KLONDEX MINES LTD - FORM 8-K - EX-99.1 - EXHIBIT 99.1 - TECHNICAL REPORT FOR THE TRUE NORTH MINE, BISSETT, MANITOBA, CANA

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Date and Signature Page

The undersigned prepared this Technical Report (Technical Report) report, titled: Technical Report for the True North Mine, Bissett, Manitoba, Canada, dated the 12th day of May 2017, with an effective date of March 31, 2017, in support of the public disclosure of Mineral Resource and Mineral Reserve estimates for the True North Project. The format and content of the Technical Report have been prepared in accordance with Form 43-101F1 of National Instrument 43-101 – Standards of Disclosure for Mineral Projects of the Canadian Securities Administrators.

Dated this 12th day of May 2017

Signed “Sarah Bull”	No. 22797, Nevada
Sarah Bull, P.E.	(Sealed)
Practical Mining LLC
495 Idaho Street, Suite 205
Elko, Nevada 89815, USA
775-345-3718 ext. 502
Email: sarahbull@practicalmining.com

Signed “Mark Odell”	No. 13708, Nevada
Mark Odell, P.E.	(Sealed)
Practical Mining LLC
495 Idaho Street, Suite 205
Elko, Nevada 89815, USA
775-345-3718 ext. 101
Email: markodell@practicalmining.com

Signed “William Stone”

Dr. William E. Stone, P. Geo.

P&E Mining Consultants

201 County Court Blvd., Suite 401

Brampton, ON, L6W 4L2

(905) 595-0575

bill@peconsulting.com

Signed “Fred Brown”

Fred H. Brown, P. Geo.

P&E Mining Consultants

201 County Court Blvd., Suite 401

Brampton, ON, L6W 4L2

(905) 595-0575

billfred@peconsulting.com

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Signed “Alfred Hayden”

Alfred S. Hayden, P. Eng.

EHA Engineering Ltd.

Consulting and Metallurgical Engineers

Box 27111, Postal Stn. B.

Richmond Hill, Ontario, L4E 1A7

Signed “David Orava”

David A. Orava, P. Eng.

P&E Mining Consultants

201 County Court Blvd., Suite 401

Brampton, ON, L6W 4L2

(905) 595-0575

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Table of Contents

Date and Signature Page			iii
List of Tables			xi
List of Figures			xiii
List of Abbreviations			xv
1.	Summary		18
	1.1.	Property Description	18
	1.2.	Tailings Reprocessing	18
	1.3.	Environmental and permitting	19
	1.4.	Geology	19
	1.5.	History	19
	1.6.	Mineral Resource Estimate	21
	1.7.	Mineral Reserve Estimate	23
	1.8.	Cash Flow Analysis and Economics	24
	1.9.	Conclusions	25
	1.10.	Recommendations	25
2.	Introduction		27
	2.1.	Terms of Reference and Purpose of this Technical Report	27
	2.2.	Qualification of the Authors	27
	2.3.	Sources of Information	27
	2.4.	Units of Measure	28
3.	Reliance on Other Experts		30
4.	Property Description and Location		31
	4.1.	Property Location	31
	4.2.	Property Description	33
	4.3.	Liabilities and Permits	33
5.	Accessibility, Climate, Vegetation, Physiography, Local Resources and Infrastructure		35
	5.1.	Access to Project	35
	5.2.	Climate	35

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	5.3.	Vegetation		35
	5.4.	Physiography		35
	5.5.	Local Resources and Infrastructure		35
6.	History			39
	6.1.	Events Prior to 1989		39
	6.2.	Events since 1989		40
7.	Geological Setting and Mineralization			49
	7.1.	Regional Geology		49
	7.2.	Local Geology		51
		7.2.1.	Host Rock Units	51
		7.2.2.	Structure	53
		7.2.3.	Veins	54
		7.2.4.	Alteration	57
8.	Deposit Types			60
9.	Exploration			61
10.	Drilling and Sampling Methodology			62
	10.1.	Diamond Core Drilling		62
	10.2.	Channel Chip Sampling		64
	10.3.	Historic Tailings Pond Drilling/Sampling		65
11.	Sample Preparation, Analysis, and Security			66
	11.1.	Core Sampling Methods		66
		11.1.1.	Surface Core Sampling Methods	66
		11.1.1.	Underground Core Sampling Methods	67
	11.2.	Face Sampling Methods		69
	11.3.	Tailings Sampling Methods		69
	11.4.	Sample Quality, Representativeness and Sample Bias		69
	11.5.	Sample Preparation, Analysis and Security		70
		11.5.1.	Core Sample Preparation and Analysis	70
		11.5.2.	Channel Chip Sample Preparation and Analysis	70
		11.5.3.	Tailings Sample Preparation and Analysis	71

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	11.6.	Core Quality Assurance and Quality Control		71
		11.6.1.	Sample Standards	71
		11.6.2.	Core Sample Blanks and Duplicates	74
	11.7.	Chip Quality Assurance and Quality Control		75
		11.7.1.	Sample Standards	75
		11.7.2.	Chip Sample Duplicates	76
	11.8.	Tailings Quality Assurance-Quality Control		77
		11.8.1.	Sample Standards	77
		11.8.2.	Tailings Sample Blanks and Duplicates	77
	11.9.	Recommendations and Conclusions		79
12.	Data Verification			81
	12.1.	Drill Data Review		81
		12.1.1.	Collar Location Checks	81
		12.1.2.	Hole Survey Checks	81
		12.1.3.	Core Assay Checks	81
		12.1.4.	Geology Checks	82
	12.2.	Channel Chip Data Review		82
		12.2.1.	Collar Location Checks	82
		12.2.2.	Downhole Survey Check	82
		12.2.3.	Assay Check	82
	12.3.	Tailings Data Review		83
		12.3.1.	Assay Checks	83
	12.4.	Due Diligence Sampling		83
	12.5.	Conclusions to Data Verification		87
13.	Mineral Processing and Metallurgical Testing			88
14.	Mineral Resource Estimates			93
	14.1.	Introduction		93
	14.2.	Previous Resource Estimates		94
	14.3.	Data Supplied		95
	14.4.	Bulk Density		95

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	14.5.	Vein Modelling		96
	14.6.	Assay Data		98
	14.7.	Compositing		101
	14.8.	Treatment of Extreme Values		104
	14.9.	Au Grade Estimation, Classification and Minimum Width		105
	14.10.	Mineral Resource Estimate		106
	14.11.	Block Model Validation		108
15.	Mineral Reserve Estimates			111
	15.1.	Methodology		111
		15.1.1.	Underground Reserves	111
		15.1.2.	Tailings Reserves	114
	15.2.	Statement of Reserves		115
16.	Mining Methods			117
		16.1.1.	Access Development	117
		16.1.2.	Geotechnical	117
		16.1.3.	Ground Support	117
		16.1.4.	Ventilation and Secondary Egress	120
	16.2.	Power Distribution and Dewatering		120
	16.3.	Mining Methods		122
		16.3.1.	Longhole Stoping	122
		16.3.2.	Captive Sub Level Longhole Stoping	125
		16.3.3.	Haulage	126
		16.3.4.	Backfill	126
	16.4.	Equipment Fleet Underground		128
	16.5.	Tailings Reprocessing		129
17.	Recovery Methods			131
18.	Project Infrastructure			134
	18.1.	Location and Access		134
	18.2.	Accommodations and Camp Facilities		135
	18.3.	Electrical Power and On-Site Distribution		135

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	18.4.	Water Supply and Reticulation		135
	18.5.	Air Compressors		135
	18.6.	Diesel Fuel and On-Site Storage Facility		135
	18.7.	Warehousing and Material Handling		135
	18.8.	Site Security		136
	18.9.	Communication		136
	18.10.	On-Site Transport and Infrastructure		136
	18.11.	Solid Waste Disposal		136
	18.12.	Parts and Mine Supply Freight		136
	18.13.	Mobile and Fixed Equipment Maintenance Facility		136
	18.14.	First Aid and Ambulance		137
	18.15.	Office and Administration Buildings		137
	18.16.	Tailings Storage		137
	18.17.	Stockpiles		138
19.	Market Studies and Contracts			139
	19.1.	Precious Metal Markets		139
	19.2.	Contracts		139
20.	Environmental Studies, Permitting and Social or Community Impact			140
	20.1.	Summary		140
	20.2.	Scope of the Project		140
	20.3.	Ongoing Exploration and Project Development		141
	20.4.	Information Review and Assessment		142
		20.4.1.	Documentation Reviewed	142
		20.4.2.	Licenses, Permits and Approvals	143
		20.4.3.	Revised Environmental License and Minor Alterations	144
		20.4.4.	Current Status / Mitigative Measures	147
		20.4.5.	Community Engagement	149
		20.4.6.	Mine Closure	150
21.	Capital and Operating Costs			152
	21.1.	Capital Costs		152

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	21.2.	Underground Operating Costs and Cut-off Grade		152
	21.3.	Cut-off Grade Calculations		155
		21.3.1.	Underground Mining Cut-off Grade	155
		21.3.2.	Tailings Reprocessing Cut-off Grade	156
22.	Economic Analysis			158
	22.1.	Life of Mine Plan and Economics		158
	22.2.	Sensitivity Analysis		160
23.	Adjacent Properties			163
24.	Other Relevant Data and Information			165
25.	Interpretation and Conclusions			166
26.	Recommendations			167
27.	References			169
28.	Glossary			170
29.	Appendix A True North Claims Information			177
30.	Appendix B: Certification of Authors and Consent Forms			184

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List of Tables

Table 1-1 Chronology of Ownership of the True Project	19
Table 1-2 Mineral Resource Statement as of March 31, 2017	22
Table 1-3 Tailings Mineral Resource as of March 31, 2017	22
Table 1-4 True North Mineral Reserves as of March 31, 2016	23
Table 1-5 Key Operating and After Tax Financial Statistics	24
Table 2-1 Qualified Professionals	27
Table 2-2 Units of Measure	29
Table 4-1 Summary of Klondex Mineral Property Holdings and Surface Areas	33
Table 6-1 Historic Production at Rice Lake Mine: 1927-1968	45
Table 6-2 Historic Production at Rice Lake Mine: 1980-2001	46
Table 6-3 Historic Production at Rice Lake Mine: 2007-2015	47
Table 10-1 Summary of Surface Exploration on the True North Mine	63
Table 10-2 Summary of Tailings Drilling at True North Mine	65
Table 11-1 Certified Gold Assay Values for Commercial Standards	72
Table 11-2 Certified Gold Assay Values for Commercial Standards	76
Table 13-1 Harmony Gold – Rice Lake Deposit Metallurgical Results	88
Table 13-2 Hinge Zone Metallurgical results	89
Table 13-3 007 Zone Metallurgical Results	89
Table 13-4 SGS Lakefield And Starkey Associates Sag Mill Testing Results	89
Table 13-5 JKTech Drop-Weight Test Summary	89
Table 13-6 More SGS Lakefield And Starkey & Associates Sag Mill Testing Results	90
Table 13-7 Results Leaching Flotation Tails for 24 Hours at 2.5 gpl NaCN Concentration	90
Table 13-8 Results Leaching Flotation Tails for 24 Hours at 0.5 gpl NaCN Concentration	91
Table 13-9 Results from Leaching Samples from Tailings Storage Facility	92
Table 14-1 Total reported mineral resources as of June 30, 2016.	94
Table 14-2 True North database records.	95
Table 14-3 Bulk Density Sample Statistics	95
Table 14-4 Modelled Veins	97
Table 14-5 Summary Drillhole Assay Statistics	98
Table 14-6 Summary channel sample assay statistics	100
Table 14-7 Summary Composite Statistics by Vein	102
Table 14-8 Composite Capping Levels	104
Table 14-9 Total Mineral Resources^1,2,3,4,5	106
Table 14-10	106
Table 14-11 Block Model Validation Grades	108
Table 14-12 Volume Comparison	109

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Table 15-1	Mineral Reserves Cut Off Grade Calculation	111
Table 15-2	Tailings Reprocessing Cut-off Grade	114
Table 15-3	True North Mineral Reserves as of March 31, 2017	115
Table 16-1	Waste Rock Backfill and Stope Voids	127
Table 16-2	Annual Production and Development Plan	127
Table 16-3	Underground Mobile Equipment Fleet	128
Table 16-4	Tailings Reprocessing Schedule	129
Table 20-1	Obtained Licenses and Key Permits and Approvals	143
Table 20-2	Potential Significant Environmental Impacts and Current Status / Mitigative Measures	147
Table 21-1	Capital Costs	152
Table 21-2	Underground Development Unit Costs	152
Table 21-3	Direct Underground Mining Cost	153
Table 21-4	Indirect Underground Mining Cost	153
Table 21-5	Processing Cost	153
Table 21-6	General and Administrative Cost	153
Table 21-7	Tailings Reprocessing Cost	154
Table 21-8	Mine Cut-off Grade Calculation	155
Table 21-9	Tailings Reprocessing Cut-off Grade Calculation	156
Table 22-1	Income Statement 2015 – 2018 ($000’s)	158
Table 22-2	Cash Flow Statement 2015 – 2019 ($000’s)	159
Table 22-3	Key Operating and After Tax Financial Statistics	159

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List of Figures

Figure 4-1	Location of the True North Mine, Bissett, Manitoba	32
Figure 4-2	Regional Klondex Mining Claim and Lease Holdings	34
Figure 5-1	Photograph of the True North Gold Mine Looking South	36
Figure 5-2	Plan View of Surface Infrastructure at the True North Gold Mine	37
Figure 5-3	Tailings Management Area at the True North Gold Mine	38
Figure 7-1	Regional Geologic Map showing the Location of True North Gold Mine in the Archean Uchi Subprovince, Manitoba	50
Figure 7-2	Geologic Map showing the Location of Gold Deposits and Lithotectonic Assemblages in True North Gold Mine Area	50
Figure 7-3	Local Geology of True North Gold Mine Area	52
Figure 7-4	The Structural Geological Setting of Gold Mineralization at True North Gold Mine	53
Figure 7-5	Shear Zones and Quartz Veins	55
Figure 7-6	Example of 16-Type Shear and 38-Type Breccia Gold Mineralized Quartz Veins in the SAM Unit at True North	56
Figure 7-7	Controls on Gold Mineralization in the 007 Zone	58
Figure 7-8	Interpreted Cross-Sections of 007 and L10 Zones Looking West	59
Figure 7-9	Level Plan of the 710 Zone showing the Location of the 710 and 711 Veins	59
Figure 8-1	Schematic Cross-Section Representation of the Geometry and Structural Setting of Shear Zone Hosted Gold-Bearing Quartz Vein Networks in Greenstone Belt Terrains like True North Gold Mine	60
Figure 11-1	Flow Chart for Surface and Underground Core Sampling Methods	68
Figure 11-2	Assay Results of Standard CDN-GS-1P5C	72
Figure 11-3	Assay Results for Standard CDN-GS-6B	73
Figure 11-4	Assay Results for Standard CDN-GS-13A	73
Figure 11-5	Assay Results for Standard CDN-GS-22	74
Figure 11-6	Assay Results for Blanks	75
Figure 11-7	Assay Results for Duplicates	75
Figure 11-8	Assay Results for Chip Sample Duplicates	77
Figure 11-9	Tailings Assay Results for Standard GS-1P5C	78
Figure 11-10	Tailings Assay Results for Standard GS-1L	78
Figure 11-11	Tailings Assay Results for Blanks	79
Figure 11-12	Tailings Assay Results for Duplicates	79
Figure 12-1	Due Diligence Sample Pulp Results for Gold	84
Figure 12-2	Due Diligence Sample Core Results for Gold	85
Figure 12-3	Due Diligence Core Reject Sample Assays	86
Figure 12-4	Due Diligence Tailings Sample Results for Gold	87
Figure 15-1	Grouping of Diluted Stope Optimizer Rings to Create Stopes	112

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Figure 15-2	26L – 710 Complex Final Reserves Plan by Year Mined	112
Figure 15-3	16L 8/10 - Reserves Mined by Year	113
Figure 15-4	26L V91 (SAM) - Reserves Mined by Year	113
Figure 15-5	Cohiba - Reserves Mined by Year	114
Figure 16-1	Longitudinal Section	119
Figure 16-2	True North Gold Mine 710 Complex - Ventilation System	121
Figure 16-3	Longhole Open Stope Sill Development	122
Figure 16-4	Longhole Open Stope Raise and Drilling	123
Figure 16-5	Longhole Open Stope Blasting	123
Figure 16-6	Longhole Open Stope Backfilling	124
Figure 16-7	Overcut and Undercut Plan View	124
Figure 16-8	True North Gold Mine Typical Longhole Drill Section	125
Figure 16-9	True North Gold Mine Sub-Level Captive Longhole Stope	126
Figure 16-10	Tailings Recovery Flow Diagram	130
Figure 16-11	Aerial View of Tailings Recovery Site	130
Figure 17-1	True North Gold Mine Process Plant Flow Sheet	133
Figure 18-1	Surface Infrastructure Plan View	134
Figure 19-1	Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average ($US/oz)	139
Figure 20-1	True North Gold Mine Site Plan	145
Figure 20-2	TMA Site Plan	145
Figure 20-3	Aerial View of the Mine, Plant Site and TMA	146
Figure 21-1	True North Total Cost Sensitivity to Production Rate	154
Figure 21-2	Cut-off Grade Sensitivity to Gold Price	156
Figure 21-3	Tailings Reprocessing Cut-off Sensitivity to Gold Price	157
Figure 22-1	5% NPV Sensitivity	161
Figure 22-2	10% NPV Sensitivity	161
Figure 22-3	Internal Rate of Return Sensitivity	162
Figure 22-4	Internal Rate of Return Sensitivity	162
Figure 23-1	Adjacent Properties to the True North Gold Mine	164

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List of Abbreviations

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A	Ampere	kA	kiloamperes
AA	atomic absorption	kCFM	thousand cubic feet per minute
A/m²	amperes per square meter	Kg	Kilograms
AGP	Acid Generation Potential	km	kilometer
Ag	Silver	km2	square kilometer
ANFO	ammonium nitrate fuel oil	kWh/t	kilowatt-hour per ton
ANP	Acid Neutralization Potential	LOI	Loss On Ignition
Au	Gold	LoM	Life-of-Mine
AuEq	gold equivalent	m	meter
btu	British Thermal Unit	m²	square meter
°C	degrees Celsius	m³	cubic meter
CCD	counter-current decantation	masl	meters above sea level
CIL	carbon-in-leach	mg/L	milligrams/liter
CoG	cut-off grade	mm	millimeter
cm	centimeter	mm²	square millimeter
cm²	square centimeter	mm³	cubic millimeter
cm³	cubic centimeter	MME	Mine & Mill Engineering
cfm	cubic feet per minute	Moz	million troy ounces
ConfC	confidence code	Mt	million tonnes
CRec	core recovery	MTW	measured true width
CSS	closed-side setting	MW	million watts
CTW	calculated true width	m.y.	million years
°	degree (degrees)	NGO	non-governmental organization
dia.	diameter	NI 43-101	Canadian National Instrument 43-101
EIS	Environmental Impact Statement	oz	Troy Ounce
EMP	Environmental Management Plan	opt	Troy Ounce per short ton
FA	fire assay	%	percent
Ft	Foot	PLC	Programmable Logic Controller
Ft²	Square foot	PLS	Pregnant Leach Solution
Ft³	Cubic foot	PMF	probable maximum flood
g	Gram	POO	Plan of Operations
g/L	gram per liter	ppb	parts per billion
g-mol	gram-mole	ppm	parts per million
g/t	grams per tonne	QAQC	Quality Assurance/Quality Control
ha	hectares	RC	reverse circulation drilling
HDPE	Height Density Polyethylene	ROM	Run-of-Mine
HTW	horizontal true width	RQD	Rock Quality Description
ICP	induced couple plasma	SEC	U.S. Securities & Exchange Commission
ID2	inverse-distance squared	Sec	second
ID3	inverse-distance cubed	SG	specific gravity
ILS	Intermediate Leach Solution	SPT	Standard penetration test

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1.	Summary

Practical Mining LLC (PM) and P&E Mining Consultants Inc. (P&E) were engaged by Klondex Mines Ltd. (KGS or Klondex or the Company), to prepare an updated Technical Report (TR) in accordance with National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators. PM’s and P&E’s evaluation of the True North Mine (True North Project, True North or the Project), located in Bissett, Manitoba, Canada, is presented herein. This TR, dated the 12th day of May 2017, with an effective date of March 31, 2017, updates the previous mineral resource and mineral reserve estimates effective June 30, 2016 (Puritch et al. 2016).

1.1.	Property Description

True North is located in southeast Manitoba, Canada at the edge of Bissett township on the north shore of Rice Lake. It lies approximately 100 miles (162 kilometers) northeast of Winnipeg, roughly 150 driving miles via all-weather Provincial highways. The town of Bissett is a long established mining community with a fluctuating population which is currently approximately 340 people.

The mine accesses quartz vein gold mineralization using shafts and underground mining methods. The Project has an on-site processing facility with ore processing circuit comprises crushing, grinding and gravity concentration and flotation, followed by Carbon-in-Pulp (CIP) processing on the gravity circuit. Refurbishing existing underground openings as well as test stope mining has been completed, as of the effective date of this Technical Report. Process plant testing programs are completed and the plant is operational and a historic tailings re-processing assessment project was completed in September 2016. The land position consists of 300 mining claims, 18 mining patents and a mineral lease totaling 99,477 acres (40,257 ha).

1.2.	Tailings Reprocessing

Under the previous ownership a percentage of the coarse grained free gold was lost to tailings. Trial processing of the reslurried tailings has demonstrated that reprocessing of the existing historic tailings can recover approximately 89% of the contained gold.

PM and P&E have produced Mineral Reserve and Mineral Resource estimates on the tailings and the Company plans to supplement the mine feed with historic tailings. The tailings reprocessing operation will be carried out concurrently with the underground mining operation until 2019 and will continue for an additional 9 years until 2028 on a stand-alone basis.

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1.3.	Environmental and Permitting

Klondex has revised the existing Environmental Act Licence 2628RR (Licence) for True North which includes approvals for minor alternations needed for operation. The San Gold Mine Closure Plan (2012) (Closure Plan) and the pledged fixed-asset financial security for mine closure were transferred to Klondex in January 2016.

Klondex collects all required environmental monitoring data including: water quality sampling, environmental effects monitoring, final effluent release reporting, and is developing procedures for its environmental management system. Klondex is also in the process of re-initiating First Nations and Aboriginal community engagement.

Based on the available information, P&E is of the opinion that there does not appear to be any significant environmental and/or social barriers to the True North operation.

1.4.	Geology

All the major gold occurrences in the True North area occur as quartz veins or quartz vein systems formed during structural deformation of the host rocks. At the Project, gold mineralization is controlled by quartz-carbonate veins and vein systems in brittle-ductile structures with related hydrothermal alteration halos within or at the margin of particular host rock units.

All of the gold mineralized zones at the Project are hosted in rocks of the Bidou Lake Assemblage which forms a north-facing stratigraphic sequence of tholeiitic basalt to intermediate volcanic flows, dacite crystal tuffs and breccias overlain by well stratified felsic epiclastic rock interpreted to be of pyroclastic and sedimentary origin. The stratigraphic sequence is intruded by tholeiitic gabbro sills and dykes and felsic porphyry dykes.

The best-known gabbro sill is the San Antonio Unit, host rock of the gold mineralization at the True North deposit. The Bidou Lake Assemblage is unconformably overlain by feldspathic sandstone of the San Antonio Assemblage.

The gold mineralized veins show a high degree of structural control and are best developed in competent mafic host rock ranging from intermediate to gabbroic in composition.

1.5.	History

Table 1-1 Chronology of Ownership and Major Events of the True North Project

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Dates	Company	Details
1911		Initial discovery at shore of Rice Lake
1927-1931	Mining syndicate	Exploration shafts and lateral test mining
1913-1968	San Antonio Gold Mines Ltd.	New company established, power lines and process facilities built
1932-1968	San Antonio Gold Mines Ltd.	Continuous production beginning at 150 tpd and increasing to 500 tpd using shrinkage stoping.
July 1968	San Antonio Gold Mines Ltd.	No. 1 Shaft surface hoist destroyed by fire, production ceased, San Antonio declared bankruptcy.
1968	New Forty Four Mines	Acquired assets
1980	New Forty Four Mines	Process plant destroyed by fire
1980-1983	Brinco Mining Limited JV	Brinco Mining underground exploration drilling, mined 100,000 tons, earned 100% interest but completed no further work
1987-1988	Agreement with Subsidiary of Inco Ltd.	Agreement with Brinco, drilled over 20,000 ft, then opted out.
1989-1994	Rea Gold Corp.	Acquired property, carried out engineering studies
1994-1997	Rea Gold Corp.	Underground rehab and exploration, feasibility studies, shaft extension, new process facility, then placed into receivership.
1998-2001	Harmony Gold (Canada) Inc.	Purchased project, installed ramp system in lower D-shaft area and attempted long hole mining. Grade was too low, project put on care and maintenance in August 2001.
2002	Option agreement with Wildcat Exploration Ltd.	Engineering studies based on shrinkage stope mining methods yielded positive results but Wildcat was unable to complete the acquisition.
2004	Rice Lake Joint Venture Inc.	San Gold Resources Corporation (Old San Gold) and Gold City Industries Ltd. JV to acquire Harmony through RLJV.
June 30, 2005	San Gold Corporation	Old San Gold and Gold City amalgamated, formed San Gold Corporation
2005-May 2015	San Gold Corporation	Exploration drilling, LiDAR survey, drove ramps, produced from three mineral trends, modernized process plant. Placed on care and maintenance.
June 2015	San Gold Corporation	Declared bankruptcy
2016	Klondex Mines Ltd.	Acquired 100% of the Rice Lake Mine, process plant complex and 400 km²exploration land package. Began underground rehab and started test mining. Commenced tailings reprocessing analysis. Prepared to process stockpiles. Changed mine name to True North. Sept. 2016 announced formal decision to resume production.

Gold was originally discovered on the Project in 1911, however, it was not until the 1920s that the construction of a shaft to a depth of 725 feet (221 metres) and approximately 2,000 feet (600 metres) of underground lateral development confirmed the presence of an economically viable mineral resource.

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Small scale production from underground mining commenced in 1932 and production increased to about 500 tons per day (450 tonnes per day) in 1948. A fire destroyed some of the surface facilities in 1968, and as a result production was suspended. Beginning in the late 1990s, production was intermittent under various ownerships, until 2016 when Klondex acquired a 100% interest in True North.

1.6.	Mineral Resource Estimate

The mineral resource estimate is based on data from 7,877 surface and underground drill holes, completed through January 5, 2017. This estimate also includes 29,214 channel samples from underground drifting.

Wire frame models were constructed for 62 vein sets. The vein models were constructed by digitally contouring surfaces along planes of data points defined by drill hole intercepts and underground surveys. Each data point is identified as a particular side of a particular vein (hanging wall or footwall), and software is used to contour surfaces between corresponding points. Hanging wall and footwall surfaces are then combined to form a solid wire frame. Assay values were composited into 10-foot lengths and truncated at the vein hanging wall and footwall. Only composites flagged as representing vein material were used in the grade estimation. A grade capping scheme based on resource category and vein was employed. Grades were assigned to individual blocks using the Inverse Distance Cubed method (ID3).

Vein models were developed by Klondex based on a scripted grid modelling workflow using Maptek Vulcan software. Grid modelling is applicable to modelling narrow, continuous geological features such as precious metal veins and coal seams and creates a surface by interpolating a regular grid of points over a modelling area. These grid points are combined with the input intercepts to create output triangulated surfaces that represent the vein hanging wall and footwall contacts. The contacts are combined to create a valid solid triangulation for use in building the resource block model.

The 62 veins were each assigned a specific search orientation based on their respective approximate dip and dip direction. Measured blocks require a minimum of four composite samples within an average anisotropic search radius of 50 feet. Indicated blocks required three drill hole intercepts within 100 feet. Inferred blocks required two drill intercepts within 500 feet. Grades were estimated only for blocks contained within the modeled veins.

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Mineral resources were estimated only for blocks within the modeled vein wireframes. In all cases, the vein boundary was treated as a “hard” boundary.

Table 1-2 Mineral Resource Statement as of March 31, 2017

Class	Grade Au opt	Grade Au g/t	Tons	Au oz
Measured	0.220	7.54	521,000	115,000
Indicated	0.214	7.34	1,276,000	273,000
Meas + Ind	0.216	7.40	1,797,000	388,000
Inferred	0.182	6.24	3,676,000	668,000

	1.	Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, marketing, or other relevant issues.
	2.	Mineral resources were estimated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.
	3.	The quantity and grade of reported Inferred resources in this estimation are uncertain in nature and there has been insufficient exploration to define these inferred resources as an Indicated or Measured mineral resource and it is uncertain if further exploration will result in upgrading them to an Indicated or Measured mineral resource category.
	4.	Contained metal may differ due to rounding.
	5.	Cut-off grade = 0.090 opt Au.

Table 1-3 Tailings Mineral Resource as of March 31, 2017

Class	Grade Au opt	Grade Au g/t	Tons (k)	Au (k) oz
Indicated	0.024	0.82	2,138	51.0
Inferred	0.022	0.75	47	1.1

	1.	Tailings Mineral Resource is inclusive of tailings Mineral Reserve.
	2.	Mineral Resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, marketing, or other relevant issues.
	3.	Mineral Resource was estimated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.
	4.	The quantity and grade of reported Inferred Mineral Resources in this estimation are uncertain in nature and there has been insufficient exploration to define these Inferred Mineral Resources as an Indicated or Measured Mineral Resource and it is uncertain if further exploration will result in upgrading them to an Indicated or Measured Mineral Resource category.
	5.	Contained metal may differ due to rounding.
	6.	Cut-off grade = 0.015 opt Au (0.51 g/t Au).
	7.	A dry bulk density of 0.044 tons per cubic foot (pcf) was utilized in the tailings Mineral Resource estimate tonnage calculation.

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1.7.	Mineral Reserve Estimate

Excavation designs for stopes, stope development drifting, and access development were created using Vulcan software. Stope designs were aided by the Vulcan Stope Optimizer Module. The stope optimizer produces the stope cross section which maximizes value within given geometric, design and economic constraints.

Design constraints included a 5 1/2-foot minimum mining width for longhole stopes with development drifts spaced at 50-foot vertical intervals. Stope development drift dimensions maintained a constant height of nine feet and a minimum width of eight feet.

Table 1-4 True North Mineral Reserves as of March 31, 2016

		Proven Reserves			Probable Reserves			Proven and Probable Reserves
District	Vein	Tons (000's)	Au opt	Au oz. (000's)	Tons (000's)	Au opt	Au Oz. (000's)	Tons (000's)	Au opt	Au Oz. (000's)
26L – 710 Complex	710	50.8	0.258	13.1	90.9	0.266	24.2	141.7	0.263	37.3
	711	12.0	0.192	2.3	70.8	0.306	21.6	82.7	0.289	23.9
	713	10.5	0.165	1.7	31.3	0.212	6.6	41.8	0.200	8.4
	708	0.1	0.148	-	18.3	0.272	5.0	18.4	0.272	5.0
	714	3.1	0.225	0.7	15.7	0.238	3.7	18.9	0.236	4.4
	712	2.0	0.149	0.3	3.3	0.187	0.6	5.3	0.173	0.9
	750	2.4	0.137	0.3	1.4	0.147	0.2	3.7	0.141	0.5
	707	-	-	-	1.5	0.208	0.3	1.5	0.208	0.3
	756	1.5	0.182	0.3	-	-	-	1.5	0.182	0.3
16L – 8/10	1030	7.9	0.264	2.1	5.7	0.194	1.1	13.6	0.235	3.2
	810	1.4	0.255	0.4	7.1	0.287	2.0	8.5	0.282	2.4
	1020	5.1	0.133	0.7	8.7	0.187	1.6	13.8	0.167	2.3
	1010	2.0	0.138	0.3	2.5	0.172	0.4	4.5	0.156	0.7
26L- SAM	V91	22.7	0.203	4.6	28.3	0.195	5.2	49.4	0.199	9.8
Cohiba	400	6.6	0.172	1.1	21.6	0.190	4.1	28.3	0.186	5.2
UG		128	0.218	27.9	306	0.251	76.9	434	0.242	104.7
Tailings					1,950	0.022	43.2	1,950	0.022	43.2
Total		128	0.218	27.9	2,256	0.053	120.1	2,384	0.062	147.9

Notes:

	1.	Mineral reserves have been estimated with a gold price of US$1,200/ounce or C$1.500/oz.
	2.	US$:CDN$ exchange rate is 0.80;
	3.	Metallurgical recovery for underground and tailings Mineral Reserves is 94% and 89% respectively;
	4.	Mineral reserves are estimated at a cut-off grade of 0.15 Au opt and an incremental cut-off grade of 0.08 Au opt;
	5.	Tailings Mineral Reserves are estimated at a cut-off grade of 0.016 opt; and
	6.	Mine losses of 5% and no unplanned mining dilution have been applied to the underground Mineral Reserves and mine losses of 8% and no unplanned dilution have been applied to the tailings Mineral Resserves.

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True North mineral reserves could be materially affected by economic, geotechnical, permitting, metallurgical or other relevant factors. Mining and processing costs are sensitive to production rates. A decline in the production rate can cause an increase in costs and cut-off grades resulting in a reduction in mineral reserves. Geotechnical conditions requiring additional ground support or more expensive mining methods will also result in higher cut-off grades and reduced mineral reserves.

The Project has the necessary permits to continue exploration and current operations. Failure to maintain permit requirements may result in the loss of critical permits necessary for continued operations.

1.8.	Cash Flow Analysis and Economics

The reserves mine plan was evaluated using constant dollar cash flow analysis, and the results are summarized in Table 1-5. The low capital requirements result in an Internal Rate of Return of 127% and a 1.8 year payback period.

Table 1-5 Key Operating and After Tax Financial Statistics

Underground Material Mined and Processed (kt)	434
Avg. Gold Grade (opt)	0.242
Contained Gold (koz)	105
Avg. Gold Metallurgical Recovery	94%
Recovered Gold (koz)	99
Reserve Life (years)	2.75
Tailings Reprocessed (kt)	1,950
Avg. Gold Grade (opt)	0.022
Contained Gold (koz)	43
Avg. Gold Metallurgical Recovery	89%
Recovered Gold (koz)	38
Reserve Life (years)	12
Cash Cost ($/oz)	$963
Total Cost ($/oz)	$1,168
Gold Price ($/oz)	US$1,200 C$1,500
Capital Costs ($ Millions)	$28.0
Payback Period (Years)	1.8
Cash Flow ($ Millions)	$25.3
5% Discounted Cash Flow ($ Millions)	$21.9
10% Discounted Cash Flow ($ Millions)	$19.0
Profitability Index (10%) ¹	1.8
Internal Rate of Return	127%

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Notes:

Profitability index (PI), is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows you to quantify the amount of value created per unit of investment. A profitability index of 1 indicates break even.

1.9.	Conclusions

P&E believes the core, channel chip and tailings sample assay data have been adequately verified for the purposes of a mineral resource estimate. All data included in the resource estimate appear to be of adequate quality.

The major capital requirements for the project are underground development and the tailings reprocessing. Capital for infrastructure and equipment is limited. Since restarting operations Klondex has rehabilitated the mine and undertaken active mining operations. The Project is technically sound with mining and processing methods proven by previous operators. Based on cash flows using estimated costs of production and revenue factors, the Project will generate a pre-tax net cash flow of $41.1 million over the LOM. At a discount rate of 10%, this corresponds to an after tax NPV of $19.0 million.

The Project is situated adjacent to a well-established mining community and has an existing infrastructure of underground openings, operating and maintenance equipment and operations personnel that can be used for future operations.

Klondex intends to apply technologies or methods that have already been accepted and previously implemented at the Project.

Klondex intends to continue to investigate extensions to the currently defined resource base.

The pre-existing mine closure plan that estimated closure costs at $4.4 million was transferred to Klondex in January 2016 along with an assignment of fixed-assets as financial security. It may be beneficial for Klondex to review the technical basis of the TMA closure approach presented in the 2012 mine closure plan and update the associated closure costs.

1.10.

Recommendations

Geology

Technical Database: All True North project data collected need to be stored and archived in a permanent and reliable retrieval manner. A full-time database administrator is recommended.

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Quality Assurance/Quality Control: Timely follow-up for any and all QA/QC assay deviations and re-assay requests should be performed in a timely manner. The process should be automated when the database is up and running.

Sample Storage and Retrieval: Half-core remaining from sample assays should be retained for reference and check assay purposes. All assay sample rejects and pulps should be stored in a safe, secure and sheltered manner and properly catalogued to ease retrieval.

Project Assay Lab: Standard operating procedures should be updated, particularly in regards to assay data generation, storage and retrieval.

Mine Operations and Planning

Production rates are directly related to the number of available work places in the mine. To achieve production rates of 700 – 800 tpd will require an inventory of at least 10 active faces. Mine planners must keep the development accesses and stope backfilling on track or the number of work places and production rate will decline with a resultant increase in unit costs.

Environmental and Mine Closure

It is recommended that Klondex review the technical basis of the TMA closure approach presented in the 2012 mine closure plan and update the associated closure costs. A provisional amount for a $250k study that would be carried out over four years commencing in 2017 is recommended. This amount has not been included in the cash flow model presented in this Technical Report. This exercise will review and confirm the technical basis of the proposed TMA closure plan and estimated costs and possibly identify opportunities to improve upon the currently proposed approach.

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2.	Introduction

2.1.	Terms of Reference and Purpose of this Technical Report

This TR provides a statement of Mineral Resources and Mineral Reserves for the Project. This evaluation includes measured, indicated, and inferred mineral resources, as well as proven and probable mineral reserves. This TR was prepared in accordance with the requirements of NI 43-101 and Form 43-101F1 (43-101F1) for technical reports.

Mineral resource and mineral reserve definitions are set forth in Section 28 of this TR in accordance with the companion policy to NI 43-101 (43-101CP) of the Canadian Securities Administrators and “Canadian Institute of Mining, Metallurgy and Petroleum (CIM) – Definition Standards for Mineral Resources and Mineral Reserves adopted by CIM Council on May 10, 2014.”

2.2.	Qualification of the Authors

This TR includes technical evaluations from six independent consultants. The consultants are specialists in the fields of open pit and underground mining.

None of the authors has any beneficial interest in Klondex or any of its subsidiaries or in the assets of Klondex or any of its subsidiaries. The authors will be paid a fee for this work in accordance with normal professional consulting practices.

The individuals who have provided input to the current TR are cited as “author” and are listed below in Table 2-1. These authors have extensive experience in the mining industry and are members in good standing of appropriate professional institutions.

Table 2-1 Qualified Professionals

Company	Name	Title	Discipline	Contributing Sections
Practical Mining, LLC	Sarah Bull	P.E.	Mining	1, 15,16, 21,22 & 24-26
Practical Mining LLC	Mark Odell	P.E.	Mining	1-3, 15, 16, 18-19, 21, 22 & 24-28
P&E Mining Consultants	Dr. William Stone	P. Geo.	Geology	1, 4-12, 23, 25 & 26
P&E Mining Consultants	Fred Brown	P. Geo.	Geology	1, 14, 25,26
P&E Mining Consultants	Alfred S. Hayden	P. Geo.	Metallurgy	1, 13, 17, 25 & 26
P&E Mining Consultants	David Oeava	P. Eng.	Mining	1, 20,25 & 26

2.3.	Sources of Information

The sources of information include data and reports supplied by Klondex staff.

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Additional information is included in the TR which is based on discussions with Klondex staff as it relates to their field of expertise at the Project. The required financial data and operating statistics were also provided by Klondex staff.

This report is based in part on internal Company technical reports and maps, published government technical reports, published scientific papers, company letters and memoranda, and public information listed in Section 27.

Sections from reports authored by other consultants have been directly quoted or summarized in this Technical Report and are indicated as such within the appropriate sections. PM and P&E held discussions with technical personnel from the Company regarding pertinent aspects of the Project. PM and P&E have not conducted detailed land status evaluations, and has relied on previous qualified reports, public documents and statements by Klondex management and legal counsel, regarding the status and legal title to True North.

Information sources are documented either within the text and cited in references, or are cited in references only. The primary author believes the information provided by Klondex staff to be accurate based on their work at the Project. The authors asked detailed questions of specific Klondex staff to help verify contributions included in this document. These contributions are clearly stated within the text.

2.4.	Units of Measure

The units of measure used in this report are shown in Table 2-2 below. U.S. Imperial units of measure are used throughout this document unless otherwise noted.

Currency is expressed in Canadian dollars unless stated otherwise. The glossary of geological and mining related terms is also provided in Section 28 of this TR.

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Table 2-2 Units of Measure

US Imperial to Metric conversions

Linear Measure

1 inch = 2.54 cm

1 foot = 0.3048 m

1 yard = 0.9144 m

1 mile = 1.6 km

Area Measure

1 acre = 0.4047 ha

1 square mile = 640 acres = 259 ha

Weight

1 short ton (st) = 2,000 lbs = 0.9071 metric tons

1 lb = 0.454 kg = 14.5833 troy oz

Assay Values

1 oz per short ton = 34.2857 g/t

1 troy oz = 31.1036 g

1 part per billion = 0.0000292 oz/ton

1 part per million = 0.0292 oz/ton = 1g/t

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3.	Reliance on Other Experts

Statements in this Technical Report regarding the status and legal title of the Project, are reliant on information provided by Klondex legal counsel.

The status of the Klondex environmental program and the permitting process was provided by the Environmental Superintendent for the Project. The corporate Manager of Metallurgy provided information regarding metallurgical testing and process operating statistics. These contributions have been reviewed, edited and accepted by the Authors of this report and are accurate portrayals of True North, as of the effective date of this Technical Report.

Operating and capital cost projections related to the production and sale of doré gold have been provided by Klondex. This information has been reviewed and accepted by the Authors as being reasonable, as of the effective date of this Technical Report.

A draft copy of this Technical Report has been reviewed for factual errors by Klondex and this Technical Report is based in part on Klondex’s historical and current knowledge of the True North Project.

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4.	Property Description and Location

4.1.	Property Location

The Project is located adjacent to the township of Bissett on the north shore of Rice Lake in southeastern Manitoba, 100 miles (160 km) northeast of the city of Winnipeg (Figure 4-1). The Project includes the mine, mill, and tailings management area. Klondex’s property holdings in Manitoba, Canada include a larger regional exploration boundary as outlined in Figure 4-2 of this Technical Report.

Bissett can be accessed from Winnipeg via all-weather provincial highways. A small emergency gravel airstrip is located 12 miles (19 km) east of Bissett. Rice Lake serves as a base for float-equipped aircraft during the ice-free months.

Geographical co-ordinates are:

	latitude 51^o01’ 19.6” N longitude 95^o40’ 44.9” W

	UTM WGS84 Zone 15U 312,110 m E 5,655,700 m N

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Figure 4-1 Location of the True North Mine, Bissett, Manitoba

The locations of all mineralized zones, mineral resources, mineral reserves, mine workings, tailings management area (TMA), and waste deposits are shown on various figures in other sections of this Technical Report.

The boundaries of the mining lease (ML-063) and of the patented mining claims have been surveyed, whereas the boundaries of other, un-surveyed mining claims, are sourced from government claim maps.

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4.2.	Property Description

The Project consists of claims, patents and a mineral lease owned 100% by the Company (Figure 4-2). The total area covered by the Project is 97,080 acres (39,287) ha (Table 4-1).

Table 4-1 Summary of Klondex Mineral Property Holdings and Surface Areas

Item	Claims/Patents/Leases	Acres	Hectares
Unpatented Mining Claims	294	95,089	38,481
Patented Mining Claims	18	731	296
Mineral Lease	1	1,893	766
TOTAL	313	97,713	39,543

Klondex retains a 100% recorded interest in mineral lease ML-063 (“ML-063”). This lease covers 1,893 acres (766 hectares (ha)) and, subject to annual payments, expires April 1, 2034.

In addition to ML-063, Klondex also holds 18 Patented Mining Claims covering an area of 731 acres (296 ha) and 294 Mining Claims covering an area of 95, 089 acres (38,481 ha). The Company owns 100% of 267 of the Mining Claims.

Klondex owns 50% of the remaining 27 Mining Claims through a Joint Venture Agreement with Greenbelt Gold Mines Inc. (see Appendix I for full listing of claims information).

4.3.	Liabilities and Permits

All environmental liabilities are disclosed in Section 20, which covers the mine closure plan. All permits required to perform work are also disclosed in Section 20.

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Figure 4-2 Regional Klondex Mining Claim and Lease Holdings

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

5.1.	Access to Project

Bissett can be accessed from Winnipeg via all-weather provincial highways. A small emergency gravel airstrip is located 12 miles (19 km) east of Bissett. Rice Lake itself serves as a base for float-equipped aircraft during the ice free months.

5.2.	Climate

This area of eastern Manitoba has average annual precipitation of approximately 17 inches (430 mm) of rain. Winter snow accumulations of up to 57 inches (145 cm) occur between October and March. Average winter temperature is 3^oF (-16°C) with extended periods of -4^oF to -13^oF (-20°C to -25^oC). Average summer temperature is 61^oF (16°C).

5.3.	Vegetation

The vegetation consists of typical Canadian Shield boreal forest. Poplar, balsam, spruce, and pine are the main tree species. Rock outcrop exposure is abundant in most areas, although there is a thin cover of organics and lichen growth that can restrict detailed observation.

5.4.	Physiography

Average relief in the Project area is approximately 130 feet to 200 feet (40m to 60m), with elongated outcrop ridges separated by low lying ground with swamps, rivers and lakes. Ground elevation of the surface facilities is roughly 840 feet and the tailings pond lies at roughly 905 feet.

5.5.	Local Resources and Infrastructure

Bissett is an established mining community, located adjacent to the mine, with a fluctuating population of approximately 340 people. The township was established to service the emerging mines that developed after 1911, but has remained home to permanent residents during periods of mine closure and now provides a healthy recreational sport base as well as servicing the Project.

Mining supplies, equipment, and services and a skilled mining and mineral exploration workforce are readily available in southern Manitoba and across the border to the established mining communities in northeast Ontario. The Project has a long history of mining, which helps to attract employees and contractors from throughout the area.

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Manitoba Hydro provides electrical power to site via twin transmission lines. Fuel is trucked in from Winnipeg and the area is well serviced by access roads.

Klondex owns 100% of the mine shaft, declines, mobile and crushing equipment, mineral processing mill, storage areas (Figure 5-1 and Figure 5-2) and TMA (Figure 5-3).

The process plant is licensed to operate at up to 2,500 tons (2,268 tonnes) per day. The Company has sufficient accommodation on-site for all personnel and provides cafeteria services for employees housed in the Project’s bunkhouse accommodation.

A small school provides education up to grade six. A bar, hotel, restaurant, and convenience store provide services for residents and visitors to the town. The township has recreational infrastructure such as a curling rink, outdoor ice skating rink and a baseball diamond.

Figure 5-1 Photograph of the True North Gold Mine Looking South

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Figure 5-2 Plan View of Surface Infrastructure at the True North Gold Mine

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Figure 5-3 Tailings Management Area at the True North Gold Mine

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6.	History

For the historic estimates in this section of the Technical Report, Qualified Persons from either Practical Mining, P&E or Klondex have not done sufficient work to classify the historical estimates as a current Mineral Resource or Mineral Reserve, and Practical Mining, P&E and Klondex are not treating these historical estimates as current Mineral Resources or Mineral Reserves. The historical estimates cannot be fully verified. These values cannot and should not be relied upon and are only referred to herein as an indication of previously defined gold mineralization. The relevance of the historical estimates is not known. Key assumptions, parameters and methods used to estimate these Mineral Resources and Mineral Reserves are not known. The historical Mineral Resource and Mineral Reserve estimates described in this Technical Report section have been superseded by the Mineral Resource and Mineral Reserve estimates described herein (see Section 14 and 15 of this Technical report).

6.1.	Events Prior to 1989

Gold was originally discovered at the shore of Rice Lake in 1911. The first attempt at underground development was undertaken by a syndicate in 1927, when No.1 exploration shaft was sunk to 164 feet (50 m) and No.2 Shaft was sunk to 300 feet (91 m). Approximately 2,000 feet (610 m) of lateral development was completed in 1927, but results failed to meet expectations. Nevertheless, during 1928 the syndicate proceeded to deepen the No.2 Shaft to 600 feet (183 m) and the No. 1 Vein was discovered on that level. However, it was not until 1929, with discovery of the No. 9 Vein on the 725-foot (221 m) level, that the deposits became economically viable.

Sufficiently encouraging underground results were obtained by 1931, and the newly formed San Antonio Gold Mines Ltd. (“San Antonio”) commenced construction of a process plant and power line. Production began in May 1932 at a rate of 150 tons (136 tonnes) per day, increasing to 350 tons (318 tonnes) per day in 1935, and subsequently increased to 550 tons (500 tonnes) per day by 1948. Access to the mine was primarily through the No.1 Shaft (now called the A-Shaft) and three internal winzes; 3A, 3B, and 3C (now called B-Shaft, C-Shaft, and D-Shaft).

Underground development was carried out by driving footwall drifts on each level. Flat exploration drill-holes on 50-foot (15 m) centres were used to establish the location of veins on the level prior to establishing drifts along the full length of ore zones. Shrinkage mining was used with a minimum mining width of 4 feet (1.2 m).

The 550-ton (500-tonne) per day process plant consisted of a crushing plant adjacent to the collar of No.1 Shaft with a conveyor to the process plant building. After grinding, concentrating, and blanket tables, an amalgam table recovered approximately 12% of the total gold. Then the material from the gravity circuit passed through a Merrill Crowe cyanide plant to recover the balance of the gold.

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The No.1 Shaft surface hoist was destroyed by fire in July 1968 and production ceased. Historic production at Rice Lake Mine through 1968 is summarized in Table 6-1. San Antonio declared bankruptcy and the assets were acquired by New Forty Four Mines (“New Forty Four”). In 1980, the process plant was destroyed by fire.

In 1980, Brinco Mining Limited (“Brinco”) entered into a Joint Venture with New Forty Four. Brinco undertook a program of underground exploration drilling during the period 1980 through 1983 and approximately 100,000 ore tons (91,000 tonnes) were mined and trucked to Hudson Bay Mining & Smelting Co Ltd. in Flin Flon, Manitoba, for processing. Brinco earned a 100% interest in the project, however, after 1983 did no significant work.

In 1987, a subsidiary of Inco Ltd. (“Inco”) entered into an agreement with Brinco and completed over 20,000 feet. (6,096m) of drilling. Inco opted out of the venture in 1988.

6.2.	Events since 1989

In 1989, Rea Gold Corp. (“Rea Gold”) acquired the Property from Brinco. Wright Engineers and Dolmage Campbell completed a due diligence study for Rea Gold prior to their acquisition of the Project in 1989. At that time, the historic (non NI 43-101 compliant) Mineral Reserve was estimated to be 1,226,000 tons (1,114,000 tonnes) grading 0.22 opt Au (7.5 g/t Au). A Pre-Feasibility study by Kilborn Engineering Ltd. (Kilborn) in 1993 recommended that the resource base be increased prior to a production decision.

In 1994, Rea Gold undertook a $3.1 million underground rehabilitation and exploration program to gain access to the lower levels of the mine and delineate additional Mineral Resources. This program resulted in an increase in the historic (non NI 43-101 compliant) Mineral Reserves to 1,977,000 tons (1,797,000 tonnes) grading 0.21 opt Au (7.2 g/t Au).

A Feasibility Study was completed by Rea Gold and Simmons Engineering Inc. in 1995, and construction and development of a 1,000 ton (907 tonne) per day mining operation was initiated. Rea Gold established a new mine access system that significantly streamlined the mining operation. Previously, the mine was accessed by A-Shaft and three internal winzes (B-Shaft, C-Shaft, and D-Shaft). Ore from the D-Shaft area had to be trammed and hoisted via four shafts in order to transport it to surface. Rea Gold deepened the principal A-Shaft to link the surface directly with the upper level of the D-Shaft area, thereby eliminating two cycles of tramming and hoisting.

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By 1997, Rea Gold had established a modern 1,000 ton (907 tonne) per day gold mining and processing facility at a total cost of approximately $90 million. Prior to the start of production, Rea Gold was placed into receivership and the receiver put the assets up for sale. Harmony Gold (Canada) Inc. (“Harmony”) was the successful bidder and took over the project in 1998.

After acquiring the assets from the receiver, Harmony invested approximately $30 million to build a ramp system in the lower part of the D-Shaft area, in order to establish a longhole mining operation. Harmony operated the mine for three years, and subsequently put the project on care and maintenance in August, 2001. Compared to the previously employed shrinkage mining operation, the Harmony operation produced fewer ounces of gold from more tons processed per day and failed to achieve the corporate objectives set by Harmony’s parent company, Harmony Gold Mines Limited of South Africa. Historic production at Rice Lake Mine from 1980 through 2001 is summarized in Table 6-2.

In January, 2002, Harmony entered into an option agreement with Wildcat Exploration Ltd. of Winnipeg, Manitoba (“Wildcat”). Wildcat’s objective was to re-establish the mine as a smaller scale shrinkage stope operation delivering ore to a surface stock pile to feed the 1,250 ton (1,136 tonne) process plant which operated on a two week-on two week-off cycle.

In April 2002, A. C. A. Howe International (“Howe”) (Titaro et al 2002) completed a report on the Harmony assets on behalf of Wildcat. The report included an audit of the mineral resources and mineral reserves, a review of the operating and capital costs, and preparation of a financial evaluation of the economic feasibility of reopening the mine. Howe concluded that a viable shrinkage mining operation could be operated at a mining rate of 550 tons (500 tonnes) per day was feasible. Ore was delivered to a surface stockpile to feed the 1,250 ton (1,136 tonne) per day process plant operating on a two week on, two week off cycle. Gold at that time was US$300/oz.

Howe further concluded that based on well-founded historical estimation practices at the Rice Lake Mine (as it was then called), that as of April 2001, the mine, had a historical (not NI 43-101 compliant) Measured and Indicated Mineral Resource of 1,267,000 tons (1,149,000 tonnes) grading 0.26 opt Au (8.9 g/t Au) plus Inferred Mineral Resource of 735,000 tons (668,000 tonnes) grading 0.31 opt Au (10.6 g/t Au). All of the above mentioned Mineral Resources were situated above the 4,630 Level (5,370 feet or 1,637 m below the collar of A-Shaft) in the C and D-Shaft areas of the Rice Lake Mine.

Within the Measured and Indicated Mineral Resources, Howe concluded that the Rice Lake Mine had Proven and Probable Mineral Reserves of 901,800 tons (820,000 tonnes) with an average grade of 0.27 opt Au (9.3 g/t Au). In determining this reserve, Howe used dilution, cutting, and cut-off practices which were based on over 38 years of mining experience at the Rice Lake Mine (now True North Gold Mine). All of these mineral reserves had existing development drifts and were accessible on levels within the C-Shaft and D-Shaft areas.

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Despite this work by Howe, Wildcat was unable to complete the acquisition of the Rice Lake Mine.

On March 5, 2004, San Gold Resources Corporation (“Old San Gold”) and Gold City Industries Ltd. (“Gold City”), entered into a joint venture agreement to acquire 100% of the issued and outstanding shares of Harmony through a newly formed corporation, Rice Lake Joint Venture Inc. (“RLJV”). RLJV was owned and controlled jointly by Gold City (50%) and Old San Gold (50%). Effective March 17, 2004, RLJV acquired the shares of Rice Lake Gold Corporation (formerly Harmony Gold Corporation (Canada) Inc.) from Harmony Gold Mining Company Limited of South Africa. The purchase price was $7,757,961, including $3,632,961 in cash and $4,125,000 in shares and warrants of Gold City and Old San Gold. On June 30, 2005 Old San Gold and Gold City amalgamated to form a new corporation called San Gold Corporation.

The exploration drilling completed between the period from 2005 to 2013, (is summarized below and more fully described in Section 10 of this Technical Report). As part of San Gold’s exploration program, a Light Detection and Ranging (LiDAR) survey was flown over the Rice Lake greenstone belt in 2009. From this a second mining trend called the Shoreline Basalt unit, which hosts the Hinge and 007 Zones, was recognized

A ramp to explore and develop the new, separate SG1 deposit commenced in the winter of 2005. Production from that deposit continued until mid-2008 when workings had reached a depth of 640 feet (195m) below surface. Work was suspended in 2008 due to diminishing economics and the mobile equipment was needed elsewhere to develop the recently discovered Hinge Zone.

A new surface ramp to explore and develop the Hinge Zone commenced in 2008 and reached the deposit in March 2009. Production started almost immediately as definition drilling continued.

In early 2010, a new internal ramp was started from a vertical depth of 800 feet (244m) in the Hinge Zone workings to access the 007 deposit. The ramp reached the 007 deposit in July 2010, and production started while definition drilling continued.

A second surface ramp was started near the old Wingold shaft in the second half of 2010. This ramp was to provide top access to the 007 deposit and provide access to develop the Cohiba deposit. The ramp reached the Cohiba mineralization at a vertical depth of 108 feet (33m) below surface.

Under the San Gold operations, ore was mined along three active underground mining trends. The Rice Lake Mine (as the Project was previously named), formed the core of mining operations and provided extensive workings that extended deep below the surface to access these deposits situated within the mineralized host system. The A-Shaft head frame and extends 4,060 feet (1,249 m) below surface.

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The second mining trend, the Shoreline Basalt mining unit, focused on the 007 deposit, began commercial production in 2010. The 007 mine portal is located 2,000 feet (600 m) from the process plant and provided the main access to San Gold’s operations along the Shoreline Basalt mining unit and the Hinge Zone. The Hinge zone is hosted in intermediate rocks, is the third mining trend host and its portal provides access to the 007 deposit.

After investing approximately C$375 million in capital since 2007, including the extensive underground development and modernizing the process plant, San Gold ceased mining in May 2015, and placed operation on care and maintenance. San Gold declared bankruptcy and announced sale of all of its assets to secured creditors in June 2015. Historic Production at Rice Lake Mine from 2007 through 2015 is summarized in Table 6-3.

In early 2016, Klondex Mines Ltd. announced acquisition of 100% of the Rice Lake Mine, process plant complex and a 400 km2 exploration land package from the creditors. In the first half of 2016, Klondex commenced refurbishment of the underground infrastructure and commenced trial mining on readily accessible ore.

Following sampling of the historic tailings storage facility, Klondex commenced a tailings reprocessing project. Reprocessing of the tailings will be carried out concurrently with processing of underground ore and when weather permits. Processing of stockpiled ROM ore is expected to commence in fourth quarter 2016.

A name change to True North Gold Mine was announced in May, 2016. In September 2016, Klondex announced the formal decision to resume production at True North.

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Table 6-1 Historic Production at Rice Lake Mine: 1927-1968

	Mill Throughput						Ore Reserves			Notes
		% Recovery of		Process		Head
				Plant
	Gold	Head	Stope	Feed	Average	Grade	Gold	Reserve	Grade
YEAR	ozs	Grade	Grade	tons	tons/day	opt	ozs	tons	opt
1927	27,008	181%	169%	30,419	83	0.49	34,992	74,450	0.47	Process Plant starts May 1932
1933	22,720	95%	94%	55,677	153	0.43	63,140	154,000	0.41
1934	21,638	93%	90%	64,294	176	0.36	87,750	225,000	0.39	Gold fixed at $35/oz from $20/oz
1935	32,250	92%	96%	102,712	281	0.34	77,070	226,675	0.34
1936	29,040	96%	86%	112,416	308	0.27	59,351	197,836	0.30
1937	30,035	93%	93%	115,765	317	0.28	71,824	256,516	0.28	Discovered 38 vein
1938	31,257	95%	96%	117,376	322	0.28	96,184	343,515	0.28
1939	34,242	94%	94%	117,787	323	0.31	152,361	491,486	0.31	Start of World War 2
1940	36,745	94%	93%	122,365	335	0.32	242,150	756,718	0.30
1941	43,121	95%	94%	138,097	378	0.33	312,501	945,609	0.30
1942	58,869	95%	95%	199,203	546	0.31	285,481	920,908	0.31
1943	48,568	95%	97%	164,307	450	0.31	256,612	916,471	0.28
1944	40,669	97%	96%	140,085	384	0.30	256,735	855,784	0.30
1945	38,326	98%	97%	135,000	370	0.29	213,562	736,419	0.29	End of World War 2
1946	43,819	97%	98%	149,875	411	0.30	214,738	715,794	0.30
1947	42,326	99%	100%	137,867	378	0.31	215,173	694,105	0.31
1948	52,764	114%	113%	154,953	425	0.30	214,826	716,087	0.30	Emergency Gold Mining Assistance started
1949	53,201	105%	104%	188,000	515	0.27	193,860	718,000	0.27
1950	51,822	101%	102%	182,397	500	0.28	198,562	709,151	0.28
1951	50,735	96%	96%	195,000	534	0.27	174,150	645,000	0.27
1952	53,120	95%	95%	200,000	548	0.28	168,112	600,400	0.28
1953	40,993	98%	99%	174,904	479	0.24	102,816	428,400	0.24	Gold free market ends
1954	43,868	97%	98%	180,599	495	0.25	86,900	347,600	0.25
1955	41,211	98%	99%	174,631	478	0.24	70,800	295,000	0.24	First operating loss
1956	33,462	98%	99%	155,595	426	0.22	54,868	249,400	0.22
1957	33,339	98%	98%	136,616	374	0.25	48,648	202,700	0.24
1958	34,300	98%	98%	124,597	341	0.28	56,650	226,600	0.25
1959	28,570	98%	98%	116,666	320	0.25	47,444	197,683	0.24
1960	31,136	96%	95%	135,642	372	0.24	47,928	199,700	0.24
1961	31,009	98%	99%	149,942	411	0.21	36,869	160,300	0.23
1962	30,339	99%	98%	133,000	364	0.23	23,218	110,560	0.21
1963	24,017	94%	94%	127,575	350	0.20	46,886	203,853	0.23
1964	28,773	98%	98%	133,764	366	0.22	44,989	187,454	0.24
1965	24,969	98%	97%	111,295	305	0.23	34,966	145,693	0.24
1966	21,630	98%	97%	85,258	234	0.26	40,241	174,961	0.23
1967	13,394	98%	98%	71,673	196	0.19	43,240	188,000	0.23

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	Mill Throughput						Ore Reserves			Notes
				Process
		% Recovery of				Head
				Plant
	Gold	Head	Stope	Feed	Average	Grade	Gold	Reserve	Grade
YEAR	ozs	Grade	Grade	tons	tons/day	opt	ozs	tons	opt
1968	6,066	87%	93%	30,218	166	0.23	38,769	161,537	0.24	Fire destroys surface hoist; production ends July, 1968.

Table 6-2 Historic Production at Rice Lake Mine: 1980-2001

	Mill Throughput						Ore Reserves			Notes
		% Recovery of		Process Plant		Head			Grade
Year	Gold				Average		Gold	Reserve
		Head	Stope	Feed		Grade
	ozs	Grade	Grade	tons	tons/day	opt	ozs	tons	opt
1980-83	13,954	100%		104,135		0.13	146,085	664,024	0.22	New Forty Four/ Brinco Joint Venture formed
	Mill destroyed by fire in 1980. Production ends May 227, 1983, drilling continues at depth
1984							111,616	534,504	0.21	Lathwell/Brinco JV conducts limited program
1985	Brinco changes name to Cassiar Mining Corporation
1986							350,196	1,522,591	0.23	Inco subsidiary drills 20,008 ft to test depth
1987	Inco opts out. Cassiar ownership 100%
1988							308,700	1,470,000	0.21	Kilborn reviews reactivation program for Mandor Gold
1989							169,641	1,225,642	0.22	Rea Gold Corp. acquires project from Cassiar.
1990	Wright Engineers and Dolmage Campbell complete due diligence on behalf of Rea Gold
1993	Pre-Feasibility of Kilborn and Tonto recommends mineable reserves be increased
1994							415,149	1,976,901	0.21	Rehab, exploration and development in lower levels of mine

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	Mill Throughput						Ore Reserves			Notes
		% Recovery of		Process Plant		Head			Grade
Year	Gold				Average		Gold	Reserve
		Head	Stope	Feed		Grade
	ozs	Grade	Grade	tons	tons/day	opt	ozs	tons	opt
1995							540,715	2,216,046	0.24	Feasibility studies by Rea Gold and Simmons completed. Drilling and development underground.
1996							558,213	2,335,621	0.24	Construction and development towards 1,000 tons per day operation
1997	9,000			60,000		0.15	674,951	2,812,297	0.24
1998	2,875			40,035		0.07				Rea Gold Corp. bankrupt. Receiver puts assets up for sale. Harmony Gold (Canada) Inc. acquires mining assets of Harmony.
1999	33,238			231,898		0.14
2000	39,476			257,605		0.15
2001	29,341	85%	79%	203,868		0.17	327,884	1,932,404	0.17	Project placed on care and maintenance August, 2001

Table 6-3 Historic Production at Rice Lake Mine: 2007-2015

Year	Tons Processed	Head opt	Grade g/t	Gold oz
2007	96,653	0.13	4.35	9,193
2008	116,835	0.09	3.20	13,845
2009	164,424	0.23	8.00	35,154

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2010	275,860	0.17	5.85	47,082
2011	461,150	0.17	5.93	79,802
2012	629,279	0.15	5.07	93,233
2013	641,711	0.13	4.32	80,828
2014	390,564	0.12	4.03	41,890
2015 (Q1)	81,427	0.11	3.91	9,261

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7.	Geological Setting and Mineralization

7.1.	Regional Geology

True North is underlain by the Archean Rice Lake greenstone belt located at the west end of the Uchi Volcanic-Plutonic Subprovince of the Superior Province (Figure 7-1 and Figure 7-2). The Rice Lake greenstone belt is bound to the north and south by the Wanipigow Shear Zone and the Manigotagan Shear Zone, respectively.

The Manigotagan Shear Zone is marked by a regionally continuous zone of faults (Manigotagan-Lake St. Joseph Fault) which separates the volcanic-plutonic terrain of the Uchi Subprovince from the English River (Ontario) - Manigotagan (Manitoba) gneissic belts. The Manigotagan gneissic belt, which occurs immediately south of the Rice Lake greenstone belt, consists of a lithologic gradation from low-grade metavolcanic and metasedimentary rocks, through paragneiss and migmatite, to quartz diorite and granodiorite gneiss.

The Wanipigow Shear Zone is marked at True North by regional-scale fault structures and elsewhere by increasing metamorphic grade into the metamorphic-plutonic terrain of the Wanipigow (Manitoba) Subprovinces and Berens River (Ontario). The Wanipigow River Plutonic Complex, which forms the northern boundary of the Rice Lake greenstone belt, is composed mainly of hornblende and biotite-bearing quartz diorite, granodiorite and locally quartz monzonite intrusions and gneisses. Several large gabbro intrusions are also present.

The rocks in the True North area were affected by at least three and possibly four major periods of deformation (Anderson, 2008). The resulting fold pattern is complex with overturned, doubly-plunging folds in the Rice Lake Group rocks. The late Archean San Antonio Formation sedimentary rocks may have been affected by only the last major period of deformation.

Many major regional fault structures are present in the True North area. The most prominent are the major structures that trend generally east-west. Movement along these structures formed conjugate shear zones which splay off to the north and south. Thrust faulting likely occurred in the early stages of the deformation, but these structures are difficult to identify.

All the major gold occurrences in the Project area occur as quartz veins or quartz vein systems formed during structural deformation of the host rocks. Significant gold production has occurred from the Uchi Subprovince in the Rice Lake area to the west in Manitoba and in the Red Lake, Birch-Uchi Lake and Pickle-Dona Lake areas to the east in Ontario (Figure 7-2).

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Figure 7-1 Regional Geologic Map showing the Location of True North Gold Mine in the Archean Uchi Subprovince, Manitoba

Figure 7-2 Geologic Map showing the Location of Gold Deposits and Lithotectonic Assemblages in True North Gold Mine Area

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7.2.	Local Geology

All of the gold mineralized zones at True North are hosted in rocks of the Bidou Lake Assemblage (Figure 7-3). The Bidou Lake Assemblage forms a north-facing stratigraphic sequence of tholeiitic basalt to intermediate volcanic flows, dacite crystal tuffs and breccias overlain by well stratified felsic epiclastic rock interpreted to be of pyroclastic and sedimentary origin. The stratigraphic sequence is intruded by tholeiitic gabbro sills and dykes and felsic porphyry dykes.

The best known gabbro sill is the San Antonio Unit, host rock of the gold mineralization at the True North deposit. The Bidou Lake Assemblage is unconformably overlain by feldspathic sandstone of the San Antonio Assemblage.

In the Project area, gold mineralization is controlled by quartz-carbonate veins and vein systems in brittle- ductile structures with related hydrothermal alteration halos within or at the margin of particular host rock units (Figure 7-4).

7.2.1.

Host Rock Units

The gold mineralized veins show a high degree of structural control and are best developed in competent host rock units. Since 2009, three main host units have produced the most gold ore at True North (Figure 7-3):

1) The SAM, a gabbro sill from which gold has been mined for more than 80 years from the True North deposit;

2) The Shoreline Basalt unit, which hosts the 007, L10 and SG zones; and

3) The Intermediate Volcanic unit, which hosts the Hinge, L13, L08 and Cohiba Zones.

The SAM unit is a layered tholeiitic gabbro sill which intrudes the Bidou Lake Assemblage and dips moderately to the north. The SAM has been interpreted to be a subvolcanic feeder for the overlying mafic volcanic rocks. SAM hosted all the gold mineralization mined prior to 2004 in the True North Mine and the Cartwright zone, for a total of 1.5 million oz.

The Shoreline Basalt is a steeply dipping mafic volcanic rock unit that is geologically similar and subparallel to the SAM unit, and hosts the 007, L10 and possibly the SG gold zones (Figure 7.4) .

The Intermediate Volcanic Unit occurs to the north of the Shoreline Basalt and hosts the Hinge and Cohiba Zones.

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In addition to these three rock units, an unnamed intermediate to mafic volcanic unit situated in the footwall to the Shoreline Basalt unit hosts the 710 Zone which was discovered in 2013.

Figure 7-3 Local Geology of True North Gold Mine Area

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Figure 7-4 The Structural Geological Setting of Gold Mineralization at True North Gold Mine

7.2.2.

Structure

The structures that control the gold mineralization are brittle-ductile shear zones which strike from parallel to transverse to the host rock units and dip steeply northwest or northeast. The shear zones are marked by intensely foliated and lineated interlayered sericite and chlorite schists, which range from <100m to 6 km long and from 1m to >10m thick (Figure 7-5A).

Structures trending east-northeast have kinematic features indicative of sinistral-reverse movement, whereas those trending northwest have kinematic features indicative of dextral-normal movement.

The sinistral and dextral structures are interpreted to have been generated during a single protracted areal deformation event – D3 (Anderson, 2011; SRK, 2013). Stretching lineation and fold plunges tend to be orthogonal to movement on the host shear zone (SRK, 2013). The structures contain a main, banded (laminated) quartz vein and subsidiary veins in the schist on either side (Figure 7-5B). The main vein can be situated anywhere in the structure.

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7.2.3.

Veins

According to Anderson (2008), shear-hosted veins include massive, laminated and brecciated varieties, commonly within the same vein, and typically pinch and swell along strike and down-dip. Thicker veins are associated with inflection points in the host shear zones, which suggests hydrothermal infill of dilational jogs.

Most of the shear zones are associated with fringing arrays of kinematically linked extension and oblique-extension quartz veins, which locally intensify into complex peripheral stockwork-breccia systems. Considered with the geometry of the vein arrays, the vein textures indicate synkinematic emplacement under brittle-ductile conditions. Most deposits are arrays of sub-horizontal extension veins, which suggests emplacement accompanied by transiently supralithostatic fluid pressures.

In the True North deposit, the gold-bearing quartz veins occur mainly as either “16-type” shear zone veins or “38-type” tensional fracture stockwork veins or, where they intersect, as a combination of the two vein types. The 16-type appear to be fault fill veins with generally higher grades and more continuity, which are laminated with pressure solution seams (stylolites) and trend north-northeast. Examples of both vein types are shown in Figure 7-5 and Figure 7-6.

The stylolites consist of intergrown pyrite-chlorite-tourmaline-muscovite. Compared to the 16-type, the 38-type are stockwork breccia veins that are wider and arranged in an en-echelon pattern along the strike and down the dip of the host gabbro unit, but gold mineralization is more irregular and grades difficult to predict. In some deposits, for example SG-1 and SG-3, the gold mineralized veins were intensely transposed during ductile deformation (Anderson, 2008), and presumably later in the D3 deposit.

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Figure 7-5 Shear Zones and Quartz Veins

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Figure 7-6 Example of 16-Type Shear and 38-Type Breccia Gold Mineralized Quartz Veins in the SAM Unit at True North

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In addition to quartz, the veins contain subordinate carbonate, minor albite, chlorite and sericite, and rare tourmaline and fuchsite (a.k.a. mariposite). The carbonate is dolomite-ankerite in composition (Ross & Rhys, 2010). Sulphide minerals consist of pyrite with minor chalcopyrite and rare sphalerite, galena and gold-silver telluride minerals. Pyrite generally comprises <5% of individual veins and occurs as scattered grains and irregular blebs within and along vein margins, and is concentrated along planar slip surfaces or stylolites.

Gold typically occurs as free grains associated with or as inclusions in pyrite. Gold grades tend to be highly erratic within individual quartz veins. The gold ores have high Au/Ag ratios of >5:1 and low concentrations of copper, lead, zinc, arsenic, bismuth, boron, antimony and tungsten, as is typical for Archean lode-gold deposits.

7.2.4.

Alteration

Wall rock alteration spatially associated with the quartz veins varies from minor to intense and is generally zoned outward from proximal albite + ankerite + sericite + quartz +pyrite through medial chlorite + ankerite ± sericite to distal chlorite + calcite (Anderson, 2008). These alteration mineral assemblages overprint the regional greenschist facies metamorphic mineral assemblage (Ames et al., 1991). Many veins show evidence of wall rock sulphidization in the form of coarse euhedral pyrite grains.

In the True North deposit, thick zones of altered and sulphidized wall rock with minor vein quartz contain ore grade gold. Complex and antithetic distribution patterns of phengitic white mica and muscovite-paragonite are reported by SRK (2013), and appear to be controlled by second order faults and near-mine shear zones. Figure 7-7 shows typical shear orientations and general alteration assemblages.

The True North and SG-1 deposits show close spatial relationship with laterally continuous zones of ankerite-sericite phyllite and phyllonite, which represent reliable guides to ore. Deformation structures in the phyllonite preserve evidence of a complex deformation history, increments of which pre-date and post-date vein formation.

Despite vertical extents of up to >2 km, the True North deposit shows only minor variation in vein mineralogy, texture and structure.

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Figure 7-7 Controls on Gold Mineralization in the 007 Zone

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Figure 7-8 Interpreted Cross-Sections of 007 and L10 Zones Looking West

Figure 7-9 Level Plan of the 710 Zone showing the Location of the 710 and 711 Veins

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8.	Deposit Types

The association of gold at True North with quartz-carbonate veins in brittle-ductile shear zones and laterally extensive hydrothermal alteration zones indicates that the deposits represent epigenetic mesothermal lode gold-type (Poulsen et al., 2000) or orogenic-type gold mineralization (Groves et al., 1998).

Such gold deposits form from metal-bearing fluids generated during accretionary processes and prograde regional metamorphism at depth in greenstone belt terrains. In this model (Figure 8.1), the resulting fluids migrate and are channeled upward along transcrustal fault systems to subsidiary shear and fracture structures developed in the middle to upper crust. Gold is deposited in quartz carbonate veins as a result of pressure-temperature, pH and other physiochemical changes, phase separation and fluid-rock reactions. The reactions commonly involve sulphidization of precursor oxide, carbonate and silicate minerals and mineral assemblages.

Figure 8-1 Schematic Cross-Section Representation of the Geometry and Structural Setting of Shear Zone Hosted Gold-Bearing Quartz Vein Networks in Greenstone Belt Terrains like True North Gold Mine

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9.	Exploration

Klondex has yet to commence near mine and regional exploration work at True North. Limited underground diamond drilling exploration has been completed for ore definition. Most of the exploration work at the Project was completed previously during the San Gold ownership.

Based on the orogenic gold model (Figure 8.1), exploration targets at True North are areas or zones selected based on the criteria listed below:

	•	Presence of gold;
	•	Favourable structure (shear zones and breccia zones);
	•	Significant quartz vein material;
	•	Hydrothermal alteration minerals and assemblages;
	•	Proximity to unconformities and disconformities; and
	•	Proximity to oxidation/reduction boundaries of regional scale.

On surface, favourable structures are identified utilizing the 2009 LiDAR survey and the 2011 airborne magnetometer survey. The LiDAR survey products include a high-resolution digital elevation model which has been used to map geological contacts, bedding planes, faults, shears, lineations and joints. Follow-up ground geological mapping is employed to identify fabrics, offsets and abrupt changes in rock types that indicate structure.

Mineral prospecting is used to identify indicative mineral alteration, particularly sericite and carbonate minerals and mineral assemblages. Many surface targets meeting some or all of the relevant criteria remain to be tested by drilling.

Underground exploration has been and continues to be guided by drilling mineralized structures along strike and up-dip and down-dip from mine workings and development, particularly within the SAM unit, Shoreline Basalt unit, Intermediate Volcanic Unit, and the 710 host mafic unit. At least seven underground targets have been identified by Klondex for exploration drilling.

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10.

Drilling and Sampling Methodology

Drilling at True North has been completed both on surface and underground. The majority of the drilling was done previously by San Gold. Klondex commenced drilling underground and sampling the historic tailings in the spring of 2016.

10.1.

Diamond Core Drilling

Underground drillholes are planned by the Geology Department using three- dimensional A mine software applying length, dip, and anticipated deviation. The front and back sights are setup by the survey department and on completion of the hole, the collar location is surveyed.

Downhole survey measurements are taken at 70 feet (20m) from the collar, and then every 100 feet (30m) for underground drillholes and 200 feet (60m) for surface drillholes. For infill drilling, the typical hole spacing is 50 feet (15m).

Underground air diamond drills produce AQ size core and underground electric diamond drills produce BQ and NQ size core. Surface diamond drills produce NQ size core, except for the first 500 feet (150m) of some of the deeper holes, for which HQ size core is produced to minimize drillhole deviation.

Underground exploration and definition drilling completed by previous owners along the SAM unit was for near and mid-term production planning purposes. Definition drilling also continued at the 007 and L10 zones within the Shoreline Basalt unit and at the L13 and Hinge Zones in order to advance Mineral Resources from Inferred to Measured and Indicated categories and to increase the Mineral Reserve.

In 2011, the L08 zone was discovered to the northwest of the Hinge Zone. The L08 zone was traced vertically in drilling from 1,000 feet to 2,300 feet below surface and is in close proximity to the Hinge Zone and True North workings at the 16 Level. In 2013, the 710 Zone was discovered as part of underground drilling of an unnamed intermediate mafic volcanic unit between the SAM unit to the south and the Shoreline Basalt unit to the north. Underground drilling of the 710 Zone and other zones continued in 2014 through the spring of 2015.

Klondex commenced underground diamond drilling which, in early 2016, was focused on the 710 Zone. By January 5, 2017 approximately 5,760 underground exploration holes were drilled collectively by San Gold and Klondex, for an approximate total of 2,516,000 feet (767,200 m).

It is important to note that the apparent drop in the Mineral Reserves and Resources from early publicly released numbers (not NI 43-101 compliant) is not due to poor drill results, but due to a more conservative approach to resource modelling and categorizing.

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Surface diamond drilling has occurred at the Project since 1912, which includes more than 2,190 holes and 2,605,000 feet (794,000 m) (Table 10-1). San Gold’s largest surface drilling exploration program was in 2011-2012 and included drilling approximately 1,024,000 feet (336,000 m) in 602 surface holes. The exploration drill program focused on the SAM unit, Shoreline Basalt unit, and Intermediate Volcanic Rock unit.

Table 10-1 Summary of Surface Exploration on the True North Mine

Year	Company	Property	Type of Work	Holes Drilled	Footage
1912	B. Thordarson	Original Sannorm	discovery by prospecting
1934	Normandy Mines Ltd.	Original Sannorm	prospecting, drilling	12	3,000
1945	Sannorm Mines Limited	Original Sannorm	magnetometer survey
1946	Sannorm Mines Limited	Original Sannorm	diamond drilling	37	20,000
1947	Sannorm Mines Limited	Original Sannorm	25' shaft; surface facilities
1949	Sannorm Mines Limited	Original Sannorm	diamond drilling	11	3,923
1961	Sannorm Mines Limited	Original Sannorm	magnetometer survey
1974	Wynne Gold Mines Ltd.	Original Sannorm	diamond drilling	5	3,923
1978	Wynne Gold Mines Ltd.	Original Sannorm	diamond drilling	3	2,177
1985	Orenda Resources Ltd.	Original Sannorm	magnetometer survey
1986	Orenda Resources Ltd.	Original Sannorm	mapping; diamond drilling	7	1,803
1987	Orenda Resources Ltd.	Original Sannorm	VLF EM; IP; diamond drilling	10	2,803
1988	Bakra Resources	Original Sannorm	diamond drilling	8	2,999
1989	Bakra Resources	Original Sannorm	diamond drilling	12	4,292
1992	Partnership	Original Sannorm	diamond drilling	12	5,429
1993	Partnership	Original Sannorm	diamond drilling	4	1,000
1994	Partnership	Original Sannorm	diamond drilling	27	6,859
1996	Partnership	Original Sannorm	diamond drilling	22	4,927
1997	Harmony Inc.	Original Sannorm	diamond drilling	12	6,988
1998	Harmony Inc.	Original Sannorm	diamond drilling	33	28,411
2000	Harmony Inc.	Original Sannorm	RC drilling
2003	Harmony Inc.	Original Sannorm	diamond drilling	17	11,496
2004	Rice Lake Joint Venture	Incl. Mine Lease	diamond drilling	47	28,347
2005	San Gold Corporation	Incl. Mine Lease	diamond drilling	101	67,094
2006	San Gold Corporation	Incl. Mine Lease	drilling on Mine Lease	152	160,276
2007	San Gold Corporation	Incl. Mine Lease	drilling on Mine Lease	186	147,333
2008	San Gold Corporation	Incl. Mine Lease	drilling on Mine Lease	191	191,808
2009	San Gold Corporation	Incl. Mine Lease	drilling on Mine Lease; LiDAR	161	192,474
2010	San Gold Corporation	Incl. Mine Lease	drilling on Mine Lease	352	367,688
2011	San Gold Corporation	Incl. Mine Lease Cougar Option	drilling on Mine Lease; AirMag diamond drilling	382 3	585,931 3,266
2012	San Gold Corporation	Incl. Mine Lease Cougar Option	drilling on Mine Lease diamond drilling	188 3	385,816 5,800

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		Wildcat Option	diamond drilling	26	43,165
2013	San Gold Corporation	Incl. Mine Lease	diamond drilling	170	316,901
2014	San Gold Corporation	Incl. Mine Lease	diamond drilling	0
2015	San Gold Corporation	Incl. mine lease	diamond drilling	0
2016	Klondex	Incl. Mine Lease	diamond drilling	0
			TOTAL	2194	2,605,929

10.2.

Channel Chip Sampling

The face sampling procedure described below is used for grade control at the Project and is derived from a 2016 document provided by Klondex. The sampling procedure consists of the following steps:

•	Ensure that the face to be sampled is secure and safe prior to chipping;
•	Wash the face thoroughly to remove loose material and expose rock type boundaries;
•	Delineate sample intervals using a measuring stick and paint. The sample line should be perpendicular to the orebody dip and run from left to right;
•	Sample widths should not exceed 3 feet in waste and 2 feet in ore, with a minimum sample width of 0.5 foot (0.15m);
•	Sample intervals should not cross lithologic boundaries; and
•	Cumulative sample intervals should encompass the entire width of the face.

Samples are collected by chipping rock from the interval utilizing a sharp rock hammer pick. Rock chips are placed in a clean plastic 12 inch x 12 inch sample bag. The bag is filled to an approximate ¼ capacity. Sample information is recorded on a sample tag and the numbered end placed in the corresponding sample bag.

On a chip sample face sheet, a scaled picture is drawn of the face and the sampled intervals. In addition, the date, sampler, heading, dimension, orientation, position relative to the nearest survey station, samples, sampled intervals, and materials (host rock or vein) are all recorded.

The channel chip sample information is entered into the database and stored as pseudo drillholes with collar, survey and assay values.

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10.3.

Historic Tailings Pond Drilling/Sampling

Prior to committing to a trial of the re-processing of the historic tailings, an extensive sampling program was undertaken to determine the distribution of recoverable gold within the tailings mass. In the spring and summer of 2016, Klondex drilled 138 holes and hand dug 214 holes on the historic tailings for a total of 352 holes with a total footage of 3,714 feet (1,132 m) (Table 10.2) . The holes range in depth from <5 feet at the margins up to 35 feet in the centre.

On the historic tailings, two phases of drilling have been completed using a 200-foot by 100-foot grid. Phase 1 drilling was completed using a stem auger. The hollow stem auger sampling method employed by SanGold is not well documented. Brief descriptions indicate that the holes were drilled and sampled in 5-ft (1.5m) increments through a hollow stem auger.

Phase 2 drilling was performed with a percussion probe (Geo Probe). Holes were planned, located, and staked with a differential global positioning system (GPS), and drilled with a dual tube system. An outer tube is pushed down inside the drill rod by percussion, and then an inner tube with a sampling polyvinyl chloride (PVC) cylinder is pushed down inside of the tube. A continuous 5¬foot sample is collected within the PVC cylinder. The inner PVC cylinder and the sample are retrieved through the outer tube, which remains in place. When drilling resumes, the drill pushes down an additional 5-foot section of outer tube, and the processes is repeated in continuous 5- foot intervals until the total depth of the hole is reached.

In addition to the drilling, samples were also collected from hand dug holes. Holes were planned on a grid spacing of 50-foot by 50-foot and staked and located with a differential GPS. The holes were dug to the diameter of the shovel blade and to a depth at which the hole is stable. The geologist logs lithologic units that are >0.5 foot in thickness, based on grain size (sand, clay, or mixed). Hand dug hole collars are imported into the database as pseudo drillholes with an azimuth of 0 degrees and dip of -90 degrees.

Table 10-2 Summary of Tailings Drilling at True North Mine

Hole Type	Grid Spacing	No. Holes	Footage	Assay Lab
Stem Auger	200 ft X 100 ft	39	818.8	AcmeLabs
Geo Probe	200 ft X 100 ft	99	2351.5	TSL Lab
Hand Dug	50 ft X 50 ft	214	543.5	Site Lab

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11.

Sample Preparation, Analysis, and Security

This section of the report summarizes the sampling methods, sample preparation, assay analysis, and security procedures for surface and underground drill core, underground face sampling, and historic tailings sampling.

For core and face sampling, the procedures developed and documented by the previous operator San Gold have largely been adopted by Klondex. Any changes made by Klondex are noted within this Technical Report. The procedures for tailings sampling were developed entirely by Klondex.

11.1.

Core Sampling Methods

Surface and underground drilling at the Project is completed by contractors. Diamond drill core is placed in labelled wooden trays and depth marker blocks are inserted by drilling contractor personnel prior to the removal of the core from the drill site by the project geologist. Upon arrival at the secure core logging facility, the core boxes are sequentially placed in a core rack and the spatial information on each box of core is checked for accuracy and consistency. If necessary, remedial action is undertaken to correct deficiencies and errors in the spatial information prior to entry into the database. The drill core is digitally photographed prior to logging and marked for sampling.

11.1.1.

Surface Core Sampling Methods

Exploration geologists log the core and record observations in a digital drill log database prior to sample selection for assay analyses (Figure 11.1) . Core intervals are selected for sampling based on the following presence of mineralization, favourable structure and quartz veining. They are then marked and measured for sampling and identified with one part of a three-part assay tag placed at the end of the sample interval.

Samples are taken by sawing the core perpendicular to the core axis, with one-half of the core returned to the core box and the other half placed in a clean plastic bag along with part two of the three-part assay tag. Information on the third part of the assay tag is entered into the database and the drill log, at which time accuracy and consistency are checked again and corrected for discrepancies.

San Gold submitted core samples for assay analysis to TSL Laboratories Inc. in Saskatoon, Saskatchewan. Check assays were performed at Accurassay Laboratories Ltd. in Thunder Bay, Ontario. As of the publication date of this Technical Report, Klondex has not completed any surface drilling. Both labs are independent of Klondex.

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11.1.1.

Underground Core Sampling Methods

Drill programs planned by the Project’s Department are typically underground definition drilling of known zones rather than exploration. The core sampling method differs from that for the surface exploration holes.

The interval to be sampled is determined and marked by the geologist logging the core (Figure 11.1) . Most samples, particularly those from known zones, range between 0.5 foot (0.15m) and 4.0 foot (1.2m) in length. Every sample is bracketed by a minimum of 1.0 ft. (300 mm) for small veins and structures and 6 feet (1.8m) in each of the footwall and hanging wall of known zones.

The entire core sample is placed in a bag by the geologist and identified with an assay tag, which has a copy that remains in the sample book, and the sample number is recorded in the database. If core is to be cut, the sampling procedure is the same as the surface exploration procedure. Approximately 10 feet (3.05m) of core above and below the sampled portion is kept to ensure that sufficient material remains if a re-bracket is required. The remainder of the core is stored at the Project.

Underground core samples are submitted to TSL Laboratories Inc. (TSL) in Saskatoon, Saskatchewan. The check assay laboratory was ALS Global (ALS) in Vancouver, British Columbia. San Gold also submitted core samples to the Project’s Assay Lab, in which case check assays were performed by TSL. Klondex submits underground core samples to TSL. Their check assay lab is ALS. Both labs are independent of Klondex.

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Figure 11-1 Flow Chart for Surface and Underground Core Sampling Methods

In general, all sections with quartz veining and/or alteration are sampled. Sample lengths in mineralized core, characterized by silicification, carbonate alteration, sulphide minerals, quartz veins and visible gold, are variable and based on geological considerations.

Blind standards are routinely inserted into the sample sequence prior to delivery to the assay laboratory. Blanks (also routinely inserted every 50 samples and after all noted visible gold) consist of intervals of un-mineralized core which are identified and flagged prior to shipment to the assay lab.

Sealed sample bags are transported to the assay laboratory in a timely manner. Upon arrival at the assay lab, samples are received by laboratory personnel and transferred to the laboratory’s chain of custody procedures and protocols. (Klondex keeps a chain of custody as well which is updated throughout the process).

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11.2.

Face Sampling Methods

The face sample preparation methodology is outlined below:

•	Channel chip samples are bagged on-site at the face as described in Section 10.2;
•	The samples are delivered to the Project’s assay lab for analysis; and
•	The internal lab inserts their own QAQC into the chip sample stream.

11.3.

Tailings Sampling Methods

Preparation of the tailings drillhole samples is summarized in the following text. The tailings samples and QAQC samples are placed in rice bags and sealed with security tags and zap-straps. Each sample bag has a sample ID and a bag number written on it. There is one sample shipment ID per hole. Each rice bag weighs approximately 30 pounds.

Security tags are recorded in an Excel tracking spreadsheet separate from the Chain of Custody. Phase 1 samples were sent to Acme Analytical Labs Ltd (“Acme”) in Vancouver, British Columbia for assay. Phase 2 samples went to TSL.

The assay lab breaks open, dries and screens the sample at 80 mesh. Approximately 1,000 grams are riffle split, pulverized to -200 mesh. Thirty grams are taken for fire assay. The reject and pulp portions are returned to the Project and stored in the QAQC compound at the SG1 Zone compressor building or retained at the TSL.

Preparation of the hand dug hole samples from the tailings pond is as follows. The samples are delivered to the Project’s Assay Lab where the lab inserts their own QAQC into the tailings sample stream. The samples then are dried and subsequently rolled with a steel rolling pin to break down any lumps. The samples are then split to size in a Jones riffle. After reducing the size of the sample to between 200 and 250 grams, they are pulverized for 30 seconds in a ring pulverizer, rolled and placed in a pulp bag for assaying. Forty grams are taken for fusion.

11.4.

Sample Quality, Representativeness and Sample Bias

• The sampling methods used by Klondex are similar to the current industry standards for mineralization of this type. P&E recognizes that there exist serious issues related to gold deposit sampling because of the non-uniform distribution of gold within the vein systems. P&E does not believe this to be the case in this instance.

• P&E consider that the sampling methods utilized meet NI43-101 standards.

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11.5.

Sample Preparation, Analysis and Security

11.5.1.

Core Sample Preparation and Analysis

The primary independent assay laboratory used by San Gold and Klondex is TSL. When pulps and rejects are returned by TSL, selected samples are sent by the Company to ALS to cross check the TSL assay results. TSL and ALS are each ISO/IEC 17025 certified laboratories and have long histories within the Canadian mining industry. Each laboratory uses similar sample preparation, analytical methods, and QAQC procedures.

On receipt by TSL, samples are sorted and verified according to the sample submittal form shipped with the samples by the Company. Security ties on the sample bags are checked with records sent electronically to TSL and the shipment is assigned a TSL reference number and worksheet. Sample labels are produced with the client sample number and the TSL reference number. Sample preparation procedures involve oscillating jaw crushing to 75% -10 mesh. A 1,000-gram sub-sample is riffle split from the -10 mesh sample and pulverized to >95% -150 mesh in a ring mill pulverizer. Between each sample, the crushers, rifflers, and pans are cleaned with compressed air. Pulverizing pots and rings are brushed, hand cleaned and air blown.

Samples without visible gold are subject to normal fire assay analytical procedures. The gold concentration is determined for a homogenized 30-gram sample using a fire assay collector and atomic absorption finish. Samples are assayed in batches of 24, comprised of 20 client samples, two duplicate client samples, one TSL standard and one TSL blank.

Each sample with visible gold is subject to total metallic and fire assay procedures. The whole sample is crushed and pulverized to 95% passing 150 mesh. The +150 mesh fraction (including the sieve cloth) is assayed for the coarse gold content and two 30-gram samples of the -150 mesh are assayed. The weighted average of the three assays determines the reported assay grade for the sample.

11.5.2.

Channel Chip Sample Preparation and Analysis

Channel chip samples are analyzed by the Project’s Assay Lab at True North. When chip samples are received in the laboratory, they are sorted and placed into numerical order and the sample tag number written on the outside of the plastic bag. Wet samples are dried. All information on the samples received is entered into the logbook.

All dry samples are put through a Rhino Crusher. The crusher reduces the size of the sample to 50% -10 mesh. Crushed samples are reduced in size to approximately 200 grams by splitting, utilizing the sample riffle until one side contains the 200 gram sample to be pulverized. The remaining sample (reject) is returned to the original sample bag and stored for 6 months.

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All crushed samples are pulverized for 90 seconds in a ring pulverizer to 90% -150 mesh. The sample pulps are rolled to ensure that they are homogenous and then placed in pulp bags for assay.

Pulp samples are subject to normal lead fire assay analytical procedures. A 20 gram sample is placed with 60 millilitres flux in a 30 gram crucible for fusion and fire assay of gold. Klondex does not insert standards or blanks with the chip samples.

The past production of greater than 1.5 million oz of gold from the True North Gold Mine supports the validity of the channel sampling and assay procedures. P&E consider the assays to be accurate. The Project’s Assay Lab inserts a standard with every set of samples and the results are checked and tracked internally. The Project’s Assay Lab also runs check assays with each batch of chip samples.

11.5.3.

Tailings Sample Preparation and Analysis

The sample cylinders are split by the driller for the geologist. The geologist wearing latex gloves photographs the sample, and then scoops it from the cylinder into a doubled sample bag using a spoon. The sample is tagged, the inner bag is rolled down and the outer bag is sealed. After each hole is completed, the geologist changes gloves and washes the spoon with distilled water.

11.6.

Core Quality Assurance and Quality Control

A QA/QC program was implemented by San Gold and adopted by Klondex to monitor the contamination, precision and accuracy at the various stages of core sample analysis. Klondex systematically inserts sample standards, blanks and duplicates into its sampling stream.

After every 25th sample, the company inserts a QA/QC control sample alternating between a standard, a field duplicate and a blank. (Standards are every 25th sample, Blanks are every 50th sample or after any noted visual gold, Duplicates are inserted every 20 samples). When assays are received, the data are plotted to ensure that all the results are within acceptable limits and any remediation, if required, is carried out.

11.6.1.

Sample Standards

Under Klondex procedures all exploration core is subject to data verification procedures through the insertion of four blind sample standards at regular intervals in every one hundred samples.

Standards consist of Standard Reference Material (SRM) purchased from CDN Resource Laboratory Ltd. located in British Columbia, Canada. Four different standards are employed with contents of gold ranging from low grade to high grade (Table 11.1) .

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Assay results for the standards are illustrated in Figures 11.2 to 11.5. Results are routinely reviewed. If the results plot outside the acceptable limits for standards or blanks, the sample batch is rerun.

Table 11-1 Certified Gold Assay Values for Commercial Standards

CRM Standard	Laboratory	Recommended Value (g/t)	2 x StdDev
CDN-GS-1P5C	CDN Resource Laboratory Ltd.	1.56	0.13
CDN-GS-6B	CDN Resource Laboratory Ltd.	6.45	0.33
CDN-GS-13A	CDN Resource Laboratory Ltd.	13.20	0.72
CDN-GS-22	CDN Resource Laboratory Ltd.	22.94	1.12

P&E considers that all potential gold mineralized zones in drill core have been sampled. Security of the samples at the core logging facility and at the analytical lab appear to be adequate to ensure the integrity of the samples and assays.

Figure 11-2 Assay Results of Standard CDN-GS-1P5C

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Figure 11-3 Assay Results for Standard CDN-GS-6B

Figure 11-4 Assay Results for Standard CDN-GS-13A

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Figure 11-5 Assay Results for Standard CDN-GS-22

11.6.2.

Core Sample Blanks and Duplicates

Sample blanks consist of un-mineralized, unaltered and un-deformed drill core from True North. Two blanks are inserted at regular intervals for every 100 samples. Additional blanks are inserted after each sample with visible gold. The blanks are employed to monitor contamination during the sample preparation step in the assay lab.

Review of assay results for 4,499 blanks indicates that only 28 (<0.7%) exceed the upper threshold assay value set by San Gold of 0.05 opt Au (Figure 11.6) . An upper threshold assay value of 0.02 opt Au is utilized by Klondex. Otherwise, the core sample blank analysis protocols remain unchanged from those used previously by San Gold.

Five duplicate samples are inserted by TSL at regular intervals in every 100 samples by cutting un-mineralized intervals of whole drill core in half. The procedure also includes TSL submitting selected sample rejects to ALS for duplicate analysis. The duplicate sample assay results reflect the heterogeneous distribution of gold in the core (Figure 11.7) .

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Figure 11-6 Assay Results for Blanks

Figure 11-7 Assay Results for Duplicates

11.7.

Chip Quality Assurance and Quality Control

11.7.1.

Sample Standards

The standards inserted by the Project’s Assay Lab consist of Certified Reference Materials (SRM) produced by Rocklabs of Auchland, New Zealand and sold in Canada previously by

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Anachemia Science and currently by VMR’s mining division (Table 11.2) . The standards are ordered with a sulphur matrix and a gold value of approximately 1 part per million (ppm) to 10 ppm.

Results are routinely reviewed by the Project’s Assay Lab.

Table 11-2 Certified Gold Assay Values for Commercial Standards

CRM Standard	Laboratory	Recommended Value (ppm)	95% Cl
SG31	Rocklabs	0.996	0.011
SG40	Rocklabs	0.976	0.009
SG66	Rocklabs	1.086	0.009
SK52	Rocklabs	4.107	0.029
SK62	Rocklabs	4.075	0.045
SL61	Rocklabs	5.931	0.057
SN38	Rocklabs	8.576	0.061
SN50	Rocklabs	8.685	0.062
SN60	Rocklabs	8.595	0.073
SN74	Rocklabs	8.981	0.065
SN75	Rocklabs	8.671	0.054

11.7.2.

Chip Sample Duplicates

Duplicate assays are run with every set of chip samples. Assay results for the original samples versus duplicates are illustrated in Figure 11.8. Results are routinely reviewed by the Project’s Assay Lab.

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Figure 11-8 Assay Results for Chip Sample Duplicates

11.8.

Tailings Quality Assurance-Quality Control

11.8.1.

Sample Standards

For QAQC of the historic tailings samples, blind standards are inserted every 25th sample. The standards utilized are Canadian Research Lab standards GS-1L, GS-1P5C, and GS-P6 on an alternating basis. At this stage, the majority of the data are available for GS-1P5C and GS-1L. Assay results for the standards are illustrated in Figures 11.9 to 11.10. Results are routinely reviewed. If the results plot outside the acceptable limits for standards and for blanks, the sample batch is rerun.

11.8.2.

Tailings Sample Blanks and Duplicates

Sand blanks are inserted at the sample numbers ending in 40 and 90. The blanks are composed of beach sand, which has been prepared and assay tested at the Project’s Assay Lab. Assay results for the blank are illustrated in Figure 11.11. Assay results for duplicates are illustrated in Figure 11.12. Blank and duplicate results are routinely reviewed.

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Figure 11-9 Tailings Assay Results for Standard GS-1P5C

Figure 11-10 Tailings Assay Results for Standard GS-1L

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Figure 11-11 Tailings Assay Results for Blanks

Figure 11-12 Tailings Assay Results for Duplicates

11.9.

Recommendations and Conclusions

P&E is of the opinion that the core, channel chip and tailings sample assay data have been adequately verified for the purposes of a mineral resource estimate. All data included in the resource estimate appear to be of adequate quality.

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Recommendations are as follows:

1) Technical Database: All Project data collected needs to be stored and archived in a permanent and reliable retrieval manner. A full-time database administrator is recommended.

2) Quality Assurance/Quality Control: Timely follow-up for any and all QA/QC assay deviations and re-assay requests should be performed in a timely manner. The process should be automated when the database is up and running.

3) Sample Storage and Retrieval: Half-core remaining from sample assays should be retained for reference and check assay purposes. All assay sample rejects and pulps should be stored in a safe, secure and sheltered manner and properly catalogued to ease retrieval.

4) Project Assay Lab: Standard operating procedures should be updated, particularly in regards to assay data generation, storage and retrieval.

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12.

Data Verification

This section of the report summarizes the results of P&E’s due diligence for the data verification for the True North Gold Mine.

12.1.

Drill Data Review

True North Gold Mine was visited by Dr. Bill Stone, P.Geo., a Qualified Person as defined in National Instrument NI 43-101 Standards of Disclosure for Mineral Projects. Dr. Stone visited the Project from May 21-25, 2016; June 20-23, 2016; September 20-22, 2016; and January 16-19, 2017. The categories of data reviewed for the three datasets include collar location surveys, down-hole surveys, assay results and geology.

12.1.1.

Collar Location Checks

Comparison of 209 underground collar survey reports to easting, northing, elevation and hole length values in the database revealed only a single error. Collar locations of the underground holes are considered to be reliable.

Comparison of 55 surface collar survey reports to easting, northing, elevation and hole length values in the database revealed no errors. Collar locations for the surface holes are considered to be reliable.

12.1.2.

Hole Survey Checks

Comparison of 120 underground hole borehole survey reports to azimuth and dip values in the database found only a few errors. Borehole surveys of the underground holes are considered to be reliable.

Comparison of 13 surface hole borehole survey reports to azimuth and dip values in the database found no errors.

12.1.3.

Core Assay Checks

P&E conducted verification of the drillhole assay database by comparison of the database entries with the assay certificates. The assay certificates were obtained in digital format directly from the assay laboratories and compiled.

Assay data ranging from 2009 to June 2016 were verified for the True North Gold Mine. Approximately 40% (6,220 out of 15,494) of the constrained drilling assay data were checked for Au against the original laboratory certificates from TSL, ALS, and the Project’s Assay Lab. A very few minor data errors were observed and corrected, with the overall impact to the database considered negligible. In addition, 63% of the assay samples in the constrained core database for July 2016 to January 2017 were checked against the TSL Lab assay certificates and no errors were found.

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12.1.4.

Geology Checks

The digital core logging system developed by San Gold has been adapted by Klondex for True North. In the lithology database, approximately 15,000 intervals logged as 75% or greater quartz were checked for vein codes. As a result, only <1% of the quartz rich intervals were coded as rock types other than quartz veins.

Many of these intervals are coded as chert or cherty sediments, or structures such as faults and shear zones which are commonly associated with quartz veins. In addition, approximately 7000 intervals logged as shears or shear zones in the structure database were also checked and only a single error was found.

12.2.

Channel Chip Data Review

12.2.1.

Collar Location Checks

The location of the channels on the original face map sheets and the Klondex database show positive correlation.

12.2.2.

Downhole Survey Check

The location of the channel chip samples on the face map sheets and the database show positive correlation.

12.2.3.

Assay Check

P&E conducted verification of the channel chip database by comparison of the database entries with the assay certificates. The assay certificates were obtained in digital format directly from the assay laboratories and compiled.

Assay data ranging from 2009 through June 2016 were verified for the True North Gold Mine. Approximately 66% (3,739 out of 5,697) of the constrained chip channel assay data were checked for Au against the original laboratory certificates from the Project’s Assay Lab. A very few minor data errors were observed and corrected, with the overall impact to the database considered negligible. In addition, 1797 of 1888 assay samples in the constrained chip channel assay database for July 2016 to January 2017 were checked against the Site Assay Lab certificates and only a few errors were found, with the overall impact to the database considered negligible.

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12.3.

Tailings Data Review

Tailings sample procedures are described in Section 10.3 of this report. All the holes are located by differential GPS and are vertical. Sand and clay lithological units >0.5 feet thick are logged.

12.3.1.

Assay Checks

P&E conducted verification of the tailings assay databases by comparison of the database entries with the assay certificates. The assay certificates were obtained in digital format directly from the assay laboratories and compiled. Approximately 87% (882 out of 1,012) of the tailings assay data were checked for Au against the original laboratory certificates from TSL, ACME and the Project’s Assay Lab. A very few minor data errors were observed and corrected, with the overall impact to the database considered negligible.

12.4.

Due Diligence Sampling

Data verification assays were carried out for four separate batches of samples:

•	Pulp samples of diamond core from 14 drillholes of three mineralized zones;
•	Half-core samples of diamond core from 6 drillholes of three mineralized zones;
•	Rejects of 20 samples from the tailings pond; and
•	Twenty reject samples of diamond core from 5 holes drilled by Klondex in 2016.

As part of the due diligence, sampled intervals of a variety of low grade, average grade, and high grade mineralized material. Selected core intervals were sampled by taking the entire pulp or reject sample or the entire half core, whichever was available.

For due diligence of the historic tailings, samples were selected from stored sample bags. The bagged materials were sampled by taking half or the entire sample, whichever was available. Prior to sampling, employees or other associates of Klondex were not informed of the location or identification of any of the samples to be collected. The objective of these check samples was to verify the presence and approximate grades of gold encountered during drilling, rather than to provide exhaustive independent verification of the original assay results.

The samples were collected by Dr. Stone or under his direct supervision. They were placed by Dr. Stone in appropriately numbered sample bags with pre-packaged standards, sealed in rice bags with lock ties and packing tape, and taken by him to Burlington for courier transport to the P&E office in Brampton, Ontario. From there, the samples were sent by courier to AGAT Laboratories Ltd. (AGAT) in Mississauga, Ontario for analysis. Gold was analyzed by fire assay on a 30 gram aliquot with an AAS finish. Samples yielding values >10 g/t Au were re-assayed and quantitatively determined by the gravimetric method.

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AGAT employs a quality assurance system to ensure the precision, accuracy and reliability of all results. The best practices have been documented and are consistent with:

	•	The International Organization for Standardization’s ISO/IEC 17025, “General Requirements for the Competence of Testing and Calibration Laboratories' and the ISO 9000 series of Quality Management standards”;
	•	All principles of Total Quality Management (TQM);
	•	All applicable safety, environmental and legal regulations and guidelines;
	•	Methodologies published by the American Society for Testing and Materials (ASTM), National Institute for Occupations Safety and Health (NIOSH), United States Environmental Protection Agency (EPA) and other reputable organizations, and;
	•	The best practices of other industry leaders.

Figure 12-1 Due Diligence Sample Pulp Results for Gold

P&E’s independent comparisons of the pulp, core and reject sample verification results to the original assay results are illustrated in Figure 12-1, Figure 12-2, and Figure 12-3 respectively. The P&E results for the pulps and the rejects are satisfactory, but those for cores exhibit a low bias consistent with what could be anticipated for different sample volumes of a relatively heterogeneous sample from a high-grade, narrow vein gold deposit like True North.

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Figure 12-2 Due Diligence Sample Core Results for Gold

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Figure 12-3 Due Diligence Core Reject Sample Assays

Comparisons of the P&E independent tailings sample verification results to the original assay results are illustrated in Figure 12-4.

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Figure 12-4 Due Diligence Tailings Sample Results for Gold

12.5.

Conclusions to Data Verification

Based on the evaluation of the QA/QC program undertaken by Klondex and the due diligence sampling and assay program performed by P&E, it is P&E’s opinion that the results are suitable for use in the current Mineral Resource Estimate.

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13.

Mineral Processing and Metallurgical Testing

Mining and ore processing has been carried out at True North intermittently since the early 1930’s. The original process used a gravity concentration step and whole-ore cyanidation using Merrill Crowe gold precipitation. Recoveries with this original plant and process were generally 96%.

In 1980, the original process plant was destroyed by fire except for the crushing plant and fine ore bin feed conveyor. A new process plant was constructed with the same throughput as the original; however, the process was changed to incorporate gravity concentration and a bulk sulphide flotation process.

It was found that by floating the sulphides, a “throwaway tail” could be achieved. The concentrate was reground and upgraded through a cleaner circuit and filtered. The high grade concentrate was then shipped to a local smelter. Recovery using this process was generally 93%.

In the mid-1990’s, the mine was restarted and the process plant was expanded by adding a larger 12.5 foot x 14 foot (3.8m x 4.3m) grinding mill and a cyanide leach circuit for concentrate leaching. The operation was short lived.

In 1998, the operation was restarted again and this time ran for three years at a rate of 1,000 tons (907 tonnes) per day. The process used two-stage crushing followed by grinding, concentration using a centrifugal concentrator, and a bulk sulphide flotation process. This flotation concentrate was reground and sent to a leach/CIP gold recovery plant. The carbon was eluted using a conventional pressure strip followed by electrowinning and subsequent refining. Recovery for the period was calculated as generally 92%.

Table 13-1 Harmony Gold – Rice Lake Deposit Metallurgical Results

Tons Milled 1990’s	Gravity (oz. Au)	EW (oz. Au)	Gold Prod’n (oz. Au)	Overall Loss (oz. Au)	Calc. Grade (opt Au)	Gravity Recovery	Overall Recovery
994,830	58,198	91,297	149,496	13,304	0.164	35.75%	91.83%

When the Hinge Zone was developed, a 3,700 ton (3,357 tonne) bulk sample was treated through the process circuit with no changes having been made to that process. Recovery from this bulk sample was generally 92%. Subsequent samples were processed in May and June of 2009 with recoveries at 96.6% and 97.2% respectively, not shown.

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Table 13-2 Hinge Zone Metallurgical results

Tons Milled	Gravity (oz. Au)	EW (oz. Au)	Gold Prod’n (oz. Au)	Overall Loss (oz. Au)	Calc. Grade (opt Au)	Gravity Recovery	Overall Recovery
154,229	6,712	16,608	23,320	1,826	.163	27.30%	92.74%
27,543	742	3,385	4,127	348	.162	16.59%	92.23%
258,469	10,462	21,418	31,880	2,605	.133	27.61%	92.45%

In August 2010, the first bulk sample from the 007 Zone ore was processed. This represented 6,245 tons (5,667 tonnes) grading 0.139 opt Au (4.77 g/t Au) gold with a general recovery of 92%. Additional samples in the months of September and October of 2010 yielded recoveries between 95% and 92%, not shown. Current process plant recovery from all ore is 93.3%, not shown.

Table 13-3 007 Zone Metallurgical Results

Tons Processed	Gravity (oz. Au)	EW (oz. Au)	Gold Prod’n (oz. Au)	Overall Loss (oz. Au)	Calc. Grade (opt Au)	Gravity Recovery	Overall Recovery
24,734	1,015	1,944	2,959	270	0.131	65.78%	91.65%
248,475	17,782	27,716	45,498	3,026	0.195	36.65%	93.76%

Although current operations employ a conventional ball mill as a primary grinding unit, the potential of Semi-Autogenous Grinding (SAG) milling was investigated. Samples of both True North and Hinge Zone mineralized material were sent to both SGS Mineral Services’ Lakefield Laboratory (SGS Lakefield) and Starkey & Associates Inc. (Starkey Associates) for testing. Results are listed below.

Table 13-4 SGS Lakefield And Starkey Associates Sag Mill Testing Results

Sample Name	Relative Density	JK Parameters		MacPherson Test		Work Indices (kWh/t)
Sample Name	Relative Density	A x b	ta	(kg/h)	(kWh/t)	AWI	RWI	BWI
Rice Lake Ore	2.77	74.5	0.34	9.7	8.2	13.9	15.7	14.9
Hinge Ore	2.71	64.4	.038	10.9	7.5	14.5	13.2	16.7

Table 13-5 JKTech Drop-Weight Test Summary

Sample Name	A	b	A x b	Hardness Percentile	ta	Hardness Percentile	Relative Density
Rice Lake Ore	61.7	0.77	47.5	50	0.34	73	2.77
Hinge Ore	91.9	1.04	64.4	30	0.38	65	2.71

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Table 13.6 shows additional SGS Lakefield and Starkey & Associates SAG Mill Testing Results.

Table 13-6 More SGS Lakefield And Starkey & Associates Sag Mill Testing Results

Project Identification		SAG Mill Data from SAG Design Test					Ball Mill Data from SAG Design Test							Total Pinion W to P80 kWg/t
Project Sample No.	Client Sample Info	Initial Weight grams	No. of Revs	Bulk SG g/cc	SG Solids g/cc	Calc SAG W to 1.7mm kWh/t	Initial Weight grams	Test Feed F80μ	Test Produc t F80 μ	Gpb (Avg last 3 cycles)	SAG Dis. Bond BWI kWh/t	Macro/ Micro Ratio	Calc BMW to P80 kWh/t	Total Pinion W to P80 kWg/t
1	Zone 1 - Hinge	7715	1123	1.71	2.71	7.72	1303	1409.7	1163	1.516	16.67	0.46	12.23	19.94
2	Zone 2 Rice	7650	1306	1.70	2.84	9.03	1294	1348.4	112.6	1.705	14.93	0.60	10.95	19.97
Averag		7682	1214	1.71	2.78	8.37	1298	1379.0	114.4	1.610	15.80	0.53	11.59	19.96
Std. deviati on		46	130	0.01	0.09	0.93	7	43.3	2.7	0.134	1.23	0.10	0.90	.002
Design data											16.67	0.54	12.23	21.25
							Bond Equation for Pinion Energy:
		SAGDesign Equation for Pinion Energy:
							W = (10Wi/P80^0.5)fines factor
		W = Revolutions * (grams+16000)/(447.3*grams)
							Note: Calc BM pinion kWh/t is based on P80 105μm
		Note: Calc SAG pinion kWh/t equation calibrated
		for feed
							Fines Factor = (P80 + 10.3)/(1.145P80) 1.00*
		F80 152mm and transfer size T80 170mm
							Note: Bond BM Wi test closing Screen 150 μm

In 2012 several flotation tails samples were leached in cyanide to understand the potential need for building a flotation tails leach circuit at the Project. The samples were tested during two test programs with both programs employing a 24-hour leach on as-received samples at a pH above 10.0. The first program used a cyanide concentration of 2.5 grams per litre cyanide (gpl NaCN) and the second program a concentration of 0.5 gpl NaCN. The results of these two programs are summarized in the tables below:

Table 13-7 Results Leaching Flotation Tails for 24 Hours at 2.5 gpl NaCN Concentration

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Date Sampled	Calc. Head, opt Au	% Recovery	opt Recoverable Gold
10/09/12	0.0100	90.04%	0.0090

10/10/12	0.0062	84.00%	0.0052
10/11/12	0.0074	86.52%	0.0064
10/12/12	0.0071	85.96%	0.0061
10/13/12	0.0100	90.04%	0.0090
10/14/12	0.0086	88.35%	0.0076
10/15/12	0.0119	66.31%	0.0079
10/16/12	0.0157	55.55%	0.0087
10/17/12	0.0071	85.96%	0.0061
10/18/12	0.0081	75.38%	0.0061
10/19/12	0.0083	87.94%	0.0073
10/20/12	0.0092	89.09%	0.0082
10/21/12	0.0095	89.43%	0.0085
10/22/12	0.0077	87.03%	0.0067
10/23/12	0.0077	87.03%	0.0067
10/24/12	0.0083	87.94%	0.0073
10/25/12	0.0105	85.77%	0.0090
10/26/12	0.0271	87.10%	0.0236
10/27/12	0.0103	90.32%	0.0093
10/28/12	0.0095	94.76%	0.0090
10/30/12	0.0071	85.96%	0.0061
10/31/12	0.0088	82.94%	0.0073
11/01/12	0.0065	76.77%	0.0050
11/02/12	0.0097	89.74%	0.0087
Avg.	0.0097	84.58%	0.0081

Table 13-8 Results Leaching Flotation Tails for 24 Hours at 0.5 gpl NaCN Concentration

Date Sampled	Calc. Head, opt Au	% Recovery	opt Recoverable Gold
11/08/12	0.0060	83.22%	0.0050
11/09/12	0.0045	77.78%	0.0035
11/12/12	0.0054	81.39%	0.0044
11/14/12	0.0060	83.22%	0.0050
11/15/12	0.0067	70.00%	0.0047
11/16/12	0.0070	71.25%	0.0050
11/17/12	0.0093	78.47%	0.0073
11/18/12	0.0065	84.71%	0.0055
11/19/12	0.0080	87.50%	0.0070
11/20/12	0.0092	89.09%	0.0082
11/21/12	0.0068	85.36%	0.0058
11/22/12	0.0086	88.35%	0.0076
11/23/12	0.0080	87.50%	0.0070
11/24/12	0.0080	87.50%	0.0070

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Date Sampled	Calc. Head, opt Au	% Recovery	opt Recoverable Gold
11/25/12	0.0092	89.09%	0.0082
11/26/12	0.0092	89.09%	0.0082
11/29/12	0.0118	91.52%	0.0108
11/29/12	0.0210	57.05%	0.0120
11/29/12	0.0117	74.47%	0.0087
11/30/12	0.0176	94.33%	0.0166
12/01/12	0.0060	83.22%	0.0050
12/02/12	0.0085	58.62%	0.0050
12/03/12	0.0068	85.36%	0.0058
12/04/12	0.0089	88.73%	0.0079
12/13/12	0.0071	85.96%	0.0061
12/14/12	0.0071	85.96%	0.0061
12/15/12	0.0054	81.39%	0.0044
12/16/12	0.0097	89.74%	0.0087
12/17/12	0.0100	90.04%	0.0090
12/18/12	0.0089	88.73%	0.0079
12/19/12	0.0092	89.09%	0.0082
12/20/12	0.0083	87.94%	0.0073
12/21/12	0.0080	87.50%	0.0070
12/22/12	0.0086	71.01%	0.0061
12/23/12	0.0146	89.74%	0.0131
12/24/12	0.0124	91.92%	0.0114
12/27/12	0.0074	86.52%	0.0064
12/28/12	0.0065	84.71%	0.0055
12/29/12	0.0074	86.52%	0.0064
12/30/12	0.0098	59.32%	0.0058
12/31/12	0.0067	85.05%	0.0057
Avg.	0.0087	82.91%	0.0072

Solid samples have been collected from the tailings storage facility and leached to document potential recoverable gold present. Three as-received samples were leached for 72 hours with 1.0 gpl NaCN concentration at a pH of above 10.0. Results from these three tests are given in the table below:

Table 13-9 Results from Leaching Samples from Tailings Storage Facility

Sample	Calc. Head, opt Au	% Recovery, 24 hr test	% Recovery, 72 hr
1	0.0408	80.1%	85.4%
2	0.0204	67.7%	72.4%
3	0.0408	78.1%	85.8%

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14.

Mineral Resource Estimates

14.1.

Introduction

The mineral resource estimate presented herein has been prepared following the guidelines of the Canadian Securities Administrators’ National Instrument 43-101 and Form 43-101F1 and in conformity with generally accepted “CIM Estimation of Mineral Resource and Mineral Reserves Best Practices” guidelines¹. Mineral resources have been classified in accordance with the “CIM Standards on Mineral Resources and Reserves: Definition and Guidelines”:

	•	Measured Mineral Resource: “A ‘Measured Mineral Resource’ is that part of a mineral resource for which quantity, grade or quality, densities, shape, and physical characteristics are so well established that they can be estimated with confidence sufficient to allow the appropriate application of technical and economic parameters, to support production planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration, sampling and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough to confirm both geological and grade continuity.”
	•	Indicated Mineral Resource: “An ‘Indicated Mineral Resource’ is that part of a mineral resource for which quantity, grade or quality, densities, shape and physical characteristics, can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes that are spaced closely enough for geological and grade continuity to be reasonably assumed.”
	•	Inferred Mineral Resource: “An ‘Inferred Mineral Resource’ is that part of a mineral resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes.”

^{____________________________________}

¹CIM DEFINITION STANDARDS - For Mineral Resources and Mineral Reserves. Prepared by the CIM Standing Committee on Reserve Definitions. Adopted by CIM Council on May 10, 2014.

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Mineral resources are not mineral reserves and do not have demonstrated economic viability. There is no guarantee that all or any part of the mineral resource will be converted into mineral reserve. Confidence in the estimate of Inferred Mineral Resources is insufficient to allow the meaningful application of technical and economic parameters or to enable an evaluation of economic viability worthy of public disclosure.

All mineral resource estimation work reported herein was carried out or reviewed by Fred Brown, P.Geo., an independent Qualified Person as defined by National Instrument 43-101 by reason of education, affiliation with a professional association and past relevant work experience. Mr. Brown visited the True North site over the period February 20, 2017 to February 24, 2017. This mineral resource estimate is based on information and data supplied by Klondex, and a draft copy of this report was reviewed by Klondex for factual errors.

Mineral resource modelling and estimation were carried out using the Maptek Vulcan software program. The effective date of this mineral resource estimate is February 14, 2017.

14.2.

Previous Resource Estimates

A previous public mineral resource estimate for the True North Mine was reported in 2016.²The 2016 mineral resource estimate was based on 36 modelled veins, and reported 293,400 Measured and Indicated Au ozs and 459,900 Inferred Au ozs at a cut-off grade of 0.090 opt (Table 14-1).

Table 14-1 Total Reported Mineral Resources as of June 30, 2016

Class	Grade Au opt	Grade Au g/t	Tons	Au oz
Measured	0.232	7.95	455,000	105,600
Indicated	0.202	6.92	931,000	187,800
Meas + Ind	0.212	7.26	1,386,000	293,400
Inferred	0.165	5.65	2,793,000	459,900

P&E notes that all previous mineral resource estimates for the True North deposits are superseded by this report.

^{_________________________________}

²Puritch E, Veresezan A, Brown F, Stone W, Hayden A, Orava D, Rodgers K (2016). Technical report and pre-feasibility study on the True North Gold Mine, Bissett, Manitoba, Vanada. Technical report for Klondex Canada Ltd. released on SEDAR with an amended effective date of June 30, 2016.

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14.3.

Data Supplied

The cut-off date for the drilling database used for this updated mineral resource estimate is January 5, 2017. The updated database encompasses records for 345,656 assay samples from 7,877 diamond drillholes and 108,127 assay samples for 29,214 channel strings (Table 14.2). Distance units are reported in feet, and assay grade units are reported as ounce per short ton (opt).

Table 14-2 True North Database Records

Data Type	Record Count	Total Footage
Drillhole Samples	345,656	717,098.5
Channel Samples	108,127	248,022.5
Total Samples	453,783	965,121,0

Industry standard validation checks were completed on the supplied databases, and minor corrections made where necessary. P&E typically validates a mineral resource database by checking for inconsistencies in naming conventions or analytical units, duplicate entries, interval, length or distance values less than or equal to zero, blank or zero-value assay results, out-of-sequence intervals, intervals or distances greater than the reported drill hole length, inappropriate collar locations, and missing interval and coordinate fields. No significant discrepancies with the supplied data were noted, and P&E considers that the databases are suitable for mineral resource estimation.

14.4.

Bulk Density

Klondex supplied a total of 7,586 bulk density measurements, with an average value of 2.76 tonnes per cubic metre (0.086 short tons per cubic foot) (Table 14.3) . Bulk density was determined by immersion of dried core samples in distilled water.

No correlation between assay grade and bulk density was noted by P&E. The average reported bulk density of 2.76 tonnes per cubic metre (0.086 short tons per cubic foot; 11.7 cubic feet per short ton) was used for mineral resource estimation.

Table 14-3 Bulk Density Sample Statistics

	Tonnes per cubic metre	Short tons per cubic foot
Minimum	2.30	0.072
Maximum	4.14	0.129
Average	2.76	0.086
Median	2.77	0.086

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	Tonnes per cubic metre	Short tons per cubic foot
Mode	2.75	0.086
Standard Deviation	0.07	0.002
Number of Samples	7,586

14.5.

Vein Modelling

The updated P&E mineral resource estimate is based on 62 modelled veins, with a total volume on the order of 291 million cubic feet. Of the 62 modelled veins reported herein, 17 modelled veins are unchanged from the 2016 mineral resource, 13 modeled veins have been depleted against the 2016 reported mineral resource, 6 modeled veins have been updated with additional information, and 26 modelled veins have been added to the 2016 mineral resource estimate.

The modelling methodology as implemented by Klondex can be summarized as follows:³

	1.	Set the vein to be modelled, its overall dip and dip direction, and the drillhole and channel databases to be used.
	2.	Extract the hanging wall (HW) and footwall (FW) vein intercepts from the drillhole and channel databases.
	3.	Combine interpreted or surveyed HW and FW points to control the vein model interpretation where required.
	4.	Use the dip and dip direction settings to rotate the intercepts to a local flat plane.
	5.	Use inverse distance to contour HW and FW grid surfaces from the input data and perform grid mathematics to ensure HW grid points are always above FW grid.
	6.	Create a triangulation of the HW contact that combines the grid model points with the input intercepts to ensure the final surface is snapped to the input data. Repeat this process for the FW contact. Model specific settings are attached as attributes to the triangulations and also written to a text file for future auditing. Where channel samples are present, channel sampling may override drillhole sampling in generating the vein model as drillhole intercepts may be found to be locally inaccurate. Drillholes to be ignored are flagged to allow channel samples will take precedence over drilling.

^{___________________________}

³Anthony Bottrill, Klondex Corporate Resources Manager, personal communication, August 16, 2016.

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	7.	Produce boundary polygons of the vein contact surfaces to create a boundary triangulation that can be appended to the vein contacts to create a valid solid triangulation.
	8.	Un-rotate the triangulations and intercepts back to their true spatial location.
	9.	Clip the solid vein triangulation to the topography and other vein surfaces as required.
	10.	Build the vein block model. The block model specifications are read directly from the vein extents and overall dip and dip direction used in the vein model creation process. Block sizes along strike and down dip are set to 5.0 ft. x 5.0 ft. to represent local variations in orientation of the vein. Block sizes across the thickness of the vein are designated as a single block across the true width of the vein to a minimum thickness resolution of 0.2 ft. This ensures the local vein orientations and volumes are representative but also rationalizes the size of the final block model.

The modelling technique as implemented by Klondex produces a series of valid triangulated wireframes oriented to the plane of the vein, as well as a corresponding rotated and plunging block model for each vein. The block model corresponding to the vein has constant strike and dip dimensions of 5.0 ft., and a variable vein thickness perpendicular to strike.

P&E has reviewed the resulting wireframes and considers them to be suitable for mineral resource estimation. The vein wireframes were used for volumetrics, sample coding, statistical analysis and compositing limits (Table 14-4).

Table 14-4 Modelled Veins

Model	Vein	Wireframe Volume (cubic feet)	Wireframe Tons	Model	Vein	Wireframe Volume (cubic feet)	Wireframe Tons
2016	v1011	1,547,795	133,110	2017	v707	1,020,740	87,784
2016	v1012	1,170,010	100,621	2017	v708	3,080,210	264,898
2016	v1040	240,220	20,659	2017	v712	2,498,942	214,909
2016	v1300	4,288,995	368,854	2017	v715	2,061,603	177,298
2016	v1305	1,115,260	95,912	2017	v750	1,714,643	147,459
2016	v1320	1,539,625	132,408	2017	v751	222,287	19,117
2016	v1330	772,756	66,457	2017	v752	1,105,123	95,041
2016	v200	23,078,214	1,984,726	2017	v755	1,102,244	94,793
2016	v210	9,920,635	853,175	2017	v756	375,939	32,331
2016	v63	3,330,105	286,389	2017	v757	310,740	26,724
2016	v731	2,482,877	213,527	2017	v770	1,106,404	95,151
2016	v732	1,115,265	95,913	2017	v98	1,155,709	99,391

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2016	v820	2,103,755	180,923	2017 Depl.	v072	4,388,856	377,442
2016	v86	3,981,750	342,431	2017 Depl.	v1010	14,318,142	1,231,360
2016	vCW2	5,575,025	479,452	2017 Depl.	v1030	11,931,908	1,026,144
2016	vCW3	6,303,789	542,126	2017 Depl.	v1310	2,079,285	178,819
2016	vCW4	9,005,981	774,514	2017 Depl.	v1331	193,962	16,681
2017	v04	8,614,068	740,810	2017 Depl.	v62	695,815	59,840
2017	v100	1,856,915	159,695	2017 Depl.	v700	12,000,717	1,032,062
2017	v101	1,337,231	115,002	2017 Depl.	v730	12,317,510	1,059,306
2017	v500	1,181,478	101,607	2017 Depl.	v800	5,977,515	514,066
2017	v505	1,755,391	150,964	2017 Depl.	v810	7,311,945	628,827
2017	v507	1,312,126	112,843	2017 Depl.	v84	6,724,538	578,310
2017	v510	1,158,255	99,610	2017 Depl.	v91	3,501,150	301,099
2017	v511	1,353,128	116,369	2017 Depl.	vSG1	54,522,427	4,688,929
2017	v512	1,051,374	90,418	2017 Update	v1020	5,975,060	513,855
2017	v513	1,448,757	124,593	2017 Update	v400	7,896,820	679,127
2017	v515	1,179,031	101,397	2017 Update	v710	9,644,615	829,437
2017	v520	1,271,638	109,361	2017 Update	v711	8,030,502	690,623
2017	v522	1,397,845	120,215	2017 Update	v713	2,114,117	181,814
2017	v530	573,778	49,345	2017 Update	v714	2,082,206	179,070

14.6.

Assay Data

Summary assay statistics were calculated separately by vein for drillhole assay samples (Table 14.5) and channel assay samples (Table 14.6). Channel sample strings have been converted to pseudo-drillholes.

Table 14-5 Summary Drillhole Assay Statistics

Vein	Count	Minimum opt	Maximum opt	Mean opt	Std. Deviation	CV
v1011	252	0.0001	4.295	0.180	0.425	2.366
v1012	62	0.0001	2.265	0.252	0.429	1.705
v1040	113	0.001	13.684	0.245	1.294	5.271
v1300	258	0.0001	1.826	0.123	0.191	1.545
v1305	42	0.0001	0.438	0.073	0.082	1.128
v1320	39	0.001	0.450	0.104	0.123	1.185
v1330	16	0.001	2.623	0.282	0.641	2.279
v200	192	0.0001	1.569	0.093	0.210	2.260
v210	112	0.0001	1.988	0.084	0.214	2.545
v63	415	0.0001	2.732	0.174	0.304	1.746

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Vein	Count	Minimum opt	Maximum opt	Mean opt	Std. Deviation	CV
v731	98	0.0001	2.825	0.152	0.414	2.727
v732	127	0.0001	5.229	0.159	0.536	3.364
v820	81	0.001	2.284	0.132	0.332	2.516
v86	338	0.0001	2.220	0.174	0.307	1.763
vCW2	262	0.0001	1.356	0.086	0.169	1.968
vCW3	350	0.0001	6.948	0.149	0.558	3.759
vCW4	331	0.0001	1.632	0.103	0.173	1.683
v04	1198	0.0001	9.283	0.140	0.513	3.664
v100	97	0.001	1.803	0.154	0.329	2.136
v101	155	0.001	0.365	0.102	0.323	3.167
v500	90	0.001	3.630	0.186	0.464	2.495
v505	101	0.001	1.709	0.122	0.202	1.656
v507	63	0.001	2.114	0.113	0.302	2.673
v510	131	0.001	2.400	0.124	0.258	2.081
v511	96	0.001	1.660	0.112	0.244	2.179
v512	68	0.001	2.760	0.151	0.372	2.464
v513	69	0.001	0.306	0.103	0.205	1.990
v515	41	0.001	0.254	0.311	0.551	1.772
v520	21	0.002	1.236	0.137	0.252	1.839
v522	9	0.001	0.790	0.242	0.262	1.083
v530	104	0.001	1.190	0.129	0.158	1.225
v707	16	0.0001	1.077	0.161	0.249	1.547
v708	92	0.001	3.670	0.111	0.407	3.667
v712	171	0.001	1.654	0.098	0.224	2.286
v715	84	0.0001	0.689	0.060	0.103	1.717
v750	140	0.001	2.043	0.149	0.266	1.785
v751	32	0.001	0.127	0.017	0.026	1.529
v752	58	0.0001	0.379	0.136	0.495	3.640
v755	23	0.0001	0.641	0.090	0.148	1.644
v756	104	0.001	2.994	0.144	0.329	2.285
v757	13	0.001	1.858	0.195	0.488	2.503
v770	154	0.001	8.271	0.134	0.680	5.075
v98	255	0.001	3.649	0.127	0.418	3.291
v072	426	0.0001	4.051	0.195	0.401	2.062
v1010	1,573	0.0001	49.673	0.268	2.258	8.422
v1030	666	0.0001	35.062	0.179	1.395	7.787
v1310	183	0.0001	24.701	0.374	1.880	5.031
v1331	15	0.001	0.989	0.208	0.295	1.423

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Vein	Count	Minimum opt	Maximum opt	Mean opt	Std. Deviation	CV
v62	69	0.0009	2.760	0.313	0.613	1.956
v700	1,986	0.0001	188.785	0.286	4.508	15.754
v730	2,562	0.0001	47.662	0.223	1.286	5.769
v800	700	0.0001	2.574	0.145	0.290	2.006
v810	453	0.0001	3.900	0.173	0.433	2.505
v84	732	0.0001	17.290	0.149	0.699	4.702
v91	469	0.0001	6.580	0.168	0.394	2.347
vSG1	1,632	0.0001	1.209	0.061	0.126	2.067
v1020	486	0.0001	3.459	0.150	0.361	2.407
v400	617	0.001	13.900	0.214	0.897	4.192
v710	1690	0.001	25.547	0.290	1.049	3.617
v711	369	0.001	4.842	0.226	0.523	2.314
v713	390	0.001	6.391	0.180	0.486	2.700
v714	164	0.001	1.781	0.134	0.283	2.112

Table 14-6 Summary Channel Sample Assay Statistics

Vein	Count	Minimum opt	Maximum opt	Mean opt	Std. Deviation	CV
v1011	211	0.001	3.477	0.218	0.488	2.242
v1012	17	0.003	0.634	0.103	0.152	1.476
v1040	12	0.001	0.052	0.008	0.016	2.002
v1300	17	0.02	0.510	0.158	0.138	0.870
v63	83	0.0001	3.780	0.320	0.632	1.971
v731	1	0.005	0.005	0.005	0.000	0.000
v732	3	0.017	0.046	0.029	0.015	0.534
v86	293	0.001	16.810	0.401	1.056	2.634
v04	990	0.001	23.142	0.425	1.313	3.089
v100	119	0.001	2.700	0.221	0.480	2.172
v101	125	0.001	3.500	0.222	0.458	2.063
v500	20	0.001	0.870	0.107	0.184	1.720
v530	3	0.02	0.030	0.023	0.005	0.217
v712	47	0.001	2.714	0.244	0.463	1.898
v715	1	0.061	0.061	0.061	0.001	0.016
v750	49	0.004	0.734	0.121	0.165	1.364
v751	16	0.001	0.369	0.100	0.115	1.150
v756	16	0.001	6.088	0.959	1.690	1.762
v757	39	0.001	3.270	0.497	0.783	1.575

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		Minimum	Maximum	Mean	Std.
Vein	Count	opt	opt	opt	Deviation	CV
v770	13	0.001	3.546	0.385	0.935	2.429
v072	833	0.0001	23.020	0.485	1.368	2.822
v1010	1,370	0.0001	45.835	0.258	1.427	5.541
v1030	830	0.0001	5.188	0.184	0.406	2.208
v1310	468	0.001	48.038	0.754	3.403	4.511
v1331	4	0.011	0.119	0.040	0.053	1.330
v62	5	0.001	0.170	0.046	0.070	1.509
v700	293	0.001	9.436	0.215	0.774	3.599
v730	201	0.0001	25.830	0.254	1.846	7.267
v800	341	0.001	19.269	0.334	1.243	3.726
v810	257	0.001	7.527	0.286	0.710	2.477
v84	556	0.001	60.500	0.748	4.216	5.632
v91	255	0.0001	2.019	0.144	0.215	1.496
vSG1	3202	0.0001	3.890	0.120	0.207	1.732
v1020	343	0.001	7.753	0.206	0.535	2.597
v400	1000	0.001	23.224	0.254	1.008	3.969
v710	972	0.001	26.210	0.343	1.339	3.904
v711	125	0.001	4.094	0.384	0.811	2.112
v713	127	0.001	1.107	0.301	0.503	1.671
v714	15	0.001	0.886	0.151	0.255	1.689

14.7.

Compositing

Drillhole assay sample lengths within the updated veins range from 0.1 ft. to 10.0 ft., with an average sample length of 1.6 ft. Channel assay sample lengths within the defined veins range from 0.3 ft. to 10.0 ft., with an average sample length of 2.4 ft. A maximum length of 10.0 ft. was selected for compositing in order to generate single intercept composites across the width of the vein. Drillhole assay samples were capped at 10.0 opt prior to compositing.

Klondex geologists have identified where drillholes oriented sub-parallel to the strike of specific veins were not representative of local mineralization, and these were excluded from the compositing process. Klondex geologists also identified a small number of assay intervals that were not captured by the vein modelling process, and these were manually flagged for inclusion in the compositing process.

Length-weighted composites were calculated within the defined vein. Missing sample intervals or zero-grade assay intervals were assigned a value of 0.001 opt. The compositing process started at the first point of intersection between the drillhole and the vein intersected, and halted on exit from the vein wireframe. All residual composites were retained. The wireframes that represent the interpreted veins were also used to back-tag a rock code field into the assay and composite workspaces. The resulting composite data were visually validated against vein wireframes. Summary statistics were calculated separately by vein for the composite samples (Table 14.7) .

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Table 14-7 Summary Composite Statistics by Vein

		Minimum	Maximum	Mean	Std.
Vein	Count	opt	opt	opt	Deviation	CV
v1011	204	0.0001	2.503	0.144	0.294	2.043
v1012	21	0.0002	0.857	0.189	0.265	1.404
v1040	38	0.001	1.600	0.150	0.284	1.894
v1300	163	0.001	0.709	0.078	0.113	1.444
v1305	35	0.00011	0.168	0.040	0.050	1.257
v1320	14	0.001	0.237	0.086	0.072	0.844
v1330	6	0.001	1.348	0.291	0.526	1.807
v200	99	0.0004	0.896	0.053	0.108	2.051
v210	31	0.0001	1.988	0.139	0.357	2.571
v63	171	0.001	2.000	0.178	0.239	1.348
v731	40	0.0001	0.670	0.080	0.145	1.808
v732	30	0.0001	0.721	0.105	0.173	1.651
v820	56	0.001	0.793	0.054	0.130	2.394
v86	433	0.001	16.810	0.239	0.915	3.824
vCW2	114	0.0001	0.675	0.070	0.104	1.484
vCW3	162	0.0001	6.948	0.139	0.562	4.059
vCW4	161	0.0001	0.868	0.073	0.118	1.610
v04	1168	0.0001	6.419	0.198	0.535	2.703
v100	95	0.0002	2.340	0.182	0.347	1.906
v101	156	0.0003	1.450	0.107	0.234	2.182
v500	53	0.001	0.956	0.151	0.215	1.423
v505	41	0.001	0.428	0.107	0.110	1.024
v507	30	0.001	0.671	0.085	0.134	1.588
v510	98	0.001	1.013	0.063	0.125	1.970
v511	87	0.0009	1.307	0.065	0.174	2.695
v512	61	0.001	1.478	0.078	0.214	2.751
v513	26	0.0009	0.297	0.056	0.093	1.650
v515	20	0.001	1.214	0.179	0.359	2.010
v520	9	0.01	0.250	0.093	0.071	0.763
v522	7	0.001	0.790	0.285	0.307	1.076
v530	57	0.001	0.528	0.089	0.100	1.118
v707	24	0.0001	0.493	0.046	0.103	2.228
v708	32	0.001	0.902	0.074	0.168	2.256

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		Minimum	Maximum	Mean	Std.
Vein	Count	opt	opt	opt	Deviation	CV
v712	166	0.0001	1.033	0.063	0.156	2.491
v715	72	0.0001	0.383	0.032	0.067	2.103
v750	97	0.0003	1.486	0.091	0.180	1.969
v751	58	0.001	0.653	0.087	0.125	1.442
v752	30	0.0002	0.932	0.059	0.174	2.956
v755	13	0.0001	0.226	0.060	0.069	1.163
v756	88	0.0001	4.815	0.210	0.698	3.324
v757	42	0.001	7.565	0.509	1.282	2.521
v770	149	0.0002	4.284	0.085	0.457	5.400
v98	939	0.0001	10.000	0.415	0.853	2.056
v072	447	0.0003	6.451	0.417	0.640	1.534
v1010	1067	0.001	22.996	0.199	0.854	4.284
v1030	658	0.001	3.407	0.115	0.271	2.344
v1310	457	0.0001	48.038	0.633	2.865	4.524
v1331	10	0.001	0.532	0.152	0.178	1.173
v62	35	0.001	2.760	0.265	0.589	2.225
v700	646	0.0001	4.719	0.135	0.374	2.767
v730	634	0.001	25.830	0.218	1.107	5.071
v800	410	0.00085	19.269	0.236	1.037	4.401
v810	240	0.0001	3.900	0.163	0.390	2.386
v84	390	0.001	10.000	0.357	0.993	2.783
v91	262	0.001	6.580	0.163	0.431	2.642
vSG1	1656	0.001	1.639	0.101	0.131	1.302
v1020	414	0.0001	2.894	0.124	0.266	2.136
v400	619	0.0003	10.000	0.228	0.671	2.943
v710	766	0.0001	7.565	0.249	0.524	2.104
v711	201	0.0003	2.570	0.196	0.365	1.867
v713	203	0.0005	2.163	0.177	0.288	1.626
v714	170	0.0001	0.841	0.048	0.123	2.536
v730	634	0.001	25.830	0.218	1.107	5.071
v800	410	0.00085	19.269	0.236	1.037	4.401
v810	240	0.0001	3.900	0.163	0.390	2.386
v84	390	0.001	10.000	0.357	0.993	2.783
v91	262	0.001	6.580	0.163	0.431	2.642
vSG1	1656	0.001	1.639	0.101	0.131	1.302
v1020	414	0.0001	2.894	0.124	0.266	2.136
v400	619	0.0003	10.000	0.228	0.671	2.943
v710	766	0.0001	7.565	0.249	0.524	2.104
v711	201	0.0003	2.570	0.196	0.365	1.867

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		Minimum	Maximum	Mean	Std.
Vein	Count	opt	opt	opt	Deviation	CV
v713	203	0.0005	2.163	0.177	0.288	1.626
v714	170	0.0001	0.841	0.048	0.123	2.536

14.8.

Treatment of Extreme Values

Grade capping analysis was conducted on the vein-coded and composited grade data in order to evaluate the potential influence of extreme values during Au grade estimation. The presence of high-grade outliers was identified by examination of histograms and log-probability plots. Composites were capped to the selected value prior to estimation.

In addition, due to the highly channelized nature of the mineralization a 25.0 ft. range restriction on samples greater than the 97.5th percentile of the local vein composite data set was used during the first Au grade estimation pass for the updated veins (Table 14.8) .

Table 14-8 Composite Capping Levels

Vein	Cap (opt)	97.5th Percentile (opt)	Vein	Cap (opt)	97.5th Percentile (opt)
v1011	1.4	0.994	v707	1.0	0.129
v1012	NA	0.500	v708	2.0	0.314
v1040	NA	0.600	v712	1.0	0.511
v1300	NA	0.382	v715	1.0	0.196
v1305	NA	NA	v750	1.0	0.436
v1320	NA	NA	v751	1.0	0.553
v1330	NA	0.100	v752	1.0	0.206
v200	NA	0.304	v755	1.0	0.226
v210	NA	0.600	v756	4.0	0.985
v63	NA	0.500	v757	5.0	3.182
v731	NA	0.667	v770	3.0	0.347
v732	NA	0.650	v98	8.0	2.850
v820	NA	0.361	v072	5.0	1.670
v86	5.0	1.350	v1010	5.0	1.184
vCW2	NA	0.462	v1030	2.0	0.812
vCW3	2.0	0.512	v1310	9.0	3.476
vCW4	NA	0.352	v1331	NA	0.100
v04	3.0	1.529	v62	NA	0.500
v100	2.0	1.209	v700	2.4	0.930
v101	1.0	0.782	v730	4.0	1.254
v500	1.0	0.870	v800	3.0	0.991
v505	1.0	0.415	v810	2.0	0.976
v507	1.0	0.227	v84	7.0	0.889

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Vein	Cap (opt)	97.5th Percentile (opt)	Vein	Cap (opt)	97.5th Percentile (opt)
v510	1.0	0.340	v91	3.0	0.715
v511	2.0	0.456	vSG1	1.0	0.426
v512	2.0	0.616	v1020	1.0	0.599
v513	1.0	0.273	v400	5.0	1.085
v515	1.0	1.214	v710	NA	1.305
v520	1.0	0.250	v711	2.0	1.237
v522	1.0	0.790	v713	NA	0.999
v530	1.0	0.323	v714	NA	0.449

14.9.

Au Grade Estimation, Classification and Minimum Width

The mineral resource estimate reported herein was constrained by vein wireframes that form hard boundaries between their respective composite sample extents. All Au block grades were estimated using a triple pass Inverse Distance cubed (“ID3”) weighting of between four and twelve capped composite grades from two or more drillholes. In addition, due to the varying sample support lengths generated by the compositing process, a variable weighting factor defined by the length of the composite samples was also implemented during Au grade estimation. A discretization level of 4 x 4 x 1 was used for the updated veins.

During the first pass, drillhole and channel samples were selected using an anisotropic search with a radius of 50.0 ft x 50.0 ft x 25.0 ft. All blocks estimated during the first pass were algorithmically assigned a classification of Measured.

During the second pass, drillhole samples were selected using an anisotropic search with a radius of 100.0 ft x 100 ft x 50 ft. For the second pass, estimation was implemented using a 25.0 ft. by 25.0 ft. by 25.0 ft. parent block size. All blocks estimated during the second pass were algorithmically assigned a classification of Indicated.

During the third pass, drillhole samples were selected using an anisotropic search with a radius of 500.0 ft x 500 ft x 150 ft. For the third pass, estimation was implemented using a 50.0 ft. by 50.0 ft. by 50.0 ft. parent block size. All blocks estimated during the third pass were algorithmically assigned a classification of Inferred.

Blocks that did not meet the minimum sample requirements were not estimated.

Klondex has defined a minimum mining width of 4.0 ft. For blocks with a width less than 4.0 ft. the estimated grade was diluted at zero grade to a 4.0 ft. width with an associated increase in tonnage, and the adjusted grade and tonnage used for mineral resource reporting.

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A Nearest Neighbor model (“NN”) was also generated simultaneously using the same parameters as the third estimation pass.

14.10.

Mineral Resource Estimate

Mineral resources are based on a cut-off grade of 0.09 opt (Table 14.9) . The cut-off grade has been calculated from the following parameters:

	•	Gold Price: US$ 1,400.00 per oz
	•	Exchange Rate: C$ to US$: 0.80
	•	Mine & Mill Cost: C$141.52 per ton
	•	Recovery: 94%
	•	Cut-off: 0.090 opt Au (3.09 grams per tonne)

Table 14-9 Total Mineral Resources^{1,2,3,4, 5}

	Grade	Grade
Class	Au opt	Au g/t	Tons	Au oz
Measured	0.220	7.54	521,000	115,000
Indicated	0.214	7.34	1,276,000	273,000
Meas + Ind	0.216	7.40	1,797,000	388,000
Inferred	0.182	6.24	3,676,000	668,000

	1.	Mineral resources which are not mineral reserves do not have demonstrated economic viability. The estimate of mineral resources may be materially affected by environmental, permitting, legal, marketing, or other relevant issues.
	2.	Mineral resources were estimated using the Canadian Institute of Mining, Metallurgy and Petroleum (CIM), CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines prepared by the CIM Standing Committee on Reserve Definitions and adopted by CIM Council.
	3.	The quantity and grade of reported Inferred resources in this estimation are uncertain in nature and there has been insufficient exploration to define these inferred resources as an Indicated or Measured mineral resource and it is uncertain if further exploration will result in upgrading them to an Indicated or Measured mineral resource category.
	4.	Contained metal may differ due to rounding.
	5.	Cut-off grade = 0.090 opt Au.

Table 14-10 Detail of the Resource by Vein

		Measured			Indicated			Mea+Ind			Inferred
Vein	grade	tons	ozs	grade	tons	ozs	grade	tons	ozs	grade	tons	ozs
v04	0.179	19,000	3,000	0.127	4,000	0	0.171	22,000	4,000	0.090	0	0
v100	0.325	14,000	4,000	0.266	13,000	4,000	0.296	27,000	8,000	0.231	12,000	3,000
v101	0.189	11,000	2,000	0.134	8,000	1,000	0.165	19,000	3,000	0.102	0	0
v1020	0.185	16,000	3,000	0.305	38,000	11,000	0.269	54,000	15,000	0.229	87,000	20,000
v400	0.190	14,000	3,000	0.283	41,000	12,000	0.259	54,000	14,000	0.250	176,000	44,000
v500	0.173	5,000	1,000	0.170	16,000	3,000	0.171	21,000	4,000	0.215	45,000	10,000
v505	0.161	5,000	1,000	0.143	23,000	3,000	0.146	29,000	4,000	0.170	61,000	10,000

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		Measured			Indicated			Mea+Ind			Inferred
Vein	grade	tons	ozs	grade	tons	ozs	grade	tons	ozs	grade	tons	ozs
v507	0.000	0	0	0.182	7,000	1,000	0.182	7,000	1,000	0.102	19,000	2,000
v510	0.147	2,000	0	0.151	11,000	2,000	0.150	13,000	2,000	0.120	14,000	2,000
v511	0.152	3,000	0	0.293	12,000	3,000	0.268	14,000	4,000	0.187	17,000	3,000
v512	0.000	0	0	0.190	11,000	2,000	0.190	11,000	2,000	0.176	12,000	2,000
v513	0.243	2,000	0	0.173	10,000	2,000	0.183	11,000	2,000	0.165	21,000	3,000
v515	0.000	0	0	0.197	5,000	1,000	0.197	5,000	1,000	0.144	37,000	5,000
v520	0.125	0	0	0.149	4,000	1,000	0.149	4,000	1,000	0.127	35,000	4,000
v522	0.000	0	0	0.000	0	0	0.000	0	0	0.361	116,000	42,000
v530	0.129	6,000	1,000	0.103	5,000	1,000	0.117	11,000	1,000	0.107	0	0
v707	0.100	0	0	0.206	4,000	1,000	0.203	4,000	1,000	0.215	17,000	4,000
v708	0.173	0	0	0.329	17,000	6,000	0.328	17,000	6,000	0.157	84,000	13,000
v710	0.254	94,000	24,000	0.272	138,000	38,000	0.265	232,000	62,000	0.224	408,000	91,000
v711	0.255	12,000	3,000	0.339	82,000	28,000	0.328	95,000	31,000	0.196	107,000	21,000
v712	0.206	3,000	1,000	0.171	13,000	2,000	0.178	16,000	3,000	0.130	10,000	1,000
v713	0.219	23,000	5,000	0.223	53,000	12,000	0.222	76,000	17,000	0.150	122,000	18,000
v714	0.213	6,000	1,000	0.221	30,000	7,000	0.220	36,000	8,000	0.113	36,000	4,000
v715	0.117	0	0	0.193	5,000	1,000	0.187	5,000	1,000	0.149	17,000	3,000
v750	0.147	8,000	1,000	0.154	24,000	4,000	0.152	32,000	5,000	0.234	23,000	5,000
v751	0.133	1,000	0	0.171	2,000	0	0.163	3,000	0	0.120	1,000	0
v752	0.116	1,000	0	0.194	11,000	2,000	0.189	12,000	2,000	0.500	20,000	10,000
v755	0.000	0	0	0.000	0	0	0.000	0	0	0.099	2,000	0
v756	0.208	5,000	1,000	0.122	5,000	1,000	0.165	10,000	2,000	0.099	2,000	0
v757	0.409	3,000	1,000	0.291	2,000	0	0.373	5,000	2,000	0.194	3,000	0
v770	0.300	1,000	0	0.195	8,000	2,000	0.206	9,000	2,000	0.153	1,000	0
v98	0.142	2,000	0	0.243	10,000	2,000	0.223	12,000	3,000	0.130	1,000	0
v62	0.307	3,000	1,000	0.230	7,000	1,000	0.253	9,000	2,000	0.178	4,000	1,000
v072	0.199	10,000	2,000	0.174	8,000	1,000	0.188	18,000	3,000	0.135	40,000	5,000
v84	0.234	25,000	6,000	0.166	11,000	2,000	0.212	36,000	8,000	0.160	17,000	3,000
v91	0.199	51,000	10,000	0.194	50,000	10,000	0.197	101,000	20,000	0.156	24,000	4,000
v700	0.188	10,000	2,000	0.135	22,000	3,000	0.151	32,000	5,000	0.170	210,000	36,000
v730	0.243	17,000	4,000	0.206	90,000	19,000	0.212	107,000	23,000	0.128	76,000	10,000
v800	0.164	3,000	0	0.153	24,000	4,000	0.154	26,000	4,000	0.129	20,000	3,000
v810	0.242	5,000	1,000	0.214	26,000	6,000	0.218	31,000	7,000	0.188	40,000	8,000
v1010	0.183	12,000	2,000	0.195	55,000	11,000	0.193	67,000	13,000	0.181	167,000	30,000
v1030	0.278	9,000	3,000	0.173	39,000	7,000	0.194	48,000	9,000	0.189	125,000	24,000
v1310	0.384	16,000	6,000	0.153	5,000	1,000	0.333	21,000	7,000	0.147	37,000	5,000
v1331	0.267	0	0	0.230	1,000	0	0.231	1,000	0	0.234	12,000	3,000
vSG1	0.131	16,000	2,000	0.140	81,000	11,000	0.139	97,000	13,000	0.131	438,000	57,000
v63	0.191	34,000	6,000	0.211	26,000	5,000	0.200	59,000	12,000	0.167	75,000	13,000
v86	0.189	13,000	3,000	0.186	18,000	3,000	0.187	32,000	6,000	0.204	14,000	3,000

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		Measured			Indicated			Mea+Ind			Inferred
Vein	grade	tons	ozs	grade	tons	ozs	grade	tons	ozs	grade	tons	ozs
v200	0.301	4,000	1,000	0.252	9,000	2,000	0.268	13,000	3,000	0.138	207,000	29,000
v210	0.147	0	0	0.194	12,000	2,000	0.193	12,000	2,000	0.209	181,000	38,000
v731	0.235	3,000	1,000	0.217	19,000	4,000	0.219	22,000	5,000	0.151	110,000	17,000
v732	0.279	5,000	1,000	0.183	19,000	3,000	0.204	24,000	5,000	0.137	30,000	4,000
v820	0.197	5,000	1,000	0.179	13,000	2,000	0.184	18,000	3,000	0.135	3,000	0
v1011	0.256	2,000	1,000	0.277	1,000	0	0.262	3,000	1,000	0.133	0	0
v1012	0.327	2,000	0	0.232	14,000	3,000	0.241	15,000	4,000	0.194	26,000	5,000
v1040	0.247	1,000	0	0.250	8,000	2,000	0.250	9,000	2,000	0.196	4,000	1,000
v1300	0.148	13,000	2,000	0.167	28,000	5,000	0.161	42,000	7,000	0.142	33,000	5,000
v1305	0.116	0	0	0.096	2,000	0	0.099	2,000	0	0.000	0	0
v1320	0.000	0	0	0.170	4,000	1,000	0.170	4,000	1,000	0.103	50,000	5,000
v1330	0.000	0	0	0.198	1,000	0	0.198	1,000	0	0.333	8,000	3,000
vCW2	0.126	0	0	0.129	10,000	1,000	0.129	10,000	1,000	0.152	80,000	12,000
vCW3	0.157	4,000	1,000	0.173	35,000	6,000	0.171	38,000	7,000	0.159	69,000	11,000
vCW4	0.183	1,000	0	0.196	30,000	6,000	0.196	31,000	6,000	0.195	70,000	14,000
TOTAL	0.220	521,000	115,000	0.214	1,276,000	273,000	0.216	1,797,000	388,000	0.182	3,676,000	668,000

14.11.

Block Model Validation

The block models were validated visually by the inspection of successive section lines in order to confirm that the block models correctly reflect the distribution of high-grade and low-grade values. Comparative histograms were also generated (see Appendices). Unadjusted block grades were compared an unadjusted Nearest Neighbor model generated using the same search criteria as that used for the mineral resource estimate (Table 14.11) .

Table 14-11 Block Model Validation Grades

Vein	ID3 Mean opt	NN Mean opt	Vein	ID3 Mean opt	NN Mean opt
v1011	0.069	0.067	v707	0.063	0.058
v1012	0.125	0.167	v708	0.095	0.109
v1040	0.222	0.205	v712	0.055	0.057
v1300	0.054	0.058	v715	0.035	0.036
v1305	0.041	0.040	v750	0.068	0.080
v1320	0.087	0.072	v751	0.096	0.097
v1330	0.077	0.234	v752	0.054	0.054
v200	0.048	0.053	v755	0.072	0.040
v210	0.083	0.127	v756	0.101	0.106

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Vein	ID3 Mean opt	NN Mean opt	Vein	ID3 Mean opt	NN Mean opt
v63	0.129	0.104	v757	0.229	0.195
v731	0.122	0.104	v770	0.049	0.052
v732	0.112	0.109	v98	0.099	0.100
v820	0.042	0.030	v072	0.152	0.163
v86	0.075	0.090	v1010	0.089	0.090
vCW2	0.071	0.061	v1030	0.081	0.079
vCW3	0.081	0.086	v1310	0.176	0.151
vCW4	0.061	0.056	v1331	0.224	0.208
v04	0.070	0.074	v62	0.090	0.188
v100	0.129	0.123	v700	0.074	0.071
v101	0.071	0.056	v730	0.066	0.067
v500	0.152	0.159	v800	0.089	0.089
v505	0.117	0.119	v810	0.071	0.069
v507	0.080	0.080	v84	0.072	0.135
v510	0.062	0.063	v91	0.114	0.110
v511	0.075	0.065	vSG1	0.052	0.049
v512	0.064	0.075	v1020	0.082	0.081
v513	0.066	0.071	v400	0.162	0.156
v515	0.146	0.123	v710	0.165	0.162
v520	0.120	0.122	v711	0.170	0.165
v522	NA	NA	v713	0.124	0.121
v530	0.080	0.080	v714	0.058	0.053

Swath plots were generated to check the block model estimation bias by comparing the nearest neighbor (NN) block estimate to the ID3 estimate for Measured and Indicated resources (see Appendices).

As a further check of the mineral resource model the total volume reported at zero Au cut-off was compared by vein with the calculated volume of the defining mineralization wireframe (Table 14-12). All reported volumes fall within acceptable tolerances.

Table 14-12 Volume Comparison

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	Wireframe			Wireframe
	Volume (cubic	Block Volume		Volume	Block Volume
Vein	feet)	(cubic feet)	Vein	(cubic feet)	(cubic feet)
v1011	1,547,795	1,547,129	v707	1,020,740	1,011,982
v1012	1,170,010	1,170,844	v708	3,080,210	3,054,931
v1040	240,220	239,926	v712	2,498,942	2,485,743
v1300	4,288,995	4,291,264	v715	2,061,603	2,040,028
v1305	1,115,260	1,115,306	v750	1,714,643	1,709,792
v1320	1,539,625	1,542,305	v751	222,287	206,776
v1330	772,756	774,116	v752	1,105,123	1,114,648
v200	23,078,214	25,549,080	v755	1,102,244	1,091,253
v210	9,920,635	11,320,757	v756	375,939	375,790
v63	3,330,105	3,425,009	v757	310,740	301,351
v731	2,482,877	2,482,616	v770	1,106,404	1,103,428
v732	1,115,265	1,115,264	v98	1,155,709	1,175,592
v820	2,103,755	2,105,643	v072	4,388,856	4,388,991
v86	3,981,750	3,988,093	v1010	14,318,142	14,507,150
vCW2	5,575,025	6,389,382	v1030	11,931,908	12,788,126
vCW3	6,303,789	7,081,996	v1310	2,079,285	2,079,014
vCW4	9,005,981	11,431,644	v1331	193,962	567,757
v04	8,614,068	8,433,816	v62	695,815	696,313
v100	1,856,915	1,826,021	v700	12,000,717	12,088,160
v101	1,337,231	1,316,472	v730	12,317,510	12,526,653
v500	1,181,478	1,183,702	v800	5,977,515	5,977,193
v505	1,755,391	1,747,869	v810	7,311,945	7,359,619
v507	1,312,126	1,303,770	v84	6,724,538	5,863,299
v510	1,158,255	1,160,933	v91	3,501,150	3,501,065
v511	1,353,128	1,363,918	vSG1	54,522,427	64,501,083
v512	1,051,374	1,059,457	v1020	5,975,060	5,974,179
v513	1,448,757	1,440,703	v400	7,896,820	8,079,136
v515	1,179,031	1,184,524	v710	9,644,615	12,693,819
v520	1,271,638	1,246,999	v711	8,030,502	8,080,407
v522	1,397,845	1,407,290	v713	2,114,117	2,114,464
v530	573,778	598,291	v714	2,082,206	2,082,142

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15.

Mineral Reserve Estimates

15.1.

Methodology

15.1.1.

Underground Reserves

Excavation designs for stopes, stope development drifting and access development were created using Vulcan software. Stope designs were aided by the Vulcan Stope Optimizer Module. The stope optimizer produces the stope cross section which maximizes value within given geometric, mining and economic constraints.

The minimum mining width for stopes is the greater of the vein width plus 1.25 feet of planned dilution on both the hanging wall and foot wall or 5.5 feet. Rather than adding a fixed percentage dilution to each stope this approach dilutes narrow veins more severely. Thus, for three-foot wide vein the planned dilution is 83% but for a ten-foot wide vein the planned dilution is 25%. The minimum stope footwall dip is set to 50° to prevent the accumulation of blasted mineralization that will not flow to the stope sill where it can be removed.

Stope development drifts are rectangular drifts eight feet wide and nine feet high driven along strike of the vein. When the drift does not encompass the entire vein the stope design is adjusted to incorporate all the mineralization.

Development, production mining and backfill tasks were created from all designed excavations. These tasks were assigned costs and productivities specific to the excavation or backfill task type. Additionally, the undiscounted cash flow for each task was calculated. All tasks were then ordered in the correct sequence for mining and backfilling. Any task sequence or subsequence that did not achieve a positive cumulative undiscounted cash flow was removed from consideration for mineral reserves. Stope development excavations necessary to the extraction of ore grade stopes and that exceed the incremental cut-off grade are also included in reserves.

Table 15-1 Mineral Reserves Cut Off Grade Calculation

Gold Sales Price	$/Ounce	$1,500
Refining and Sales Expense	$/Ounce	-
Royalty		0%
Metallurgical Recovery		94%
Total Cost	$/ton	$212
Incremental Cut Off Grade		0.08
Cut-off Grade	opt	0.15
Minimum Mining Width	feet	5.5
Grade Thickness cut-off	Eq. opt-ft.	0.825

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Figure 15-1 Grouping of Diluted Stope Optimizer Rings to Create Stopes

Figure 15-2 26L – 710 Complex Final Reserves Plan by Year Mined

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Figure 15-3 16L 8/10 - Reserves Mined by Year

Figure 15-4 26L V91 (SAM) - Reserves Mined by Year

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Figure 15-5 Cohiba - Reserves Mined by Year

15.1.2.

Tailings Reserves

Open pit methods were applied to the historic tailings to determine reserves. A Lerch-Grossman optimal pit was created using the tailings block model and the economic parameters in Table 15-2. Maximum embankment slopes were limited to 30°.

Thirty-percent of the material within the optimum pit is below the 0.016 opt cut-off grade. Since it is likely that this material cannot be segregated during recovery operations this sub-grade material is treated as dilution at the indicated block model grade of 0.014 opt and included in the process stream and the Mineral Reserve.

Table 15-2 Tailings Reprocessing Cut-off Grade

Gold Sales Price	$/Ounce	$1,500
Refining and Sales Expense	$/Ounce	-
Royalty		0%
Metallurgical Recovery		89%
Total Operating Cost	$/ton	$19.32
Sustaining Capital	$/ton	$2.56
Total Cost	$/ton	$21.88
Cut-off Grade	opt	0.016

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15.2.

Statement of Reserves

True North underground and tailings Mineral Reserves are summarized by mining district and vein in Table 15-3. These reserves have been estimated using the methodology described above.

Table 15-3 True North Mineral Reserves as of March 31, 2017

								Proven and Probable
		Proven Reserves			Probable Reserves				Reserves
	Vein	Tons		Au oz.	Tons		Au Oz.	Tons		Au Oz.
District		(000's)	Au opt	(000's)	(000's)	Au opt	(000's)	(000's)	Au opt	(000's)
	710	50.8	0.258	13.1	90.9	0.266	24.2	141.7	0.263	37.3
	711	12.0	0.192	2.3	70.8	0.306	21.6	82.7	0.289	23.9
	713	10.5	0.165	1.7	31.3	0.212	6.6	41.8	0.200	8.4
	708	0.1	0.148	-	18.3	0.272	5.0	18.4	0.272	5.0
	714	3.1	0.225	0.7	15.7	0.238	3.7	18.9	0.236	4.4
	712	2.0	0.149	0.3	3.3	0.187	0.6	5.3	0.173	0.9

26L – 710 Complex	750	2.4	0.137	0.3	1.4	0.147	0.2	3.7	0.141	0.5
	707	-	-	-	1.5	0.208	0.3	1.5	0.208	0.3
	756	1.5	0.182	0.3	-	-	-	1.5	0.182	0.3
	1030	7.9	0.264	2.1	5.7	0.194	1.1	13.6	0.235	3.2
	810	1.4	0.255	0.4	7.1	0.287	2.0	8.5	0.282	2.4
16L – 8/10	1020	5.1	0.133	0.7	8.7	0.187	1.6	13.8	0.167	2.3
	1010	2.0	0.138	0.3	2.5	0.172	0.4	4.5	0.156	0.7
26L-
SAM	V91	22.7	0.203	4.6	28.3	0.195	5.2	49.4	0.199	9.8
Cohiba	400	6.6	0.172	1.1	21.6	0.190	4.1	28.3	0.186	5.2
UG		128	0.218	27.9	306	0.251	76.9	434	0.242	104.7
Tailings					1,950	0.022	43.2	1,950	0.022	43.2
Total		128	0.218	27.9	2,256	0.053	120.1	2,384	0.062	147.9

	Notes:
	7.	Mineral reserves have been estimated with a gold price of US$1,200/ounce or C$1.500/oz
	8.	US$:CDN$ exchange rate is 0.80;
	9.	Metallurgical recovery for underground and tailings Mineral Reserves is 94% and 89% respectively;
	10.	Underground Mineral Reserves are estimated at a cut-off grade of 0.15 Au opt and an incremental cut-off grade of 0.08 Au opt;
	11.	Tailings Mineral Reserves are estimated at a cut-off grade of 0.016 opt, and;
	12.	Mine losses of 5% and no unplanned mining dilution have been applied to the underground Mineral Reserves and mine losses of 8% and no unplanned dilution have been applied to the tailings Mineral Reserves.

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16.

Mining Methods

16.1.1.

Access Development

True North is an underground mining operation that has been in almost continuous operation since the early 20th century. Over the years, the mine has employed many mining methods including shrinkage, sublevel stoping, cut and fill and panel stoping.

Currently, the Project has two main production levels, namely 16L and 26L, which are accessed via a 4,400 foot (1,341m) two compartment shaft (A-Shaft). The 710 Zone mining complex is the main mining area and is located approximately 6,600 feet (2,000m) from the main shaft along the 26L access level.

The main haulage track drift on 26L is used to access the 710 Vein. In order to access above and below the 26L, the 710 Vein has a 12 foot by 12 foot (3.7m x 3.7m) incline and decline driven at a maximum gradient of +/-15% with access cross-cuts into the ore body every 60 feet (18m) vertically. Additional infrastructure along the incline includes a vertical ventilation and escape raise and an ore and waste pass system.

A longitudinal section through the mine is provided in Figure 16.1.

16.1.2.

Geotechnical

Rock characteristics at True North are typical northern Canadian shield conditions with very little water and very competent with RMR ranging from 65% to 85%. In large areas and stopes all necessary joint sets are mapped and a ground support standards are reviewed using Unwedge.

The Uniaxial Compressive Strength (UCS) of the ground is 200 million pascal (MPa), due to the depth of the 710 vertical stress will range from 30MPa to 40MPa with maximum horizontal stress being 1.3 times the vertical stress.

16.1.3.

Ground Support

The ground conditions at True North are typical of those found elsewhere in the northern shield, with typically dry and very competent conditions. The main ground support system is resin encapsulated 6 ft (1.8m) #6 rebar bolted in a 4 foot by 4 foot (1.2m by 1.2m) pattern, supporting 4 inch (0.1m) welded wire mesh. In areas wider than 18 feet (5.5m), 8 feet (2.4m) #6 rebar bolts replace the 6 foot (1.8m) rebar bolts.

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Where more adverse conditions are encountered in longhole stopes, 20 feet (6m) and 30 feet (9m) long grout encapsulated cable bolts are installed as well as in intersections with span between 20 feet and 30 feet (6m to 9m). The cable bolt pattern is determined from the specific conditions and are drilled with a longhole drill and then fully grouted.

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Figure 16-1 LongitudinalSection

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16.1.4.

Ventilation and Secondary Egress

Underground mining relies on diesel equipment in the process to extract the ore and waste rock and to transport backfill to the stopes. The Project is ventilated through an intake connection to surface. This air route includes travel through historic mining areas and from there, it is directed to the 24L and distributed to the main 710 Zone mining horizon via horizontal and vertical openings. All the air eventually exhausts out through A- shaft. Fans located in development headings ventilate working faces. The surface intake fan are two 150 horsepower (hp) (110 kilo Watt [kW]) fans and the 24L main ventilation fans deliver 75,000 cubic feet per minute (cfm) (5.7 m³/sec) to the 710 Zone. A ventilation schematic is provided in Figure 16.2.

Two means of secondary egress are available at the project. These include a man-way with ladders and landings in A-shaft and a second man-way connection to surface is via a timbered raise from 26L to 16L, which continues on to a man-way up B shaft to 8L and then on to a series of other raises from 8L to surface.

16.2.

Power Distribution and Dewatering

Electrical power to the mine is provided by a 4,160-volt feeder connection which is stepped down to 480 volts for distribution. Step down transformers and circuit protection are provided by 22 load centers located throughout the mine. Excess mine water is dewatered from 26L to 16L to 10L and then to the process plant where is it sent to tailings. The mine purges water at a rate of approximately 300 gallon (1.36m ³) per week, however, most water is recycled and inflow from the surrounding rock is minimal.

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Figure 16-2 True North Gold Mine 710 Complex- VentilationSystem

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16.3.

Mining Methods

The primary mining method at the Project is longhole stoping which is a cost-effective method to mine the complex geology at the Project, and benefits from a quick stope cycle time. In areas where mineralization does not warrant the development of a ramp access system, Klondex employs captive sublevel stoping methods

16.3.1.

Longhole Stoping

Longhole stoping is the lowest cost method used at the Project and generally also provides the lowest total cost per ounce of gold produced.

Level accesses are driven perpendicular towards the ore body every 60 feet (18m) vertical. From these access drifts, 8 foot x 9 foot (2.4m x 2.7m) sills are developed along the strike of the mineralized zone (Figure 16-3).

Figure 16-3 Longhole Open Stope Sill Development

Once the levels are developed, a slot raise is driven between the levels which provides the free face necessary for longhole blasting. Subsequently, longhole drilling is carried out with 2.5 inch (64mm) holes on a 3foot (0.9m) ring burden. The actual drill pattern is determined by the stope shape. (Figure 16-4).

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Figure 16-4 Longhole Open Stope Raise and Drilling

Once all longhole drilling is completed, the stope is loaded with explosives and blasted. A diesel powered load-haul-dump machine (LHD) is used to move the blasted material from the undercut. The LHD is equipped with line-of-sight remote control mechanism to allow the removal of all blasted rock without exposing operating personnel to the open stope and the potential risk of ground falls (Figure 16-5).

Figure 16-5 Longhole Open Stope Blasting

After all blasted material has been extracted, the remaining void is backfilled with waste rock (Figure 16-6).

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Figure 16-6 Longhole Open Stope Backfilling

Figure 16-7 shows a typical over-cut and undercut access and sill in plan view, and Figure 16-8 shows a typical drill ring section.

Figure 16-7 Overcut and Undercut Plan View

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Figure 16-8 True North Gold Mine Typical Longhole Drill Section

16.3.2.

Captive Sub Level Longhole Stoping

Captive sub level longhole stoping is used in situations where an up-ramp access cannot be economically justified. This mining method uses three 6 foot by 6 foot (1.8m by 1.8m) raises up to 60 feet (18m) in length. One raise acts as a slot raise for the longhole blasts to slash into; the second raise acts as a mill-hole for muck from the sublevel where broken ore is slushed into; and the last is a access man-way. At the top of the raises an 8 foot by 9 foot (2.4m by 2.7m) sublevel is driven along the strike of the ore body. An air slusher is used to move the development muck from the sublevel over to and down the mill hole. If the mineralization warrants, another series of raises and a sublevel are developed.

After all of the raises and sublevels are developed, longhole drilling is carried out with an air drill capable of drilling 2.5 inch (64mm) holes on a 3 foot (0.9m) ring burden. These rings are blasted into the slot raises. The material drops to the bottom level and a remote controlled LHD removes the blasted ore without exposing the operator to the open stope and the potential risk of ground falls.

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Figure 16-9 shows an arrangement of two sub-levels of a sub level captive longhole stope.

Figure 16-9 True North Gold Mine Sub-Level Captive Longhole Stope

16.3.3.

Haulage

Ore and waste rock generated from the incline of the 710 Zone mining complex is hauled with LHDs to the vertical ore and waste chute system which connects to the 26L haulage drift. The material is then loaded into rail cars and hauled from the rock passes along the haulage drift with a diesel locomotive and 5 ton (4.5 tonne) rail cars. The cars are dumped at a grizzly equipped with a rock breaker and the material is sent to the loading pocket below 26L, from where it is hoisted in the shaft to surface in a 5 ton (4.5 tonne) capacity skip.

In the 710 Zone mining complex decline, ore and waste are hauled to the 26L haulage drift via two 13 ton (12 tonne) rubber tired underground mining haul trucks. These haul trucks deliver the ore and waste rock to the 710 ore and waste bins. From these ore bins, the follows the same route as described in the above paragraph.

16.3.4.

Backfill

Waste rock is moved from development to stoping whenever possible and major fill zones are created in the mining through the down hole longhole method on the 710 incline. The decline will not create fill void and will require sill pillars at certain level intervals. With the mining in the incline being filled, a percentage of waste rock will need to be removed from the mine but there will be sufficient material to fill the created void as per Table 16-1.

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Table 16-1 Waste Rock Backfill and Stope Voids

Period	2017	2018	2019	Total
Waste Rock Backfill Requred – (yd³)	20,000	57,500	47,000	124,500
Waste Mined – (yd3)	39,500	61,000	103,000	204,000

For all waste lateral development and stoping scheduling, the following parameters are utilized.

Waste development:

	•	Ramp – 20 feet/day
	•	Ramp Auxiliary – 20 feet/day

Ore development:

•

Ore Sills – 12 feet/day

Conventional raising:

•

8 feet/day

Drop Raise

•

20 feet/day

Backfilling:

•

150 tons/day

Longhole drilling:

600 feet/day

Table 16-2 Annual Production and Development Plan

Calendar Year	2017¹	2018	2019	Total

Reserves Mined
Proven Ore Mined (000's Tons)	49.3	58.5	20.2	128.0
Gold Grade (Ounce/Ton)	0.204	0.229	0.221	0.218
Contained Gold (000's Ounces)	10.1	13.4	4.4	27.9

Probable Ore Mined (000's Tons)	52.4	159.3	94.0	305.7
Gold Grade (Ounce/Ton)	0.235	0.243	0.276	0.251

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Calendar Year	2017¹	2018	2019	Total
Contained Gold (000's Ounces)	12.3	38.6	25.9	76.9

Total Ore Mined (000's Tons)	101.7	217.8	114.1	433.6
Gold Grade (Ounce/Ton)	0.220	0.239	0.266	0.242
Contained Gold (000's Ounces)	20.8	52.0	27.9	100.6

Production Mining
Ore and Incremental Development Mining (000's Tons)	16.8	26.1	3.4	46.2
Longhole Stope Mining (000's Tons)	85.0	191.7	110.8	387.5
Reserves Mined (000's Tons)	101.7	217.8	114.1	433.6

Backfill
Total Backfill (000's Tons)	37.3	107.4	88.3	233.0

Waste Mining
Expensed Waste Drifting including Waste Sills (Feet)	2,541	2,572	161	5,274
Expensed Waste (000's Tons)	14.9	15.1	0.9	31.0
Primary Capital Drifting (Feet)	5,223	8,114	141	13,478
Capital Raising (Feet)	356	502	-	858
Capitalized Mining (000's Tons)	58.1	98.0	1.5	157.6

Total Tons Mined (000's Tons)	174.7	330.9	116.5	622.2

The mine plan for 2017 includes only projections for the period from April 1, 2017 to December 31, 2017.

16.4.

Equipment Fleet Underground

Table 16-3 Underground Mobile Equipment Fleet

	Description	Quantity
	3.5yd³(2.7m³) scoop	4
	2.5yd³(1.9 m³) scoop	1
Development	Single boom jumbo	3
	Scissor decks	2
	13 ton (12 tonne) haul trucks	2
	Air powered longhole drill	2
Production	2.5yd³(1.9m³) remote controlled
	scoop	4
	Locomotive	4
Tramming	5 ton (4.5 tonne) rail cars	20

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	Description	Quantity
	Scissor lift	1
	Boom truck	1
Ancillary	Grease truck	1
	Telehandler	1
	Kubota RTV or Toyota Land Cruiser	3

16.5.

Tailings Reprocessing

Tailings reprocessing will be carried out from 2017 to 2028, initially during the underground mining operations (2017 to 2019) and subsequently following cessation mining operations (2020 to 2028), as a stand-alone operation. This seasonal reprocessing operation will be conducted at a process plant throughput of 1,200 tpd (1,189 tonnes per day).

Dry tailings material is extracted throughout the historic tailings and moved to the pump station using a dozer. The pump station consists of a 50 HP (37 kW) pump, a reclamation water supply and a water reservoir. During operation, dry material can be moved at a rate of 39 tons (35 tonnes) per hour from the tailings pump. The material is fed into a reservoir and is diluted into slurry with high pressure water jets. The slurry is pumped from the historic tailings pond and into a holding tank located at the tailings substation, on the bank of the historic tailings pond. The tailings substation consists of a holding tank with a 75 HP (56kW) pump inside, valves to direct process water, electrical controls for the pumps and auxiliary equipment and a diesel powered generator.

From the holding tank, the slurry is pumped through a 10 inch (250mm) line to the grinding bay at the process plant. A flow rate is maintained using the pump located in the tank and a 50 HP (37 kW) booster pump located approximately 3,000 feet (1 km) along the reprocessing line. Process water is added to the holding tank as required in order to maintain a consistent suspended solids level, heading to the process plant. Tailings slurry is pumped to the process plant with a flow rate above 600 gallons (2.7m3) per minute maintained and a typical suspended solids rating of 20%. Tailings material is sent to a de-watering cyclone located in the process plant grinding bay where the overflow from the cyclone is returned to the new tailings pond and the underflow is sent into the grinding circuit.

A schedule for tailings reprocessing is shown Table 16-4.

Table 16-4 Tailings Reprocessing Schedule

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Period	2017	2018	2019	2020	2021-2028
Tons Reprocessed	80,000	115,000	115,000	195,000	1,445,000
Au opt Reprocessed	0.022	0.022	0.022	0.022	0.022
Au Oz Reprocessed	1,800	2,500	2,500	4,300	32,000
Au Oz Recovered	1,600	2,300	2,300	3,900	28,500

Figure 16-10 Tailings Recovery Flow Diagram

Figure 16-11 Aerial View of Tailings Recovery Site

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17.

Recovery Methods

In 2011, the process crushing plant was expanded by adding a new primary 30 inch by 42 inch (760mm by 1067mm) jaw crusher, a GP300 gyratory crusher as a secondary breaker and a Barmac B8000 VSI (vertical shaft impact) crusher as a tertiary unit. There units are all in line with a 7 foot by 20 foot (2.1m by 6.1m) triple deck vibrating screen. See Figure 17.1

The process plant feed is ground in an Allis CHalmers 12 ½ foot by 14 foot (3.8m by 4.3m), 1,250 HP (933 kW) ball mill to 67% passing Tyler 200 mesh (74 microns). A portion of the process plant circulating load is passed through one of two 20” (500mm) gravity concentrators. Concentrate from these units is upgraded on an 8 foot (2.4m) shaking table. The table concentrate is direct smelted. Tails from the concentrators and the shaking table are returned to the head end of the grinding mill. The fines from the grinding circuit are fed to one of two rows of 10m3 OUTOTEC tank cells producing both a rougher concentrate grading between 5 and 10 opt (171 and 343 g/t Au) and a scavenger concentrate that is very low grade. The scavenger concentrate is circulated to the main grinding circuit. Rougher concentrate is collected and reground through an 8 foot by 6 foot (2.4m by 1.8m) ball mill to 98% passing Tyler 400 mesh (37 microns), thickened and leached in a three stage leach circuit of 12 foot by 24 foot (3.6m by 7.2 m) tanks. Dissolved gold is recovered using a six stage carbon-in-pulp circuit using 12ft. x 14ft (3.6m x 4.3m) vessels. The carbon is then eluted in a stainless steel pressure strip vessel. The elution liquor is passed through an electrowinning cell fitted with stainless steel anodes and cathodes. Gold sludge from this cell is then smelted in an electric induction furnace.

Current process plant recovery is name plated at 93.5% based on a feed grade of 0.16 opt (5.5 g/t). This recovery is grade dependent and has been as high as 96.5% with higher feed grades.

Past and current processing analyses have shown that the process plant feed is clean displaying no evidence of any deleterious constituents such as arsenic, mercury or antimony that would otherwise affect gold recovery in the leach circuit. Copper in solution is sometimes high.

The process plant operates on a 14 day on, 14 day off, 12 hour per day schedule utilizing four crews. See process plant production schedule in Table 17.1.

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An economic assessment is underway looking into leaching the flotation tails and solids from the tailing storage facility. According to the potential future flow sheet the existing flotation concentrate leach circuit would remain intact with flotation tails reporting to a new pre-leach thickener. Slurry would be pumped from the existing tailings storage facility over a new trash screen. The flotation tails and trash screen undersize would be combined and thickened using the new thickener. The thickener underflow would be pumped to a series of six CIL tanks for cyanide leaching. The CIP tails from the flotation concentrate leach circuit will be pumped to the new CIL circuit. Carbon would be transferred counter current from the slurry with the carbon from the first CIL tank being pumped either to the acid wash vessel or to the flotation concentrate CIP circuit depending upon operational conditions. Material from the tailings storage facility would be pumped back to the process plant for processing as weather allows. It is estimated that the tailings pumping would operate approximately six months of the year. During the winter months only the flotation underflow alone would report to the new thickener and then pumped to the new CIL circuit. The CIP tails from the flotation concentrate leach would continue to be pumped to the new CIL circuit.

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Figure 17-1 True North Gold Mine Process Plant Flow Sheet

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18.

Project Infrastructure

True North has been an active mine for almost 90 years except for some periods of inactivity. During this timeframe, the onsite infrastructure has been updated, upgraded and improved continuously by its respective historic owners. Figure 18.1 illustrates the current layout of the surface infrastructure.

Figure 18-1 Surface Infrastructure Plan View

18.1.

Location and Access

The Project is located adjacent to the town of Bissett Manitoba, which is 160 km northeast from Winnipeg Manitoba. Bissett and the Project are accessible via provincially maintained public roads connecting to Winnipeg. Bissett provides employee housing and support services to the Project.

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18.2.

Accommodations and Camp Facilities

The Project has a 300 room camp facility located near the main administration offices which includes a kitchen and dining facility, and recreation and fitness facilities. The majority of employees and contractors working on site are currently accommodated at this facility during shift rotations. In addition, the town of Bissett offers options for employee room and board.

18.3.

Electrical Power and On-Site Distribution

The Project is supplied power by Manitoba Hydro grid through two power lines which provides 20MW to the Project transformer station. The twin power line provides a redundancy such that in the event of a single line power outage, the mine, process plant and surface facilities can still function in a limited capacity on 10MW.

18.4.

Water Supply and Reticulation

Potable water is supplied from the town of Bissett’s water supply.

Process water for the mine is reclaimed from the tailings pond and water recovered from the underground workings.

18.5.

Air Compressors

Compressed air for the underground workings is provided by five 300hp (224kW) and two 150hp (122kW) compressors located in a central compressor house. The compressed air is distributed throughout the Project through a network of 10 inch and underground and plant smaller airlines

18.6.

Diesel Fuel and On-Site Storage Facility

Diesel fuel is supplied to the onsite storage tanks by commercial road tanker from a major fuel supplier’s central depot in Winnipeg. The diesel fuel for the underground machinery is transported from the onsite storage tanks to the underground SasStat fuel storage facility via fuel cars on the mine cage.

18.7.

Warehousing and Material Handling

The Project is serviced from a two-story, heated, 4,800 ft²(445m²) warehouse building, a 223 m²2,400 ft²(223 m²) cold storage area, as well as three cold storage tents and a 100,000 ft²(9,290 m²) secured yard storage. There is also a gravel surfaced storage area that is unsecured.

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18.8.

Site Security

The Company employs an external security contractor, who monitors the Project from a central security outpost at the main gate. There are also roaming security personnel. Currently, the Project is surrounded by chain link fencing.

18.9.

Communication

Voice and data communications are routed through the Bissett Manitoba Telephone System microwave tower. This tower also provides cell phone coverage for the Project and town site. On-site and underground communications is via a radio over leaky feeder network which is maintained and extended as need it by the Project personnel.

18.10.

On-Site Transport and Infrastructure

The Company provides bus transportation from Winnipeg to site on scheduled shift rotations. Light vehicles and pickups are provided on-site to transport mine workers from accommodations to their respective work areas.

18.11.

Solid Waste Disposal

Waste is managed in dumpsters and other appropriate waste containers. Waste and materials for recycling are disposed of off-site by an external contractor located in Pine Falls. Additionally, the external contractor removes waste hydrocarbons for disposal or recycling.

18.12.

Parts and Mine Supply Freight

All supplies and other consumables required to operate the mine, process plant and surface facilities are brought in via all season access road from Winnipeg, Manitoba by various freight-forwarding contractors.

18.13.

Mobile and Fixed Equipment Maintenance Facility

There are 5 maintenance bays, welding and tire facilities at the Project which have been upgraded by the previous owner to accommodate and provide an enclosed facility for all maintenance activities. This is especially useful during winter season when temperature can plunge as low as -30°F (-35 °C).

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18.14.

First Aid and Ambulance

The Project has a first-aid clinic, ambulance and trained personnel on stand-by for any medical attention or emergency that may arise. An air ambulance service is readily available from the nearby Winnipeg Emergency Rescue Service.

18.15.

Office and Administration Buildings

The Project hosts a recently constructed (by the previous operator) modern office and administration facility that can accommodate the necessary engineering, geological, accounting, safety, environmental, and administrative personnel.

18.16.

Tailings Storage

The Tailings Management Area (TMA) is located approximately 1 mile (1.6 km) north of the process plant in an area naturally defined by bedrock ridges around the perimeter of a previously flat boggy area. The original ground surface of the bog was near elevation 889 feet above sea level (asl) (271 m asl) (geodetic) with bedrock ridges on the south and west sides up to 920 feet above sea level (asl) (280 m asl) and bounded to the north by bedrock up to elevation 985 feet asl (300 m asl).

Since the development of the TMA, tailings have been pumped from the process plant to the TMA via an approximate 1 mile (1.6 km) pipeline. It is understood that during mine operation, the tailings are transported as slurry, with 34% (approx.) solids by weight. The TMA currently consists of 8 dykes with a number of the embankments separated by bedrock outcroppings such that they follow an A/B nomenclature. The embankments have been designed and constructed in various stages and phases from 1997 onwards to the most recent raises and improvements completed from 2012 to 2014. The current configuration of the TMA consists of a tailings pond and polishing pond, separated by dyke 7. The west half (approx.) of the tailings pond has reached its capacity, with tailings placed up to the crest of dykes 1, 2, 8 and a portion of dyke 3, while the east half of the tailings pond contains tailings submerged beneath water ranging in depth from less than 3 feet (1m) to several feet (metres). No spillway or low level outlet structures are present in the TMA. It is understood that the TMA has been designed to safely retain water from the mill discharge, runoff, and storm events.

In order to increase the capacity to retain tailings beyond the current capacity of the TMA, a new area, termed the East Tailings Management Area (ETMA) was currently undergoing development until the most recent mining operations ceased in 2015. The ETMA is located directly east of the TMA and currently consists of dyke 9 along its south perimeter, with dyke 6 of the TMA forming the containment along the west side. The natural contours to the north and east provide containment of the remainder of the ETMA.

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Dyke 9 has an overall length of nearly 5,000 feet (1,500m) and at its current constructed elevation, has a height of 10 feet to 13 feet (3m to 4 m). No spillway or low level outlet structures are present in the ETMA. It is understood that the TMA has been designed to safely retain water from the process plant discharge, runoff, and storm events.

A dam safety review was conducted by Stantec Consulting Ltd. geotechnical engineers in 2015.

18.17.

Stockpiles

The True North site has an existing waste rock stockpile which currently contains approximately 200,000 tons (180,000 tonnes) on an area of 4.6 acres (1.9 ha). This waste material is utilized to construct the tailings containment berms

The site is permitted to stockpile up to 10,000 tons (9,000 tonnes) of ore permanently and 50,000 tons (45,000 tonnes) of ore on surface as outlined in the 2016 temporary permit.

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19.

Market Studies and Contracts

19.1.

Precious Metal Markets

Gold and silver markets are mature with reputable smelters and refiners located throughout the world. As of March 2017, the 36-month trailing average gold price was US$1,217 per ounce, the 24-month trailing average price was US$1,207 while the monthly average had increased to US$1,265. The silver price trend shows similar behavior and both are shown in Figure 6 1 This report uses a gold price of US1200/oz (C$1,500/oz) for Mineral Reserves and US$1,400/oz (C$1,750) for Mineral Resources. An exchange rate of 0.8 Canadian dollars to one US dollar is also used.

Figure 19-1 Historical Monthly Average Gold and Silver Prices and 36 Month Trailing Average ($US/oz)

19.2.

Contracts

As part of normal mining activities, Klondex has entered into contracts with several mining industry suppliers and contractors. The terms of these agreements are customary for mines similar to the True North Project.

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20.

Environmental Studies, Permitting and Social or Community Impact

20.1.

Summary

The previous mine operator San Gold held an Environmental Act License covering mining, processing and tailings management area operations for the Project. San Gold also held an accepted Mine Closure Plan and had pledged certain fixed assets to provide financial security for closure. Klondex has since obtained a revised Environmental Act Licence for the Project and approvals of minor alternations required for its Project. The San Gold Mine Closure Plan (2012) and the pledged fixed-asset financial security for the mine closure plan were transferred to Klondex in January 2016.

Klondex has been conducting required environmental monitoring including water quality sampling and environmental effects monitoring work; developing procedures for its environmental management system; and is in the process of re-initiating First Nations and Aboriginal community engagement including final effluent release reporting.

In consideration of the historic activities and planned activities at this Project, the Author has reviewed historic and current information on the Project including: current legislation affecting mine permitting, operations and closure in Manitoba; the revised Environment Act License for mining, processing and tailings management recently issued to Klondex; relevant reports prepared as part of a harmonized federal-provincial environmental assessment for a tailings management expansion project approved in 2013 including public comments and First Nations and Aboriginal community consultation conducted at that time; and other information. The Author also contacted the Author of the metallurgical components of this report in regard to the process plant complex and Klondex’s Environmental Superintendent about the current environmental–social status of the Project.

Based on the available information, P&E is of the opinion that there do not appear to be any insurmountable environmental and/or social barriers to the Project.

20.2.

Scope of the Project

The scope of True North Gold Mine includes:

Underground mine development and production. True North includes six underground mines (i.e. Cohiba Zone, SG-1, 710/711 Zone, 007 Zone, Hinge Zone, Rice Lake), a vertical shaft, two decline ramps, a mill, an ore feed pad, mill feed crushing and conveying, a waste rock management area, and a tailings management area (“TMA”).

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Klondex commenced ramp-up activities in early 2016 including test stope mining and stockpiling in advance of a production decision. This underground work has included underground mining using conventional narrow vein longhole stoping methods at the 710/711 and Cohiba Zones. It is projected that underground mining would produce about 800 tpd using conventional drill and blast methods, rail load and haul technologies, and a vertical shaft and two decline ramps.

The re-processing of historic process plant tailings for gold and silver recovery commenced in 2016 with approval from the Province of Manitoba - historic tailings are being dredged, pumped to the process plant for treatment. The re-processed tailings are re-deposited in another section of the TMA.

20.3.

Ongoing Exploration and Project Development

Klondex acquired Mineral Lease ML-63 (formerly known as the Rice Lake Gold Mine) at Bissett, Manitoba in 2015 when the Project was inactive under a temporary suspension of operations. Mineral exploration and mining activities have been undertaken at the Project since about 1932 with ownership having changed numerous times over time.

In the present process plant flowsheet, the comminution stage includes crushing and a closed grinding circuit with hydrocyclones. The hydrocyclone overflow is directed to gravity concentrators and the resulting gravity concentrate is further processed using a shaking table with the table concentrate sent to the smelting furnace. The crushing circuit feeds the flotation circuit. The rougher concentrate is re-ground, dewatered and pumped to the carbon-in-leach (“CIL”) circuit. The rougher cell tailings are pumped to scavenger cells, and the scavenger concentrate is fed back to the grinding circuit. Rougher flotation concentrate is leached in a three tank leach circuit followed by a six stage CIL circuit. Loaded carbon is eluted in a conventional strip circuit and gold is electrowon from the eluate. An Inco SO2-air cyanide destruction process is used to treat CIL circuit wastewater. Natural degradation is used to destroy residual cyanide in the tailings pond water. The six-stage CIL circuit is currently being used as part of Klondex’s tailings re-processing project.

The present TMA includes a designed tailings storage containment area and a water polishing pond. Water quality is monitored through sampling and excess pond water is pumped and released to No Name Creek over a specified effluent release time line. TMA water pond levels, water quality, available water storage capacity and available freeboard are monitored by Klondex’s environmental staff.

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20.4.

Information Review and Assessment

20.4.1.

Documentation Reviewed

The documentation reviewed by the Author included:

The regulatory regime affecting mine permitting, operations and mine closure in Manitoba.

Revised Environment Act License 2628 RRR issued to Klondex Canada Ltd. for the “True North Gold Mine” on September 16, 2016, and Klondex’s April 2016 request for a minor alteration to allow early discharge from the East Tailings Pond and tailings re-processing – the latter document provides East Tailings Pond water quality data for a February 18, 2016 water sampling event. The QP also reviewed historical Environment Act License 2628 R which applied to Rice Lake Gold Corporation’s “Bissett Gold Mine” operations in 2004.

Relevant parts of the Environmental Assessment Proposal (EAP) filed in 2012 for a Class 2 development comprised of the expansion and operation of the TMA. That development included the construction of an additional main tailings pond, a polishing pond and three access roads. Treated water from the new polishing pond would be pumped to the existing polishing pond for discharge to No Name Creek from June 15 to November 30. The Author also reviewed comments on the EAP received from:

	•	Environment Canada, the Canadian Environmental Assessment Agency and Health Canada;
	•	Manitoba departments and branches including Manitoba Conservation & Water Stewardship, Climate change and Environmental Protection Division, Mines Branch, Community Planning Services, Sustainable Resource and Policy Management Branch, Aboriginal Relations, Workplace Safety and Health Division;
	•	The Kookum’s of Hollow Water First Nation and a trapper from the Hollow Water First Nation; and
	•	The Wanipigow Lake East End Cottager’s Association.

Environmental Act License 2628 RR issued in 2012 allowed for the construction and operation of the East TMA. Stage 1 of the East TMA was completed in November 2014 and provided a year of tailings storage capacity based on a process plant throughput of 2,500 tons per day (2,268 tonnes per day).

A 2010 Notice of Alteration for San Gold’s Cartwright Mine and Hinge Zone Bulk Sample Collection submitted to Manitoba Conservation. It included an assessment of environmental impacts and proposed mitigation measures in regard to air, noise, runoff and wastewater.

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Other information describing the existing infrastructure, environment, and the Project.

20.4.2.

Licenses, Permits and Approvals

The licenses, permits and approvals obtained to operate the Project are shown in Table 20.1.

Table 20-1 Obtained Licenses and Key Permits and Approvals

License/Permit/Approval		Act/Regulation	Description	Issued to

	Manitoba
License 2628 RRR.	Sustainable		Environmental Act	Klondex
Minor alteration.	Development	Environment Act.	License – main	(September
	Environmental		license.	2016)
	Approvals

	Manitoba		Early discharge of
	Sustainable		the East Tailings	Klondex
	Development		Pond and tailings
Minor alteration.	Environmental	Environment Act.	reprocessing.
	Approvals			(May 2016)

	Manitoba
	Sustainable		Ore stockpile	Klondex
Minor alteration.	Development	Environment Act.	increase to 50,000
	Environmental		tons.	(May 2016)
	Approvals

	Manitoba
	Sustainable		Tailings	Klondex
Minor alteration.	Development	Environment Act.	reprocessing –
	Environmental		trucking tailings.	(August 2016)
	Approvals

	Manitoba	Water Rights Act,
Water Rights License 2016	Sustainable	Water Rights	License to use
– 003.	Development Water	Regulation.	water from lake.	Klondex
	Licensing

	Manitoba	Hazardous Waste
	Sustainable	Regulation,
Hazardous Waste	Development	Dangerous Goods	Hazardous waste
registration.	Environmental	Handling &	registration.	Klondex
	Services	Transportation
		Regulation.

Petroleum Storage Facility	Manitoba	Storage and Handling	Above ground
Permit	Sustainable	of Petroleum Products	storage tanks with
	Development	and Allied Products	a capacity of 5,000	Penner Oil
	Manitoba	Regulation, Technical	L or more
	Conservation	Bulletins.
	Environmental
	Services

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License/Permit/Approval		Act/Regulation	Description	Issued to
			Ventilation raise
Crown Lands Permit			building situated	Rice Lake
	Crown Lands and		within the Town	Gold
GP0003073.	Property Agency	Crown Lands Act.	of Bissett	Corporation*

Crown Lands Permit				Rice Lake
	Crown Lands and	Crown Lands Act.	TMA	Gold
GP0005737.	Property Agency			Corporation*

*Klondex has requested the permit holder name be changed to Klondex Canada Ltd. Source: Klondex Canada Ltd.

20.4.3.

Revised Environmental License and Minor Alterations

Revised Environmental Act License 2628 RRR

The Environmental Stewardship Division, Environmental Approvals Branch of Manitoba Sustainable Development issued revised Environmental Act License No. 2628 RRR to Klondex Canada Ltd. on September 16, 2016 for the operation of the “Development” being a 2,273 tonnes per day (2,500 tons per day) gold and silver mining, processing and refining operation known as the True North Gold Mine and including the existing and expanded TMA. Plans of mine, process plant, and the TMA are shown in Figures 20.1 and 20.2. Figure 20.3 provides an aerial view of the mine, process plant and TMA.

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Figure 20-1 True North Gold Mine Site Plan

Source: Revised Environmental Act License No. 2628 RRR for the True North Gold Mine, Appendix B

Figure 20-2 TMA Site Plan

Revised Environmental Act License No. 2628 RRR for the True North Gold Mine, Appendix A

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Figure 20-3 Aerial View of the Mine, Plant Site and TMA

Source: Google Maps

Minor Alterations

Section 14 of The Environment Act requires notification and approval for alterations to a licensed Development. A notice of alteration submitted by a License holder is assessed by the Director as either minor, having insignificant environmental effects, or major, having significant environmental effects. Minor alterations may be approved through a revised Environment Act License or by a letter from the Director for Class 1 and 2 projects. Recent examples of requested alterations and approvals are provided below.

In March 2016, the Director of Manitoba Mineral Resources approved Klondex’s notice to recommence mining and processing operations. Since receiving approval Klondex has undertaken a range of preparatory works such as shaft guide replacement, underground track repair, narrow vein longhole layout testing, development of a new underground shop and mining gear storage cut-outs, and metallurgical test work on historical tailings based on using the existing flow sheet.

In April 2016 Klondex issued a notice of alteration to the Development as licensed for early discharge of effluent (to commence on May 9, 2016) from the East TMA and the implementation of a tailings re-processing project. The ETP contained about 780,000 m3 of water which met effluent discharge criterion. The proposed tailings re-processing alteration included the construction and operation of a tailings dredging system, a pumping station, booster pumps and a pipe line to be used to pump the tailings slurry to the existing process plant for re-processing. It was projected that 80,000 m3 to 150,000 m3 of tailings would be dredged in 2016 commencing on June 1st. The regulator determined that the potential effect of the proposed alteration would be a minor in accordance with the Environment Act and approved the proposed alteration.

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In May 2016 Klondex issued a notice of alteration for a one-year temporary increase in the ore stockpile limit from 10,000 tons to 50,000 tons (9,000 tonnes to 45,000 tonnes). The regulator determined that the potential effect request was a minor alteration in accordance with the Environment Act and approved the alteration.

20.4.4.

Current Status / Mitigative Measures

The current status of the Project as well as associated environmental aspects and mitigative measures are summarized in Table 20.2 based on the information obtained and P&E’s experience at other mine properties internationally.

Table 20-2 Potential Significant Environmental Impacts and Current Status / Mitigative Measures

Area	Current Status / Mitigative Measures

Environmental Management System	The revised Environmental Act License requires Klondex to establish and implement an Environmental Management System (“EMS”). An EMS is a comprehensive system that would be expected to require without being limited to the development and communication of an environmental policy, the identification of significant environmental aspects, the identification of legal and other requirements, procedures, training, records, change management, consultation and complaint response, monitoring, EMS and compliance reviews, a corrective and preventative measures procedure to deal with a non-conformance, and emergency preparedness and response planning.

	As part of other conditions of the environmental license it is expected that the EMS would also include solid waste reduction and recycling efforts; contingency plans for spills, ruptures and unexpected TMA seepage losses; and require spill recovery equipment. In addition, solid waste and hazardous wastes are to be disposed of in accordance with regulatory requirements; petroleum products are to be stored in accordance with regulatory requirements; and the sewage management system is subject to the Onsite Wastewater Management Systems Regulation.

Acid Rock Drainage	Test results included in the documentation reviewed by the Author indicate that waste rock and tailings are not acid generating. The revised environmental license requires ongoing scheduled acid:base account testing.

Final effluent	Klondex is to reclaim as much water as possible from the TMA to supply the process water demands of the mill.

	Mine water is directed to the TMA. Treated effluent can only be released from the TMA polishing pond to No Name Creek and subsequently to the Wanipigow River between June 15 and November 30 each year at a rate not to exceed 0.20 m³/sec. Treated effluent cannot be released if the quality or toxicity of the effluent results in, or is likely to directly or cumulatively results in, a downstream water quality degradation beyond a maximum 10% mixing zone (by volume) within No Name Creek and/or the Wanipigow River relative to the Manitoba Water Quality Standards, Objectives and Guidelines Regulation under the Water Protection Act.

	Elevated levels of ammonia in mine water / polishing pond water occurred in years past possibly in part as a result of the dissolution of mine explosives and blasting agents. Best practices including improved blasting practices and reducing / avoiding ANFO use are now used to help avoid this potential issue.

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Area	Current Status / Mitigative Measures
Air emissions	Klondex would maintain its diesel-powered equipment and mine air heater.

	As required by the environmental license: distinct plume forming fugitive emissions are not to exceed 5% opacity whilst non-plume forming fugitive emissions are to be not visible. Downwind off Project, point of impingement suspended particle matter ground level concentrations are not to exceed a 24 hour average of 120 µg/m³or an annual geometric mean of 70 µg/m³.

Cyanide transportation and storage	Cyanide transfer, storage and mixing activities would be conducted in conformance with regulatory requirements and Klondex’s procedures and EMS requirements.

Tailings management	Tailings from the original San Antonio Gold Mine were discharged into Rice Lake from about 1932 to 1968. Tailings produced when the mine was reopened in 1981 to 1983 were placed in a containment constructed over the previously disposed tailings. The TMA is located north north-east of the mine and plant site and includes a tailings pond and a polishing pond. The final treated effluent is pumped and annually released to No Name Creek which flows to the Wanipigow River. The historic tailings that are being re- processed as part of the Project are located in the TMA. The revised Environmental Act License 2628 RRR requires Klondex to engage the services of licensed professional geotechnical engineers for engineering and quality control during dyke construction and submit a construction performance and quality control report to the Director for approval.

	Klondex’s Environmental Superintendent would monitor dykes and assess conditions / geotechnical monitoring data with input from competent geotechnical engineers.

Waste rock storage	New waste rock is to be stockpiled in the designated “waste rock stockpile area”. Ore is to be stored in the designated “ore rock stockpile area”. The License also requires the company to conduct acid:base accounting testing as indicated above.

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Area	Current Status / Mitigative Measures

Environmental monitoring	Klondex uses the existing TMA and the water polishing ponds to manage surface water storage / release. An Inco SO2-air process and natural degradation continue to be used for cyanide destruction. TMA water management controls include polishing pond levels, water quality monitoring and a surface water management program.

	Surface water samples (i.e. mine water samples to be collected from the tailings ponds, polishing pond, treated effluent, downstream receiving water quality sampling stations) are to be sampled at frequencies and for parameters specified in the License while groundwater quality is to be monitored at specified groundwater wells and at additional wells as may be requested by the Director. Treated effluent toxicity testing is also required. Sediment core samples are to be collected at two downstream water quality sampling station locations and analyzed for total metals, total organic carbon, moisture content and pH.

	Klondex will continue to conduct scheduled downstream water quality sampling, sediment sampling, and environmental effects monitoring. Klondex has undertaken two environmental effects monitoring studies and has scheduled a third.

Solid waste	Solid non-hazardous waste that is not re-used / recycled would be disposed in an off-site licensed solid waste landfill.

Hazardous waste	Hazardous waste would be disposed of in accordance with regulatory requirements.

	Klondex is currently working to confirm that there are no electrical transformers that contain PCB in use or stored on the Project and that asbestos had been removed several years ago from all surface buildings with the exception of one secured unused old building.

Terrestrial and avian wildlife	Klondex is aware of its responsibilities to protect wildlife. It is expected that this would be reflected in the EMS procedures.

Social consultation	Klondex has as a priority re-initiated community engagement activities with local First Nations, the Town of Bissett, other interested communities and persons, and regulatory authorities.

20.4.5.

Community Engagement

As part of 2012 TMA expansion project EAP, the Manitoba Mines Branch requested that First Nations and Aboriginal communities whose traditional activities could be impacted be identified and engaged and that community issues be incorporated into the environmental assessment for the TMA. Winsor (2013) reported on the outcome of that consultation process and reported that:

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Most of the concerns from constituents were a result of misinformation.

The Manitoba Mines Branch had determined that there are three Anishinaabe (Ojibwa) First Nations situated within an 80 mile (130 km) radius of the mine: the Hollow Water First Nation, the Little Black River First Nation, and the Sagkeeng First Nation. The Hollow Water First Nation is situated downstream of the confluence of No Name Creek and the Wanipigow River which flows to Lake Winnipeg.

It was recommended that an Environment Act Licence be issued to San Gold in 2013 for the proposed TMA expansion subject to it accommodating community concerns and issues. As such, the draft License included a clause requiring the operator to submit an environmental monitoring report to the Hollow Water First Nation after each effluent discharge campaign summarizing monitoring data and impacts on the receiving waterways.

The Hollow Water Chief and Council encouraged direct negotiations between San Gold and two trappers. San Gold negotiated a confidential compensation settlement with one of the trappers (Trap Line #11) for loss of opportunity to trap in the proposed TMA expansion area, and at the time of reporting was negotiating a settlement with the second trapper (Trap Line #12).

San Gold had met with Hollow Water First Nation residents in June 2012 and discussed the potential impacts of the proposed TMA expansion. San Gold participated in the Hollow Water First Nation’s Traditional Area Advisory Committee (TAAC). The Kookom’s who opposed the TMA expansion did not have official standing in the community. San Gold had attempted to arrange an information meeting with the Little Black River First Nation situated in the O'Hanley and Black Rivers area on the eastern shore of Lake Winnipeg.

Klondex is in the process of re-initiating community engagement activities and aims to maintain contact and communication with local First Nations and Aboriginal communities, and continue to sponsor community events and undertakings and recruit and train local First Nations and Aboriginal workers. The Town of Bissett is also kept informed of environmental matters that could potentially impact residents or community services. Klondex participates in Town of Bissett council meetings and has held community information sessions.

20.4.6.

Mine Closure

Mine closure planning and financial security provisions that apply to advanced exploration and mining projects are described in the Mine Closure Regulation (MR 67/99) under The Mines and Minerals Act (C.C.S.M. c.M162).

Responsibility for the accepted San Gold Corporation Mine Closure Plan (2012) for Mineral Lease ML-63 which includes pledged fixed assets financial security was transferred to Klondex

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Canada Ltd. on January 13, 2016 by the Mines Branch Director (Manitoba, 2016). The Author reviewed sections of the Mine Closure Plan (2012) and estimated mine reclamation and rehabilitation costs presented in Gibson (2015). The estimated reclamation and rehabilitation costs amounted to $4.4M. There is a possibility that the Mines Branch may require an alternate form of financial security at some point in the future when the closure plan is updated.

As indicated in Section 10 “Expected Site Conditions” of the closure plan, the Project would be rehabilitated to a predevelopment state as a wilderness area with primarily conservation and recreational value. An additional crown pillar assessment would be completed at close-out to assure surface stability. The surface of the tailings in the TMA would be revegetated at closure. Pond water quality would continue to be monitored and excess water would be pumped / released to No Name Creek until the pond water quality is shown to improve to a level whereby a weir system could be used to direct excess run-off to No Name Creek. Environmental monitoring would continue to be conducted through each stage of closure to ensure that the mine remains compliant with environmental and safety requirements.

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21.

Capital and Operating Costs

21.1.

Capital Costs

Life of Mine (LOM) constant dollar capital expenditures are detailed in Table 21-1. Mine development comprises 55% of total capital requirements; tailings reprocessing 22% and site infrastructure 10%. Mine development unit costs are shown in Table 21-2.

Table 21-1 Capital Costs

			Cost (000's)
				2020
	2017¹	2018	2019	-2028	Total
Mine Development	$6,045	$9,350	$156		$15,551
Capital Drilling	$1,237	$940			$2,177
Mobile Equipment	$481	$365			$846
Site Infrastructure and Plant	$1,589	$1,208			$2,797
Underground Services	$287	$217			$504
Tailings Reprocessing	$511	$735	$400	$4,500	$6,146
Total	$10,150	$12,815	$556	$4,500	$28,021

1. 2017 includes only April through December estimates.

Table 21-2 Underground Development Unit Costs

	Width	Height	Unit Cost
Description	(ft)	(ft)	($/ft)
Capital Drifting	12	12	$1.104
Capital Raising	8	8	$782

Sustaining capital for the nine years following mine closure when historic tailings reprocessing will continue, on a stand-alone basis, will be approximately $400,000 per year. In addition, an allowance for winterization of the processing plant and reopening in the spring has been included at $100,000 per year. These total annual $500,000 costs are included in Table 21-1.

Mine closure & reclamation costs are not included in Table 21-1. They are estimated by Klondex to be $4.4 million commencing in 2026-2028.

21.2.

Underground Operating Costs and Cut-off Grade

LOM operating costs during the period of underground operations are presented in Table 21-3 through Table 21-6 below. Costs were estimated by Klondex using local labor and commodity prices, historic commodity consumption and labor productivities. Contracted rates were used where appropriate. Labor and related payroll costs in all categories is treated as a fixed cost.

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Contractor costs and consumables are classified as variable. PM has reviewed the operating cost estimate and finds it to be reasonable for a small scale underground narrow vein mine.

Table 21-3 Direct Underground Mining Cost

Description	Fixed Cost ($000’s/year)	Variable Cost ($/ton)
Longhole Stoping	-	$35.77
Stope Development Drift	-	$59.33
Material Movement	$1,043	$1.08
Direct Mining Cost	$1,043	$43.90 ¹
A-Hoist	$1,436	$1.91

1. Weighted average of longhole stoping and stope development drifting.

Table 21-4 Indirect Underground Mining Cost

Description	Fixed Cost ($000’s/year)	Variable Cost ($/ton)
Underground Support Services	$570	$9.19
Underground Dewatering	$113	-
Mobile Maintenance	$796	$5.11
Electrical Maintenance	$628	$4.08
Geology	$1,275	$17.22
Engineering/Survey	$1,028	-
Indirect Mining Cost	$4,408	$35.60

Table 21-5 Processing Cost

Description	Fixed Cost ($000’s/year)	Variable Cost ($/ton)
Operation Administration	$2,732	$9.15
Maintenance Administration	$1,117	$0.83
Consumables	-	$5.24
Plant Maintenance	-	$5.98
Processing Cost	$3,849	$21.20

Table 21-6 General and Administrative Cost

Description	Fixed Cost ($000’s/year)	Variable Cost ($/ton)
Site Administration	$8,746	-
Surface Support Services	$2,114	-
Health and Safety	$1,020	-
Environmental	$269	-
IT	$335	-
General and Administrative	$12,524	-
Corporate Allocations ($2.44%)	$542	$2.50

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Using the fixed and variable costs listed above the total cost sensitivity as a function of plant throughput has been calculated and is shown in Figure 21-1.

Figure 21-1 True North Total Cost Sensitivity to Production Rate

In 2019, the underground mining operation will cease, however, the processing plant will continue to treat tailings at a rate of 195,000 tons per year (1,200 tpd) as a seasonal operation from May to October. From July to September 2016 Klondex processed 33,434 tons of historic tailings to determine operating parameters and costs. The costs shown in Table 21-7 are derived from the actual costs incurred during the tailings reprocessing test. General and administrative costs will reduce to $1.24M per year with the reduction in site personnel and services.

Table 21-7 Tailings Reprocessing Cost

Description	Fixed Cost ($000’s/year)	Variable Cost ($/ton)
Operation Administration	$884	-
Maintenance Administration	$249	-
Consumables	-	$3.99
Other Costs	$516	-
Processing Cost	$1,649	$3.99
General and Administrative	$1,248	-
Sustaining Capital	$500	-
Corporate Allocations (2.5%)	$46	$0.10

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21.3.

Cut-off Grade Calculations

21.3.1.

Underground Mining Cut-off Grade

Using the operating costs and parameters above, cut-off grades were calculated at varying gold prices. These are shown in Table 21-8 and Figure 21-2. The incremental cut-off represents the required minimum grade of mineralization to be profitable to hoist and process after it has been mined. Mineralization from development excavations is included in the LOM plan if it exceeds the incremental cut-off since processing the incremental material improves the Project cash flow over the alternative using this material as backfill.

Table 21-8 Mine Cut-off Grade Calculation

Gold Sales Price	$/Ounce	$1,500
Refining and Sales Expense	$/Ounce	-
Royalty		0%
Metallurgical Recovery		94%
Total Cost	$/ton	$212
Incremental Cut Off Grade		0.08
Cut-off Grade	opt	0.15
Minimum Mining Width	feet	5.5
Grade Thickness cut-off	Eq. opt-ft.	0.825

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Figure 21-2 Cut-off Grade Sensitivity to Gold Price

21.3.2.

Tailings Reprocessing Cut-off Grade

Using the operating costs and parameters for tailings reprocessing, cut-off grades were calculated at varying gold prices. These are shown in Table 21-9 and Figure 21-3. The incremental cut-off represents the

Table 21-9 Tailings Reprocessing Cut-off Grade Calculation

Gold Sales Price	$/Ounce	$1,500
Refining and Sales Expense	$/Ounce	-
Royalty		0%
Metallurgical Recovery		89%
Total Operating Cost	$/ton	$19.32
Sustaining Capital	$/ton	$2.56
Total Cost	$/ton	$21.88
Cut-off Grade	opt	0.016

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Figure 21-3 Tailings Reprocessing Cut-off Sensitivity to Gold Price

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22.

Economic Analysis

The LOM plan and technical and economic projections in the LOM plan model include forward looking statements that are not historical facts and are required in accordance with the reporting requirements of the Canadian Securities Administrators. These forward-looking statements are estimates and involve risks and uncertainties that could cause actual results to differ materially.

The estimates of capital and operating costs have been developed specifically for the Project and are summarized in Section 16. These costs are estimated by Klondex using historical information along with local labor rates and commodity pricing and where appropriate include adjustments applicable to the planned production rates.

The cash flow estimate includes only costs, taxes and other factors applicable to the project and corporate obligations, financing costs, and taxes are excluded. The cash flow estimate includes 15% Federal income tax, 12% Manitoba Provincial Income Tax and 10% Manitoba Mining Tax after appropriate deductions for depreciation. No consideration has been given for carry forward losses incurred prior to 2017.

Reclamation costs of $4.4M are realized in the last three years of the project. All revenues and costs stated in the economic analysis are in Canadian Currency unless stated otherwise.

22.1.

Life of Mine Plan and Economics

Constant dollar cash flow analysis of the reserves production and development plan shown in Table 16-2 is presented in the income and cash flow statements of Table 22-1 and Table 22-2, respectively. Table 22-3 lists the life of mine key operating and financial indicators. The low capital requirements over the mine life results in a high profitability index (PI) of 1.8 calculated with an 10% discount rate and a 10% NPV of $19M. PI is the ratio of payoff to investment of a proposed project. It is a useful tool for ranking projects because it allows quantification of the amount of value created per unit of investment. A profitability index of one indicates break even.

Table 22-1 Income Statement 2015 – 2018 ($000’s)

Year	2017¹	2018	2019	2020	2021-2029	Total
Income Statement (000's)
Revenue
Gold Sales	$33,907	$76,698	$46,173	$5,780	$42,767	$205,323
Operating Costs
Mining	($10,734)	($21,740)	($10,868)	-	-	($43342)
A-Hoist	($743)	($1,590)	($833)	-	-	($3,166)
Processing	($4.625)	($9,196)	($5,503)	($2,434)	($18,034)	($39,791)
Site Administration & Overhead	($9,436)	($12,255)	($12,255)	($1,248)	($9,776)	($45,509)

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Year	2017¹	2018	2019	2020	2021-2029	Total
Corporate Allocations	($638)	($1,126)	($743)	($92)	($695)	($3,295)
Total Operating	($25,538)	($45,050)	($29,729)	($3,682)	($27,810)	($131,808)

General & Administrative
Refining & Sales (Included with Processing
Costs)	-	-	-	-	-	-
Royalty	-	-	-	-	-	-
Total Cash Cost	($25,538)	($45,050)	($29,729)	($3,682)	($27,810)	($131,808)
EBITA	$8,369	$31,648	$16,444	$2,098	$14,957	$73,515
Reclamation Accrual	($727)	($1,644)	($990)	($124)	($917)	($4,400)
Depreciation	($1,176)	($9,525)	($6,005)	($811)	($10,003)	($28,020)
Total Cost	($27,940)	($56,619)	($36,723)	($4,617)	($38,729)	($199,259)
Pre-Tax Income	$5,966	$20,479	$9,449	$1,163	$4,038	$41,096
Federal and Provincial Tax	($1,975)	($8,576)	($3,942	-	($1,333)	($15,826)
Net Income	$3,992	$11,903	$5,507	$1,163	$2,705	$25,269

2017 includes only April through December.

Table 22-2 Cash Flow Statement 2015 – 2019 ($000’s)

Year	2017¹	2018	2019	2020	2021-2029	Total
Net Income	$3,992	$11,903	$5,507	$1,163	$2,705	$25,269
Depreciation	$1,176	$9,525	$6,005	$811	$10,003	$28,020
Reclamation	$727	$1,644	$990	$124	$917	$4,400
Working Capital (6 weeks)	($2,947)	($2,251)	$1,769	$3,005	$425	$0
Operating Cash Flow	$3,448	$20,820	$14,269	$5,103	$9,649	$53,289
Capital Costs
Capital Development	($6,045)	($9,350)	($156)	-	-	($15,551)
Mine and Plant	($3,593)	($2,730)	-	-	-	($3,526)
Tailings Reprocessing	($511)	($735)	($400)	($500)	($4,000)	($6,146)
Total Capital	($10,149)	($12,815)	($556)	($500)	($4,000)	($28,020)
Net Cash Flow	($6,701)	$8,005	$13,713	$4,603	$5,649	$25,269
Cumulative Cash Flow	($6,701)	$1,305	$15,017	$19,620	$25,269

2017 includes only April through December.

Table 22-3 Key Operating and After Tax Financial Statistics

Material Mined and Processed (kt)	434
Avg. Gold Grade (opt)	0.242
Contained Gold (koz)	105
Avg. Gold Metallurgical Recovery	94%
Recovered Gold (koz)	99
Reserve Life (years)	3.0
Tailings Reprocessed (kt)	1,950
Avg. Gold Grade (opt)	0.022

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Contained Gold (koz)	43
Avg. Gold Metallurgical Recovery	89%
Recovered Gold (koz)	38
Cash Cost ($/oz).	$963
Total Cost ($/oz)	$1,168
Gold Price ($/oz)	US$1,200 C$1,500.00
Capital Costs ($ Millions)	$28.0
Payback Period (Years)	1.9
Cash Flow ($ Millions)	$25.3
5% Discounted Cash Flow ($ Millions)	$21.9
10% Discounted Cash Flow ($ Millions)	$19.0
Profitability Index (10%) ¹	1.8
Internal Rate of Return	127%

Notes:

Profitability index (PI) is the ratio of payoff to investment of a proposed project. It is useful for ranking project as a measure of the amount of value created per unit of investment. A PI of 1 indicates break even.

22.2.

Sensitivity Analysis

The Project’s net present value at 5% and 10% (NPV), IRR and profitability index from the cash flow model presented above were analyzed for sensitivity to variations in revenue, operating and capital cost assumptions. This analysis is presented graphically in Figure 22-1 through Figure 22-4 below. The mine exhibits the greatest sensitivity to gold price and slightly less to operating cost. Capital costs exhibit the least sensitivity with the project remaining profitable at up to 40% capital cost increases.

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Figure 22-1 5% NPV Sensitivity

Figure 22-2 10% NPV Sensitivity

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Figure 22-3 Internal Rate of Return Sensitivity

Figure 22-4 Internal Rate of Return Sensitivity

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23.

Adjacent Properties

There are a number of mineral exploration properties adjacent to Klondex’s True North (Figure 23.1), but most appear to be inactive.

Gold Pocket Resources Ltd. owns the Bissett Project exploration property south adjacent to True North. Maps on their website show numerous gold mineralized zones, drillhole collar locations, and historic shafts. In 1998, Golden Pocket drilled 131 diamond holes for a total of (68,652 feet) 20,925 metres. The drilling returned high grade gold intersects, particularly from the Nevada Zone. Currently, Golden Pocket has sufficient assessment credits to keep the Bissett Property in good standing.

Bison Gold Exploration Inc. owns the Cryderman Central property about 15 miles (25 km) to the southeast of True North, but adjacent to Klondex owned ground in that area. An NI 43-101 technical report dated November 15, 2013 for the Ogama-Rockland Mine deposit lists Inferred Mineral resources of 1.28 million tons (1.16 million tonnes) grading 0.24 opt Au (8.17 g/t Au). The style of gold mineralization is dominated by gold-bearing quartz-carbonate veins associated with shear zones in granite host rocks.

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Figure 23-1 Adjacent Properties to the True North Gold Mine

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24.

Other Relevant Data and Information

Practical Mining and P&E are not aware of any other relevant data or information as of the effective date of this Technical Report.

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25.

Interpretation and Conclusions

This Technical Report demonstrate the success of Klondex’s exploration and development drilling programs since acquiring the True North Mine. Both Mineral Resources and Mineral Reserves have been increased substantially in the past nine months.

The Project hosts a Proven and Probable underground Mineral Reserve of 434,000 tons at an average grade of 0.242 opt (8.30 g/t), at a 0.15 opt (5.14 g/t) Au cut-off grade. In addition, the tailings reprocessing project contains a Mineral Reserve of 1.95 million tons at an average grade of 0.022 opt Au (0.75 g/t Au).

The operation of the Project appears to be financially and technically sound. Underground development and tailings reprocessing are the two largest capital items. Plant, equipment and infrastructure are complete with only minor sustaining capital required during the life of the underground operation. The mining and processing methods in use at True North have been proven effective by the history of previous operations. Geotechnical risks are limited as shown by the large extent of underground workings.

Klondex is aware of the importance of an effective community engagement process to the Project. Klondex is currently in the process of re-initiating community engagement activities with local Aboriginal communities, the Town of Bissett, other interested stakeholders, and regulatory authorities, on a priority basis.

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26.

Recommendations

This Technical Report and describes a viable mining and processing operation at the True North Gold Mine.

Technically, the Project presents no fatal flaws and due to the minimal capital requirements and history of underground mining it is relatively low risk..

It is recommended that the Project continue with its current production plan with a combination of longhole mining methods as detailed in Section 16 and the on-site processing of ore. It is also recommended that Klondex continue with its plans to reprocess of the existing tailings on site.

Specifically, it is recommended that Klondex take the following actions to develop and operate the Project:

Geology

	1.	Technical Database: All True North project data collected need to be stored and archived in a permanent and reliably retrieval manner. A full-time database administrator is recommended.

	2.	Quality Assurance/Quality Control: Timely follow-up for any and all QA/QC assay deviations and re-assay requests should be performed in a timely manner. The process should be automated when the database is up and running.

	3.	Sample Storage and Retrieval: Half-core remaining from sample assays should be retained for reference and check assay purposes. All assay sample rejects and pulps should be stored in a safe, secure and sheltered manner and properly catalogued to ease retrieval.

	4.	Project Assay Lab: Standard operating procedures should be updated, particularly in regard to assay data generation, storage and retrieval.

Environmental and Mine Closure

It is recommended that Klondex review the technical basis of the TMA closure approach presented in the 2012 mine closure plan and update the associated closure costs. A provisional amount for a $250k study that would be carried out over four years commencing in 2016 is recommended. This amount has not been included in the cash flow model presented in this Technical Report. This exercise will review and confirm the technical basis of the proposed TMA closure plan and estimated costs and possibly identify opportunities to improve upon the currently proposed approach.

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Mine Operations and Planning

Other

PM and P&E also recommend that other exploration targets in the area continue to be identified and investigated to provide supplemental process plant feed in the future.

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27.

References

Ames, D. E., Franklin, J.M and Froese, E. 1991. Zonation of hydrothermal alteration at the San Antonio Gold Mine, Bissett, Manitoba, Canada. Economic Geology 86, 600-619.

Anderson, S.D. 2008. Geology of the Rice Lake area, Rice Lake greenstone belt, south-eastern Manitoba (Parts of NTS 52L13, 52M4), Manitoba Science, Technology, Energy and Mines, Manitoba Geological Survey, Geoscientific Report GR2008-1, 97p.

Anderson, S.D. 2011. Geology and structure of the Rice Lake Mine trend, Rice; greenstone belt, south-eastern Manitoba (part of NTS 52M4) Manitoba Innovation, Energy and Mines. Preliminary Map PMAP2011-3, 2011.

CMG Airborne 2011. Report on a Helicopter-Borne Magnetic Gradiometer and Radiometric Survey Bissett Mine; Project 2011-004; for San Gold Corporation; Dec. 21, 2011.

Ginn, D. and Michaud, M. 2013. Technical Report on the Rice Lake Mining complex, Bissett, Manitoba.

Gibson, 2015. Exhibit “J” referred to in an affidavit of Greg Gibson dated January 16, 2015. Available under San Gold Corporation at the MNP Ltd. website – www.mnpdebt.ca. January 16, 2015.

Groves, D.I., Goldfarb, R.J., Gebre-Mariam, M., Hagemann, S.G. and Robert, F. 1998. Orogenic gold deposits: a proposed classification in the context of their crustal distribution and relationship to other gold deposit types. Ore Geology Reviews 13, p. 7–27.

Manitoba Mines Branch, 2016. Consent to Closure Plan, Mineral Lease ML-63. January 13, 2016.

Manitoba Sustainable Development, 2016. Revised Environment Act Licence No. 2628 RRR, revised September 16, 2016. Issued to Klondex Canada Ltd. Available at www. gov.mb.ca. September 16, 2016.

Poulsen, K.H., Robert, F. and Dube, B. 2000. Geological classification of Canadian gold deposits. Geological Survey of Canada Bulletin 540. 106p.

Ravenelle, J.-F. and Fonseca, A. 2013. Structural controls on gold mineralization at the and surrounding area. Unpublished report for San Gold Corporation by SRK Consulting (Canada) Inc., 77p.

Ross, K. and Rhys, D. 2010. Petrographic and SEM study of vein samples from the and Hinge Zone, Bissett, Manitoba. Unpublished report for San Gold Corporation by Panterra Geoservices Inc., 50p.

San Gold Corporation 2015. DRAFT – Rice Lake Mining complex – Summary of Mineral Resource Estimation. April 2015.

Winsor, Jennifer, 2013. Summary of Comments/Recommendations for the San Gold Tailings Management Area Expansion. File No. 2435.40. Available at www.gov.mb.ca. July 9, 2013.

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28.

Glossary

Assay: The chemical analysis of mineral samples to determine the metal content.

Asbuilt: (plural asbuilts), a field survey, construction drawing, 3D model, or other descriptive representation of an engineered design for underground workings.

Composite: Combining more than one sample result to give an average result over a larger distance.

Concentrate: A metal-rich product resulting from a mineral enrichment process such as gravity concentration or flotation, in which most of the desired mineral has been separated from the waste material in the ore.

Crushing: Initial process of reducing material size to render it more amenable for further processing.

Cut-off Grade (CoG): The grade of mineralized rock, which determines as to if it is economic to recover its gold content by further concentration.

Dilution: Waste, which is unavoidably mined with ore.

Dip: Angle of inclination of a geological feature/rock from the horizontal.

Fault: The surface of a fracture along which movement has occurred.

Footwall: The underlying side of a mineralized body or stope.

Gangue: Non-valuable components of the ore.

Grade: The measure of concentration of valuable minerals within mineralized rock.

Hanging wall: The overlying side of a mineralized body or stope.

Haulage: A horizontal underground excavation which is used to transport mined rock.

Igneous: Primary crystalline rock formed by the solidification of magma.

Kriging: A weighted, moving average interpolation method in which the set of weights assigned to samples minimizes the estimation variance.

Level: A main underground roadway or passage driven along a level course to afford access to stopes or workings and to provide ventilation and a haulage way for the removal of broken rock.

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Lithological: Geological description pertaining to different rock types.

Milling: A general term used to describe the process in which the ore is crushed, ground and subjected to physical or chemical treatment to extract the valuable minerals in a concentrate or finished product.

Mineral/Mining Lease: A lease area for which mineral rights are held.

Mining Assets: The Material Properties and Significant Exploration Properties.

Sedimentary: Pertaining to rocks formed by the accumulation of sediments, formed by the erosion of other rocks.

Sill1: A thin, tabular, horizontal to sub-horizontal body of igneous rock formed by the injection of magma into planar zones of weakness.

Sill2: The floor of a mine passage way.

Stope: An underground excavation from which ore has been removed.

Stratigraphy: The study of stratified rocks in terms of time and space.

Strike: Direction of line formed by the intersection of strata surfaces with the horizontal plane, always perpendicular to the dip direction.

Sulfide: A sulfur bearing mineral.

Tailings: Finely ground waste rock from which valuable minerals or metals have been extracted.

Thickening: The process of concentrating solid particles in suspension.

Total Expenditure: All expenditures including those of an operating and capital nature.

Variogram: A plot of the variance of paired sample measurements as a function of distance and/or direction.

Mineral Resources

Mineral Resources are sub-divided, in order of increasing geological confidence, into Inferred, Indicated and Measured categories. An Inferred Mineral Resource has a lower level of confidence than that applied to an Indicated Mineral Resource. An Indicated Mineral Resource has a higher level of confidence than an Inferred Mineral Resource but has a lower level of confidence than a Measured Mineral Resource.

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A Mineral Resource is a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction.

The location, quantity, grade or quality, continuity and other geological characteristics of a Mineral Resource are known, estimated or interpreted from specific geological evidence and knowledge, including sampling.

Material of economic interest refers to diamonds, natural solid inorganic material, or natural solid fossilized organic material including base and precious metals, coal, and industrial minerals.

The term Mineral Resource covers mineralization and natural material of intrinsic economic interest which has been identified and estimated through exploration and sampling and within which Mineral Reserves may subsequently be defined by the consideration and application of Modifying Factors. The phrase ‘reasonable prospects for eventual economic extraction’ implies a judgment by the Qualified Person in respect of the technical and economic factors likely to influence the prospect of economic extraction. The Qualified Person should consider and clearly state the basis for determining that the material has reasonable prospects for eventual economic extraction. Assumptions should include estimates of cut-off grade and geological continuity at the selected cut-off, metallurgical recovery, smelter payments, commodity price or product value, mining and processing method and mining, processing and general and administrative costs. The Qualified Person should state if the assessment is based on any direct evidence and testing.

Interpretation of the word ‘eventual’ in this context may vary depending on the commodity or mineral involved. For example, for some coal, iron, potash deposits and other bulk minerals or commodities, it may be reasonable to envisage ‘eventual economic extraction’ as covering time periods in excess of 50 years. However, for many gold deposits, application of the concept would normally be restricted to perhaps 10 to 15 years, and frequently to much shorter periods of time.

Inferred Mineral Resource

An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade or quality continuity.

An Inferred Mineral Resource has a lower level of confidence than that applying to an Indicated Mineral Resource and must not be converted to a Mineral Reserve. It is reasonably expected that the majority of Inferred Mineral Resources could be upgraded to Indicated Mineral Resources with continued exploration.

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An Inferred Mineral Resource is based on limited information and sampling gathered through appropriate sampling techniques from locations such as outcrops, trenches, pits, workings and drill holes. Inferred Mineral Resources must not be included in the economic analysis, production schedules, or estimated mine life in publicly disclosed Pre-Feasibility or Feasibility Studies, or in the Life of Mine plans and cash flow models of developed mines. Inferred Mineral Resources can only be used in economic studies as provided under NI 43-101.

There may be circumstances, where appropriate sampling, testing, and other measurements are sufficient to demonstrate data integrity, geological and grade/quality continuity of a Measured or Indicated Mineral Resource, however, quality assurance and quality control, or other information may not meet all industry norms for the disclosure of an Indicated or Measured Mineral Resource. Under these circumstances, it may be reasonable for the Qualified Person to report an Inferred Mineral Resource if the Qualified Person has taken steps to verify the information meets the requirements of an Inferred Mineral Resource

Indicated Mineral Resource

An Indicated Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics are estimated with sufficient confidence to allow the application of Modifying Factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit.

Geological evidence is derived from adequately detailed and reliable exploration, sampling and testing and is sufficient to assume geological and grade or quality continuity between points of observation.

An Indicated Mineral Resource has a lower level of confidence than that applying to a Measured Mineral Resource and may only be converted to a Probable Mineral Reserve.

Mineralization may be classified as an Indicated Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such as to allow confident interpretation of the geological framework and to reasonably assume the continuity of mineralization. The Qualified Person must recognize the importance of the Indicated Mineral Resource category to the advancement of the feasibility of the project. An Indicated Mineral Resource estimate is of sufficient quality to support a Pre-Feasibility Study which can serve as the basis for major development decisions.

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Measured Mineral Resource

A Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of Modifying Factors to support detailed mine planning and final evaluation of the economic viability of the deposit.

Geological evidence is derived from detailed and reliable exploration, sampling and testing and is sufficient to confirm geological and grade or quality continuity between points of observation.

A Measured Mineral Resource has a higher level of confidence than that applying to either an Indicated Mineral Resource or an Inferred Mineral Resource. It may be converted to a Proven Mineral Reserve or to a Probable Mineral Reserve.

Mineralization or other natural material of economic interest may be classified as a Measured Mineral Resource by the Qualified Person when the nature, quality, quantity and distribution of data are such that the tonnage and grade or quality of the mineralization can be estimated to within close limits and that variation from the estimate would not significantly affect potential economic viability of the deposit. This category requires a high level of confidence in, and understanding of, the geology and controls of the mineral deposit.

‘Modifying Factors’ are considerations used to convert Mineral Resources to Mineral Reserves. These include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.

Mineral Reserve

Mineral Reserves are sub-divided in order of increasing confidence into Probable Mineral Reserves and Proven Mineral Reserves. A Probable Mineral Reserve has a lower level of confidence than a Proven Mineral Reserve.

A Mineral Reserve is the economically mineable part of a Measured and/or Indicated Mineral Resource. It includes diluting materials and allowances for losses, which may occur when the material is mined or extracted and is defined by studies at Pre-Feasibility or Feasibility level as appropriate that include application of Modifying Factors. Such studies demonstrate that, at the time of reporting, extraction could reasonably be justified.

The reference point at which Mineral Reserves are defined, usually the point where the ore is delivered to the processing plant, must be stated. It is important that, in all situations where the reference point is different, such as for a saleable product, a clarifying statement is included to ensure that the reader is fully informed as to what is being reported.

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The public disclosure of a Mineral Reserve must be demonstrated by a Pre-Feasibility Study or Feasibility Study.

Mineral Reserves are those parts of Mineral Resources which, after the application of all mining factors, result in an estimated tonnage and grade which, in the opinion of the Qualified Person(s) making the estimates, is the basis of an economically viable project after taking account of all relevant Modifying Factors. Mineral Reserves are inclusive of diluting material that will be mined in conjunction with the Mineral Reserves and delivered to the treatment plant or equivalent facility. The term ‘Mineral Reserve’ need not necessarily signify that extraction facilities are in place or operative or that all governmental approvals have been received. It does signify that there are reasonable expectations of such approvals.

‘Reference point’ refers to the mining or process point at which the Qualified Person prepares a Mineral Reserve. For example, most metal deposits disclose mineral reserves with a “mill feed” reference point. In these cases, reserves are reported as mined ore delivered to the plant and do not include reductions attributed to anticipated plant losses. In contrast, coal reserves have traditionally been reported as tonnes of “clean coal”. In this coal example, reserves are reported as a “saleable product” reference point and include reductions for plant yield (recovery). The Qualified Person must clearly state the ‘reference point’ used in the Mineral Reserve estimate.

Probable Mineral Reserve

A Probable Mineral Reserve is the economically mineable part of an Indicated, and in some circumstances, a Measured Mineral Resource. The confidence in the Modifying Factors applying to a Probable Mineral Reserve is lower than that applying to a Proven Mineral Reserve.

The Qualified Person(s) may elect, to convert Measured Mineral Resources to Probable Mineral Reserves if the confidence in the Modifying Factors is lower than that applied to a Proven Mineral Reserve. Probable Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study.

Proven Mineral Reserve (Proved Mineral Reserve)

A Proven Mineral Reserve is the economically mineable part of a Measured Mineral Resource. A Proven Mineral Reserve implies a high degree of confidence in the Modifying Factors.

Application of the Proven Mineral Reserve category implies that the Qualified Person has the highest degree of confidence in the estimate with the consequent expectation in the minds of the readers of the report. The term should be restricted to that part of the deposit where production planning is taking place and for which any variation in the estimate would not significantly affect the potential economic viability of the deposit. Proven Mineral Reserve estimates must be demonstrated to be economic, at the time of reporting, by at least a Pre-Feasibility Study. Within the CIM Definition standards the term Proved Mineral Reserve is an equivalent term to a Proven Mineral Reserve.

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Pre-Feasibility Study (Preliminary Feasibility Study)

The CIM Definition Standards requires the completion of a Pre-Feasibility Study as the minimum prerequisite for the conversion of Mineral Resources to Mineral Reserves.

A Pre-Feasibility Study is a comprehensive study of a range of options for the technical and economic viability of a mineral project that has advanced to a stage where a preferred mining method, in the case of underground mining, or the pit configuration, in the case of an open pit, is established and an effective method of mineral processing is determined. It includes a financial analysis based on reasonable assumptions on the Modifying Factors and the evaluation of any other relevant factors which are sufficient for a Qualified Person, acting reasonably, to determine if all or part of the Mineral Resource may be converted to a Mineral Reserve at the time of reporting. A Pre-Feasibility Study is at a lower confidence level than a Feasibility Study.

Feasibility Study

A Feasibility Study is a comprehensive technical and economic study of the selected development option for a mineral project that includes appropriately detailed assessments of applicable Modifying Factors together with any other relevant operational factors and detailed financial analysis that are necessary to demonstrate, at the time of reporting, that extraction is reasonably justified (economically mineable). The results of the study may reasonably serve as the basis for a final decision by a proponent or financial institution to proceed with, or finance, the development of the project. The confidence level of the study will be higher than that of a Pre-Feasibility Study.

The term proponent captures issuers who may finance a project without using traditional financial institutions. In these cases, the technical and economic confidence of the Feasibility Study is equivalent to that required by a financial institution.

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29.

Appendix A True North Claims Information

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Practical Mining LLC

May 12, 2017

Klondex Mines	Technical Report for the True North Mine, Bissett,	Page 181
Ltd	Manitoba, Canada

Practical Mining LLC

May 12, 2017

Page 182	Appendix A True North Claims Information	Klondex Mines
		Ltd.

Practical Mining LLC

May 12, 2017

Klondex Mines	Technical Report for the True North Mine, Bissett,	Page 183
Ltd	Manitoba, Canada

Practical Mining LLC

May 12, 2017

Page 184	Appendix B: Certification of Authors and Consent	Klondex Mines
	Forms	Ltd.

30.

Appendix B: Certification of Authors and Consent Forms

Practical Mining LLC

May 12, 2017

CERTIFICATE OF AUTHOR

Re: Technical Report for the True North Mine, Bissett, Manitoba, Canada, dated the 12th day of May 2017, with an effective date of March 31, 2017 (the "Technical Report").

I, Sarah M Bull, P.E., do hereby certify that:

As of May 12, 2017, I am a consulting mining engineer at:

Practical Mining LLC
495 Idaho Street, Suite 205
Elko, Nevada 89801
775-345-3718

	1)	I am a Registered Professional Mining Engineer in the State of Nevada (# 22797).

	2)	I am a graduate of The University of Alaska Fairbanks, Fairbanks, Alaska with a Bachelor of Science Degree in Mining Engineering in 2006.

	3)	Since my graduation from university I have been employed as a Mine Engineer at an underground gold mining operation and as Senior Mine Engineer for a consulting engineering firm. My responsibilities have included mine ventilation engineering, stope design and mine planning.

	4)	I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI 43- 101") and certify that by reason of my experience and qualifications and good standing with proper designation within a recognized professional organization I fully meet the criteria as a Qualified Person as defined under NI 43-101.

	5)	I am a contract consulting engineer for the issuer and project owner: Klondex Mines Ltd.

	6)	I am responsible for preparation of section 15, 16, 21 and 22, along with those parts of the, and sections 1, 25 and 26 pertaining thereto of this Technical Report.

	7)	I last visited the True North Project on March 8 and 9, 2017.

	8)	I am independent of Klondex Mines Ltd. within the meaning of Section 1.5 of NI 43-101.

	9)	I was paid a daily rate for engineering consulting services performed in evaluation of The True North Project for Klondex Mines Ltd. and do not have any other interests relating to the True North Project. I do not have any interest in adjoining properties in True North Project area.

	10)	I have read NI 43-101 and Form 43-101F1, and the sections of the Technical Report for which I am responsible have been prepared in accordance with that instrument and form.

	11)	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

	12)	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 12th day of May 2017.

“Signed” Sarah Bull

Sarah M Bull, P.E.

Practical Mining LLC

495 Idaho Street, Suite 205

Elko, Nevada 89801

775-304-5836

sarahbull@practicalmining.com

CERTIFICATE of QUALIFIED PERSON

Re: Technical Report for the True North Mine, Bissett, Manitoba, Canada, dated the 12th day of May 2017, with an effective date of March 31, 2017 (the "Technical Report"):

I, Mark A. Odell, P.E., do hereby certify that:

As of May 12, 2017, I am a consulting mining engineer at:
Practical Mining LLC
495 Idaho Street, Suite 205
Elko, Nevada 89801
775-345-3718

	1)	I am a Registered Professional Mining Engineer in the State of Nevada (# 13708), and a Registered Member (#2402150) of the Society for Mining, Metallurgy and Exploration (SME).

	2)	I graduated from The Colorado School of Mines, Golden, Colorado with a Bachelor of Science Degree in Mining Engineering in 1985. I have practiced my profession continuously since 1985.

	3)	Since 1985, I have held the positions of mine engineer, chief engineer, mine superintendent, technical services manager and mine manager at underground and surface metal and coal mines in the western United States. The past 9 years, I have worked as a self-employed mining consultant with clients located in North America, Asia and Africa. My responsibilities have included the preparation of detailed mine plans, geotechnical engineering, reserve and resource estimation, preparation of capital and operating budgets and the economic evaluation of mineral deposits.

	4)	I have read the definition of "qualified person" set out in National Instrument 43-101 ("NI 43- 101") and certify that by reason of my experience and qualifications and good standing with proper designation within a recognized professional organization fully meet the criteria as a Qualified Person as defined under NI 43-101.

	5)	I am a contract consulting engineer for the issuer and project owner, Klondex Mines Ltd. (the "Issuer"). I have not visited the True North Mine.

	6)	I am responsible for sections 2, 3, 15-16, 19, 21-22 and 24 along with those parts of sections 1, 25 and 26 pertaining thereto of this Technical Report.

	7)	I am independent of the Issuer within the meaning of Section 1.5 of NI 43-101.

	8)	I was paid a daily rate for consulting services performed in evaluation of the True North Mine for the Issuer and do not have any other interests relating to the True North Mine. I do not have any interest in adjoining properties in the True North area.

	9)	I have read NI 43-101 and Form 43-101F1, and the sections of the Technical Report for which I am responsible have been prepared in accordance with that instrument and form.

	10)	I consent to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, of the Technical Report.

	11)	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all the scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 12th day of May 2017.

“Signed” Mark A. Odell

Mark A. Odell, P.E.

Practical Mining LLC

markodell@practicalmining.com

CERTIFICATE OF QUALIFIED PERSON

FRED H. BROWN, P.GEO.

I, Fred H. Brown, of Suite B-10, 1610 Grover St., Lynden WA, 98264 USA, do hereby certify that:

1. I am an independent geological consultant and have worked as a geologist continuously since my graduation from university in 1987.

2. This certificate applies to the technical report titled “Technical Report for the True North Mine, Bissett, Manitoba, Canada” (the “Technical Report”), with an effective date of March 31, 2017.

3. I graduated with a Bachelor of Science degree in Geology from New Mexico State University in 1987. I obtained a Graduate Diploma in Engineering (Mining) in 1997 from the University of the Witwatersrand and a Master of Science in Engineering (Civil) from the University of the Witwatersrand in 2005. I am registered with the Association of Professional Engineers and Geoscientists of British Columbia as a Professional Geoscientist (171602) and the Society for Mining, Metallurgy and Exploration as a Registered Member (#4152172)..

4. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101

This report is based on my personal review of information provided by Klondex Canada Ltd. and on discussions with its representatives. My relevant experience for the purpose of the Technical Report is:

•	Underground Mine Geologist, Freegold Mine, AAC	1987-1995
•	Mineral Resource Manager, Vaal Reefs Mine, Anglogold	1995-1997
•	Resident Geologist, Venetia Mine, De Beers	1997-2000
•	Chief Geologist, De Beers Consolidated Mines	2000-2004
•	Consulting Geologist	2004-Present

5. I visited the True North site over the period February 20, 2017 to February 24, 2017.

6. I am responsible for authoring Section 14 of this Technical Report.

7. I am independent of Klondex Mines Ltd. applying the test in Section 1.5 of NI 43-101.

8. I was previously involved with the project as a coauthor of the technical report “Amended and Restated Technical Report and Pre-Feasibility Study on the True North Gold Mine, Bissett, Manitoba, Canada”, with an effective date of June 30, 2016.

9. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith.

10. As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading..

Effective Date: March 31, 2017
Signed Date: May 12, 2017

{SIGNED AND SEALED}

[Fred H. Brown]

Fred H. Brown, P. Geo.

CERTIFICATE OF QUALIFIED PERSON

WILLIAM STONE, Ph.D., P. Geo.

I, William Stone, Ph.D., P. Geo, residing at 4361 Latimer Crescent, Burlington, Ontario, do hereby certify that:

	1.	I am an independent geological consultant
	2.	This certificate applies to the technical report titled “Technical Report for the True North Mine, Bissett, Manitoba, Canada” (the “Technical Report”), with an effective date of March 31, 2017.
	3.	I am a graduate of Dalhousie University with a Bachelor of Science (Honours) degree in Geology (1983). In addition, I have a Master of Science in Geology (1985) and a Ph.D. in Geology (1988) from the University of Western Ontario. I have worked as a geologist for a total of 31 years since obtaining my M.Sc. degree. I am a geological consultant currently licensed by the Association of professional Geoscientists of Ontario (License No 1569).
	4.	I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify, by reason of my education, affiliation with a professional organization (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. My relevant experience for the purpose of the Technical Report is:

•	Contract Senior Geologist, LAC Minerals Exploration Ltd	1985-1988
•	Post-Doctoral Fellow, McMaster University	1988-1992
•	Contract Senior Geologist, Outokumpu Mines and Metals Ltd	1993-1996
•	Senior Research Geologist, WMC Resources Ltd	1996-2001
•	Senior Lecturer, University of Western Australia	2001-2003
•	Principal Geologist, Geoinformatics Exploration Ltd	2003-2004
•	Vice President Exploration, Nevada Star Resources Inc	2005-2006
•	Vice President Exploration, Goldbrook Ventures Inc	2006-2008
•	Vice President Exploration, North American Palladium Ltd	2008-2009
•	Vice President Exploration, Magma Metals Ltd.	2010-2011
•	President & COO, Pacific North West Capital Corp	2011-2014
•	Consulting Geologist	2013-Present

	5.	I visited the Property that is the subject of this report from May 21-25, June 20-23, September 20- 22, 2016 and January 16-19, 2017.
	6.	I am responsible for authoring Sections 4-12 and 23 of this Technical Report along with those parts of the Executive Summary, and Sections 25 and 26 pertaining thereto.
	7.	I am independent of Klondex Mines Ltd. applying the test in Section 1.5 of NI 43-101.
	8.	I was previously involved with the project as a coauthor of the technical report “Amended and Restated Technical Report and Pre-Feasibility Study on the True North Gold Mine, Bissett, Manitoba, Canada”, with an effective date of March 31, 2017.
	9.	I have read NI 43-101 and Form 43-101F1 and this Technical Report has been prepared in compliance therewith.
	10.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: March 31, 2017
Signed Date: May 12, 2017

{SIGNED AND SEALED}

[William Stone]

Dr. William E. Stone, P. Geo.

CERTIFICATE OF QUALIFIED PERSON

ALFRED S. HAYDEN, P. ENG

I, Alfred S. Hayden, P. Eng., residing at 284 Rushbrook Drive, Ontario, L3X 2C9, do hereby certify that:

1.	I am currently President of:

		EHA Engineering Ltd.,
		Consulting Metallurgical Engineers
		Box 2711, Postal Stn. B.
		Richmond Hill, Ontario, L4E 1A7

2.	This certificate applies to the technical report titled “Technical Report for the True North Mine, Bissett, Manitoba, Canada” (the “Technical Report”), with an effective date of March 31, 2017.
3.	I graduated from the University of British Columbia, Vancouver, B.C. in 1967 with a Bachelor of Applied Science in Metallurgical Engineering. I am a member of the Canadian Institute of Mining, Metallurgy and Petroleum and a Professional Engineer and Designated Consulting Engineer registered with Professional Engineers Ontario. I have worked as a metallurgical engineer for over 40 years since my graduation from university.
4.	I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

My summarized career experience is as follows:

•	EHA Engineering Ltd: (President)	1990-Present
•	EH Associates: (Partner)	1985-1990
•	A.H. Ross & Associates Ltd. (Senior Associate)	1976-1985
•	Eldorado Nuclear Limited (Chief Metallurgist/Mill Engineer)	1966-1976

5.	I visited the Property that is the subject of this report on September 20 and 21, 2016.
6.	I am responsible for authoring Sections 13 and 17 of this Technical Report along with those parts of the Executive Summary, and Sections 25 and 26 pertaining thereto.
7.	I am independent of Klondex Mines Ltd. applying the test in Section 1.5 of NI 43-101.
8.	I was previously involved with the project as a coauthor of the technical report “Amended and Restated Technical Report and Pre-Feasibility Study on the True North Gold Mine, Bissett, Manitoba, Canada”, with an effective date of March 31, 2017.
9.	I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith.
10.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: March 31, 2017
Signed Date: May 12, 2017

{SIGNED AND SEALED}

[Alfred Hayden]

Alfred S. Hayden, P.Eng.

CERTIFICATE OF QUALIFIED PERSON

DAVID A. ORAVA, P. ENG.

I, David A. Orava, M. Eng., P. Eng., residing at 19 Boulding Drive, Aurora, Ontario, L4G 2V9, do hereby certify that:

1.	I am an Associate Mining Engineer at P&E Mining Consultants Inc. and President of Orava Mine Projects Ltd.
2.	This certificate applies to the technical report titled “Technical Report for the True North Mine, Bissett, Manitoba, Canada” (the “Technical Report”), with an effective date of March 31, 2017.
3.	I am a graduate of McGill University located in Montreal, Quebec, Canada at which I earned my Bachelor Degree in Mining Engineering (B.Eng. 1979) and Masters in Engineering (Mining - Mineral Economics Option B) in 1981. I have practiced my profession continuously since graduation. I am licensed by the Professional Engineers of Ontario (License No. 34834119).
4.	I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

My summarized career experience is as follows:

•	Mining Engineer – Iron Ore Company of Canada.	1979-1980
•	Mining Engineer – J.S Redpath Limited / J.S. Redpath Engineering.	1981-1986
•	Mining Engineer & Manager Contract Development – Dynatec Mining Ltd.	1986-1990
•	Vice President – Eagle Mine Contractors	1990
•	Senior Mining Engineer – UMA Engineering Ltd.	1991
•	General Manager - Dennis Netherton Engineering	1992-1993
•	Senior Mining Engineer – SENES Consultants Ltd.	1993-2003
•	President – Orava Mine Projects Ltd.	2003 to present
•	Associate Mining Engineer – P&E Mining Consultants Inc.	2006 to present

5.	I have not visited the Property that is the subject of this report.
6.	I am responsible for authoring Section 20 of the Technical Report along with those parts of the Executive Summary, and Sections 25 and 26 pertaining thereto.
7.	I am independent of Klondex Mines Ltd. applying all of the tests in Section 1.5 of NI 43-101.
8.	I was previously involved with the project as a coauthor of the technical report “Amended and Restated Technical Report and Pre-Feasibility Study on the True North Gold Mine, Bissett, Manitoba, Canada”, with an effective date of March 31, 2017.
9.	I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance therewith.
10.	As at the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Effective Date: March 31, 2017
Signed Date: May 12, 2017

{SIGNED AND SEALED}

[David Orava]

David Orava, M. Eng., P. Eng.

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