow vein orebodies without drilling and blasting.

26.3

Metallurgy and Processing

Bulk flotation of gold and sulfide minerals has been successfully operated since the process plant was commissioned in 2011. Bulk flotation testing will continue in the following areas.

●

New mineralized materials from NOS (north ore shoot) and Kestane Beleni North-West

●

Current plant throughput is about 30% over the original design. Impacts of coarser grind size and shorter flotation retention time on flotation gold recovery and concentrate quality need to be determined.

●

At present, rougher/scavenger concentrate is not reground. Effect of regrinding on the final concentrate quality will be investigated.

●

Dilution of mine waste rocks in the mill feed remains high due to mining in the narrow ore zones. Investigation will continue with respect to ore sorting and dense media separation to reject mine waste rocks before or after SAG mill grinding.

●

The amount of non-sulfidic gangues in the final concentrate is supposed to be below 20% when the concentrate is adequately cleaned. Such a benchmark was not always achieved during operation. Larger amount of non-sulfidic gangues in the concentrate means lower gold grade and lower sulfur content, both of which affect unfavourably concentrate marketing and values. Column flotation testing has demonstrated that the amount of non-sulfidic gangues in the concentrate could be reduced significantly. Implementation of column flotation will start in the third quarter of 2020.

●

The upcoming column flotation project will increase gold grade and sulfur content in the concentrate. However, contents of deleterious elements will increase. Continued metallurgical testing is recommended to address to determine whether it is feasible to reject a portion of these elements out of the bulk flotation concentrate.

●

Research into the viability of selective sequential flotation to help minimize deleterious element content should also be ongoing. If successful, pre-concentration of some of these elements, like the base metals (lead and zinc) may not only minimize potential penalty charges but create a saleable product.

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SECTION ● 27          References

Baker, T., 2019, Gold±Copper Endowment and Deposit Diversity in the West Tethyan Magmatic Belt, Southeast Europe: Implications for Exploration: Economic Geology, In Press.

Boucher, K., 2016, The structural and fluid evolution of the Efemçukuru epithermal gold deposit, western Turkey: M.Sc thesis, Vancouver, BC, Canada, The University of British Columbia, 466p.

Dilek, Y. and Altunkaynak, Ş., 2009, Geochemical and temporal evolution of Cenozoic magmatism in western Turkey: mantle response to collision, slab break-off, and lithospheric tearing in an orogenic belt: Geological Society, London, Special Publications, v. 311, p. 213-233.

Ersoy, Y.E., Helvacı, C. and Sözbilir, H., 2010, Tectono-stratigraphic evolution of the NE–SW trending superimposed Selendi basin: Implications for late Cenozoic crustal extension in Western Anatolia, Turkey: Tectonophysics, v. 488, p. 210-232.

Gessner, K., Markwitz, V. and Güngör, T., 2017, Crustal fluid flow in hot continental extension: tectonic framework of geothermal areas and mineral deposits in western Anatolia: Geological Society, London, Special Publications, v. 453, p. 1-23.

Menant, A., Jolivet, L., Tuduri, J., Loiselet, C., Bertrand, G. and Guillou-Frottier, L., 2018, 3D subduction dynamics: A first-order parameter of the transition from copper-to gold-rich deposits in the eastern Mediterranean region: Ore Geology Reviews, v. 94, p. 118-135.

Okay, A.İ., Zattin, M. and Cavazza, W., 2010, Apatite fission-track data for the Miocene Arabia Eurasia collision: Geology, v. 38, p. 35-38.

Sillitoe, R.H., and Hedenquist, J.W., 2003, Linkages between volcanotectonic settings, ore-fluid compositions, and epithermal precious metal deposits: Society of Economic Geologists, Special Publication 10, p. 315–343.

Simmons, S.F., White, N.C., and John, D.A., 2005, Geological characteristics of epithermal precious and base metal deposits: Economic Geology 100th Anniversary Volume, p. 485–522.

van Hinsbergen, D.J., 2010, A key extensional metamorphic complex reviewed and restored: the Menderes Massif of western Turkey: Earth-Science Reviews, v. 102, p. 60-76.

Wardrop Engineering Incorporated, 2007, Technical Report on the Efemçukuru Project, NI 43-101 Technical Report, August 2007

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SECTION ● 28 Date and Signature Page

Date and Signature Page

The effective date of this report entitled “Technical Report, Efemçukuru Gold Mine, Turkey” is December 31, 2020. It has been prepared for Eldorado Gold Corporation by David Sutherland, P. Eng., Imola Götz, P.Eng, Paul Skayman, FAusIMM, Ertan Uludag, P.Geo. 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” Imola Götz Imola Götz, M.Sc., P. Eng.
“Signed” Paul J.Skayman Paul J. Skayman, FAusIMM	“Signed and Sealed” Ertan Uludag Ertan Uludag P. Geo.
“Signed and Sealed” Sean McKinley Sean McKinley, P. Geo.

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., Vancouver in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019.

I am a member of the Engineers & Geoscientists of British Columbia. I graduated from the Lakehead University with aBachelor 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 Efemçukuru Gold Mine on numerous occasions with my most recent visit occurring on May 18, 2018 (1 day).

I was responsible for coordinating the preparation of the technical report. I am responsible for the preparation or supervising the preparation of items 1, 2, 3, 4, 5, 6, 18, 20, 24, 25, 26, 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.

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, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019, 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”

David Sutherland

David Sutherland, P. Eng.

CERTIFICATE OF QUALIFIED PERSON

Imola Götz, M. Sc., P. Eng.

1188 Bentall 5, 550 Burrard St.

Vancouver, BC

Tel: (604) 601-6658

Fax: (604) 687-4026

Email: imola.gotz@eldoradogold.com

I, Imola Götz, am a Professional Engineer, employed as Manager, Mine Engineering (Underground), of Eldorado Gold Corporation and reside at 1127 St Andrews Avenue in North Vancouver, in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019.

I am a member of the Engineers & Geoscientists of British Columbia. I graduated from the Technical University of Petrosani in Romania with a Master of Science in Mine Engineering in 1986, and subsequently obtained a Bachelor of Education from Lakehead University in 1993.

I have practiced my profession continuously since 1996 and I have been involved in mine planning, mine design, mineral resource and mineral reserve estimation, cut-off grade optimization in underground and open pit gold properties in Canada, Romania, Greece, Turkey, Mexico, Guatemala, Chile and Argentina.

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

I have visited the Efemçukuru Gold Mine on numerous occasions with my most recent visit occurring on January 14 to 18, 2019.

I was responsible for the mineral reserves and the preparation of related sections on mineral reserves calculation, mining methods and sections related to costs of the technical report. I am responsible for the preparation or supervising the preparation of items 15, 16, 21 and 22 in the technical report.

I have not had 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, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019, 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”

Imola Götz

Imola Götz, M.Sc., P. Eng.

CERTIFICATE OF QUALIFIED PERSON

Paul 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 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, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019.

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 Efemçukuru Gold Mine on numerous occasions with my most recent visit occurring on November 18, 2019 (1 day).

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.

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, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019, 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”

Paul J. Skayman

Paul Skayman, FAusIMM

CERTIFICATE OF QUALIFIED PERSON

Ertan Uludag, P.Geo.

1188 Bentall 5, 550 Burrard St.

Vancouver, BC

Tel: (604) 601-6658

Fax: (604) 687-4026

Email: ertan.uludag@eldoradogold.com

I, Ertan Uludag, am a Professional Geoscientist, employed as Resource Geologist, of Eldorado Gold Corporation and reside at 6779 Kitchener Street in the City of Burnaby, in the Province of British Columbia.

This certificate applies to the technical report entitled Technical Report, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019.

I am a member of the Engineers & Geoscientists British Columbia (formerly the Association of Professional Engineers and Geoscientists of British Columbia). I graduated from Middle East Technical University in Ankara Turkey with Bachelor of Science in Geological Engineering in July 1994.

I have practiced my profession continuously since 1996. I have been involved in ore control, mine geology and resource modelling work on gold, copper, zinc, lead and silver underground and open pit properties in Turkey, China, Greece, Canada and Romania, and rock mechanics in 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 Efemçukuru Gold Mine on numerous occasions with my most recent visit occurring on June 10 to 11, 2019.

I was responsible for the mineral resources and the preparation of related sections on sample preparation and analyses, data verification and mineral resource estimation for this technical report. I am responsible for the preparation or supervising the preparation of items 11, 12 and 14 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, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019, 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”

Ertan Uludag

Ertan Uludag, P.Geo.

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, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019.

I am a member of the Engineers & Geoscientists British Columbia (formerly the Association of Professional Engineers and Geoscientists of 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 Efemçukuru Gold Mine on numerous occasions with my most recent visit occurring on February 6 to 7, 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, Efemçukuru Gold Mine, Turkey, with an effective date of December 31, 2019, 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”

Sean McKinley

Sean McKinley, P.Geo.