Exhibit 96.1

  TECHNICAL REPORT
SUMMARY
ELK CREEK PROJECT
NEBRASKA EFFECTIVE DATE: June 30, 2022 SIGNATURE DATE: September 2, 2022 PREPARED BY: Dahrouge Geological Consulting USA Ltd. Understood Mineral Resources Ltd. Optimize Group Inc. Tetra Tech Adrian Brown Consultants Inc. Magemi Mining Inc. L3 Process Development Olsson A2GC Metallurgy Concept Solutions Scott Honan, M.Sc., SME-RM, NioCorp Everett Bird, P.E., Cementation Matt Hales, P.E., Cementation Mahmood Khwaja, P.E., CDM Smith Martin Lepage, P.Eng., Ing., Cementation Wynand Marx, M.Eng., BBE Consulting

Elk Creek Project S-K 1300

Technical Report Summary i

Contents

1.	EXECUTIVE SUMMARY			1
	1.1	Principal Outcomes		1
	1.2	Property Description and Ownership		2
	1.3	Geological Setting and Mineralization		2
	1.4	History		3
	1.5	Drilling		3
	1.6	Mineral Resource Estimation		4
	1.7	Mineral Reserve Estimation		9
	1.8	Environmental Studies, Permitting and Social or Community Impact		10
	1.9	Capital Cost Estimate		12
	1.10	Operating Cost Estimate		14
	1.11	QP Conclusions and Recommendations		14
2.	INTRODUCTION			15
	2.1	Terms of Reference and Purpose of the Technical Report Summary		15
	2.2	Sources of Information		15
	2.3	Details of Inspection		15
	2.4	History		17
3.	PROPERTY DESCRIPTION			18
	3.1	Property Location		18
	3.2	Property Description and Land Tenure		19
		3.2.1	Nature and Extent of Issuer’s Interest	21
	3.3	Royalties, Agreements and Encumbrances		21
		3.3.1	Required Permits and Status	21
	3.4	Other Significant Factors and Risks		22
4.	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			23
	4.1	Physiography		23
	4.2	Accessibility and Transportation to the Property		23
	4.3	Climate and Length of Operating Season		24
	4.4	Infrastructure		25
		4.4.1	Personnel and Supplies	25
		4.4.2	Electrical Power	25
		4.4.2.1	Electrical Power Line & Substation	25
		4.4.2.2	Electrical Power Distribution - Plant and Facilities	25

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Technical Report Summary ii

		4.4.2.3	Electrical Power Distribution - Underground	25
		4.4.2.4	Emergency Power Generation	26
		4.4.3	Water	26
		4.4.3.1	Process Water	26
		4.4.3.2	Fire Water	27
		4.4.3.3	Potable Water	28
5.	HISTORY			29
	5.1	Exploration History		29
		5.1.1	USGS, 1964	29
		5.1.2	Discovery, 1970-1971	30
		5.1.3	Cominco American, 1974	32
		5.1.4	Molycorp, 1973-1986	32
		5.1.5	Geophysical Surveys	32
		5.1.6	Drilling	32
		5.1.7	Molycorp Data Verification, 1973-1986	33
6.	GEOLOGICAL SETTING, MINERALISATION AND DEPOSIT			35
	6.1	Regional Geology		35
	6.2	Property Geology		37
	6.3	Elk Creek Carbonatite		38
		6.3.1	Age Dating	42
	6.4	Carbonatite Lithological Unit		43
	6.5	Marine Sedimentary Rocks		44
	6.6	Structural Geology		46
	6.7	Mineralization		46
		6.7.1	Niobium and Titanium Mineralization	47
		6.7.2	Scandium Mineralization	48
		6.7.3	Rare Earth Element Mineralization	49
	6.8	Deposit Types		50
7.	EXPLORATION			53
	7.1	Geophysical Exploration		54
	7.2	Drilling		54
		7.2.1	Type and Extent	54
		7.2.2	Molycorp, 1973-1986	56
		7.2.3	Quantum, 2011	57
		7.2.4	NioCorp 2014 Program	58

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Technical Report Summary iii

		7.2.4.1	Procedures (NioCorp 2014 Program)	60
		7.2.4.2	Collar Surveys	61
		7.2.4.3	Downhole Surveys	62
		7.2.5	Interpretation and Relevant Results	63
	7.3	Geotechnical Design Parameters		64
	7.4	Hydrogeology Design Parameters		80
		7.4.1	Conceptual Geohydrology	81
		7.4.2	Mine Inflow Control	88
8.	SAMPLE PREPARATION, ANALYSES, AND SECURITY			94
	8.1	Sample Preparation and Security		94
		8.1.1	Molycorp, 1973 – 1986	94
		8.1.2	NioCorp Drilling Program, 2011 - Current	96
		8.1.3	Historical Re-Sampling Programs	100
		8.1.3.1	NioCorp (Quantum 2010) Historical Re-Sampling Program	100
		8.1.3.2	NioCorp (2014-2016) Historical Re-Sampling Program	101
		8.1.3.3	NioCorp (2021) Historical Re-Sampling Programs	102
	8.2	Sample Analysis Procedures, 2011 – Current		103
	8.3	Quality Assurance/Quality Control Programs		106
		8.3.1	Re-Sampling/Verification of Historical Assays	106
		8.3.2	NioCorp 2011 - Current	107
		8.3.3	Quality Assurance & Quality Control Results	110
		8.3.3.1	Field Quartz Blanks	110
		8.3.3.2	Certified & Standard Reference Material	112
		8.3.3.3	Reject Duplicates	129
		8.3.3.4	Field 1/4 Core Duplicates	132
		8.3.3.5	Third-Party Duplicate Check Analysis	135
	8.4	Qualified Person’s Opinion on the Adequacy of Sample Preparation, Security and Analytical Procedures		140
	8.5	Specific Gravity		140
9.	DATA VERIFICATION			142
	9.1	Understood and Optimize Group Data Validation		142
		9.1.1	Site Visit	142
		9.1.1.1	Drill Core Review	142
		9.1.1.2	Collar Verification	143
		9.1.1.3	Core Processing Protocols	144

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		9.1.2	Database Validation	144
		9.1.3	Review of NioCorp QA/QC	144
	9.2	Limitations		144
	9.3	Qualified Person’s Opinion		145
10.	MINERAL PROCESSING AND METALLURGICAL TESTING			146
	10.1	Mineral Processing		146
	10.2	Hydrometallurgy		147
		10.2.1	Testing and Procedures	147
		10.2.2	Relevant Results	160
		10.2.3	Significant Factors	161
	10.3	Pyrometallurgy		161
11.	MINERAL RESOURCE ESTIMATES			163
	11.1	Introduction		163
	11.2	Source Database		164
		11.2.1	Drill Holes	167
	11.3	Geological Domaining		168
	11.4	Exploratory Data Analysis		172
		11.4.1	Compositing	172
		11.4.2	Declustering	173
		11.4.3	Outlier Capping	173
		11.4.4	Representative Distributions Statement	175
	11.5	Exploratory Data Analysis		176
	11.6	Variography		180
	11.7	Block Model Resource Estimation		182
		11.7.1	Estimation Overview	182
		11.7.2	Block Model Definition	182
		11.7.3	Estimation Strategy and Testing	183
		11.7.3.1	Estimation Strategy	183
		11.7.3.2	Testing and Strategy Refinement	184
		11.7.4	Estimation/Interpolation Methods	187
	11.8	Model Validation		188
		11.8.1	Rare Earth Considerations	188
		11.8.2	Global Checks	189
		11.8.3	Visual Inspection	192
		11.8.4	Swath Plots	194

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		11.8.5	Correlation Review	195
	11.9	Mineral Resource Classification		196
	11.10	Reasonable Prospects of Eventual Economic Extraction		197
	11.11	Cut-Off Grade		197
	11.12	Mineral Resource Tabulation		198
	11.13	Mineral Resource Uncertainty		202
		11.13.1	Specific Identified Risks	202
		11.13.2	Generic Mineral Resource Uncertainty	203
	11.14	Mineral Resource Sensitivity		203
	11.15	Relevant Factors		205
12.	MINERAL RESERVE ESTIMATES			206
	12.1	Conversion Assumptions, Parameters and Methods		206
		12.1.1	Dilution	206
		12.1.2	Recovery	208
		12.1.3	Cut-Off Grade Calculation	208
		12.1.4	Mine Design	211
	12.2	Reserves		213
	12.3	QP Opinion and Relevant Factors		214
13.	MINING METHODS			215
	13.1	Geotechnical Design Parameters		215
	13.2	Hydrogeology Design Parameters		215
	13.3	Mine Design		215
		13.3.1	Selection of Mining Method	215
		13.3.2	Stope Optimization	216
		13.3.3	Stope Design	217
		13.3.4	Development Design	219
		13.3.5	Mine Access	224
		13.3.5.1	Shaft Layouts	225
	13.4	Production Schedule		228
		13.4.1	Productivity	228
		13.4.2	Shaft Sinking – Production Shaft and Ventilation Shaft	231
		13.4.3	Development and Production Schedule	231
	13.5	Mining Operations		234
		13.5.1	Production	234
		13.5.2	Development	235

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		13.5.3	Truck and LHD Haulage	235
		13.5.4	Backfilling	239
		13.5.5	Ground Support	241
		13.5.6	Grade Control and Reconciliation	241
		13.5.7	Workforce	245
		13.5.8	Equipment	247
	13.6	Ventilation		249
		13.6.1	Airflow Requirements	249
		13.6.2	Ventilation Controls	250
		13.6.3	Ventilation Model	252
		13.6.4	Auxiliary Ventilation	253
		13.6.5	Recommended Ventilation Infrastructure	253
		13.6.6	Ventilation Power Consumption	254
		13.6.7	Air Heating	254
		13.6.8	Thermal Exposure	255
	13.7	Mine Infrastructure & Services		255
		13.7.1	Material Handling System	255
		13.7.2	Mine Dewatering System	255
		13.7.3	Compressed Air System	257
		13.7.4	Underground Water Supply	257
		13.7.5	Underground Fuel Storage and Distribution	257
		13.7.6	Workshop, Maintenance Bays, and Warehouse	258
		13.7.7	Explosives Storage	258
		13.7.8	Refuge Stations/Chambers	258
		13.7.9	Hoist House Substation Surface Electrical Distribution	258
		13.7.10	Underground Electrical Distribution	259
		13.7.11	Overhead Pole Line Electrical Distribution	259
		13.7.12	Hoisting Plants	259
		13.7.13	Dust Suppression System	262
		13.7.14	Communications System	263
		13.7.15	Safety and Health	263
14.	PROCESSING AND RECOVERY METHODS			264
	14.1	Process Plant Design Criteria		264
		14.1.1	Surface Crushing, Ore Storage & Mineral Processing Plant	264
		14.1.2	Hydrometallurgical Plant	264

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		14.1.3	Pyrometallurgical Plant	264
		14.1.4	Acid Plant	271
	14.2	Flowsheets and Process Description		272
		14.2.1	Surface Crushing, Ore Storage & Mineral Processing Plant	272
		14.2.2	Hydrometallurgical Plant	273
		14.2.3	Pyrometallurgical Plant	280
		14.2.4	Acid Plant	282
	14.3	Process Equipment		283
		14.3.1	Surface Crushing, Ore Storage & Mineral Processing Plant	283
		14.3.2	Hydrometallurgical Plant	284
		14.3.3	Pyrometallurgical Plant	290
		14.3.4	Acid Plant	291
	14.4	Power Requirements		292
		14.4.1	Surface Crushing, Ore Storage & Mineral Processing Plant	292
		14.4.2	Hydrometallurgical Plant	292
		14.4.3	Pyrometallurgical Plant	292
		14.4.4	Acid Plant	293
	14.5	Plant Water		293
		14.5.1	Water Treatment Plant	293
		14.5.1.1	Flow Equalization	294
		14.5.1.2	Softening and Clarification	294
		14.5.1.3	Multimedia Filtration	295
		14.5.1.4	Reverse Osmosis (RO) System	295
		14.5.1.5	Cooling Tower Makeup System (CTMU)	296
		14.5.1.6	Sludge Handling	297
		14.5.1.7	Evaporation and Crystallization System	297
		14.5.1.8	Crystallizer Brine Flow	297
		14.5.2	Process Water	300
		14.5.3	Fire Water	301
		14.5.4	Potable Water	301
15.	INFRASTRUCTURE			302
	15.1	Electrical Power		302
		15.1.1	Electrical Power Line & Substation	302
		15.1.2	Electrical Power Distribution - Plant and Facilities	303
		15.1.3	Electrical Power Distribution - Underground	303

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		15.1.4	Emergency Power Generation	303
	15.2	Natural Gas		303
		15.2.1	Natural Gas Pipeline to Site	303
		15.2.2	Natural Gas Distribution on Site	303
	15.3	Plant Water		304
		15.3.1	Water Treatment Plant	304
		15.3.1.1	Flow Equalization	305
		15.3.1.2	Softening and Clarification	305
		15.3.1.3	Multimedia Filtration	306
		15.3.1.4	Reverse Osmosis (RO) System	306
		15.3.1.5	Cooling Tower Makeup System (CTMU)	307
		15.3.1.6	Sludge Handling	307
		15.3.1.7	Evaporation and Crystallization System	307
		15.3.1.8	Crystallizer Brine Flow	308
		15.3.2	Process Water	311
		15.3.3	Fire Water	312
		15.3.4	Potable Water	312
	15.4	Roads		312
		15.4.1	Main Access Road to Site	312
		15.4.2	Secondary Site Access Roads	312
		15.4.3	Secondary Site Roads (to tailings, etc.)	313
	15.5	Tailing Storage and Associated Facilities		314
	15.6	Salt Management Cells		326
	15.7	Paste Backfill Plant and Underground Distribution		327
		15.7.1	Surface Plant	327
		15.7.2	Backfill Testwork	328
		15.7.3	Paste Plant Process	330
		15.7.4	Underground Distribution of Paste Backfill	332
	15.8	Freeze Plant		332
16.	MARKET STUDIES			335
	16.1	Market Studies		335
		16.1.1	Niobium Market Overview	335
		16.1.1.1	Niobium Supply	335
		16.1.1.2	Niobium Demand	336
		16.1.2	Titanium Dioxide Market Overview	339

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		16.1.2.1	Titanium Dioxide Demand	340
		16.1.3	Scandium Trioxide Market Overview	340
		16.1.4	Key Aspects of OnG Commodities Report	341
		16.1.5	Rare Earth Market Overview	349
	16.2	Contracts and Status		356
17.	ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS			357
	17.1	Environmental Studies		357
		17.1.1	Soils	357
		17.1.2	Climate/Meteorology/Air Quality	357
		17.1.3	Cultural and Archeological Resources	357
		17.1.4	Vegetation	358
		17.1.5	Wildlife	359
		17.1.6	Threatened, Endangered, and Special Status Species	359
		17.1.7	Land Use	359
		17.1.8	Hydrogeology (Groundwater)	360
		17.1.9	Hydrology (Surface Water)	361
		17.1.10	Wetlands/Riparian Zones	362
		17.1.11	Geochemistry	362
		17.1.12	Known Environmental Issues	364
		17.1.13	Tailings	364
		17.1.14	Project Waste Disposal	364
		17.1.15	Site Monitoring	364
		17.1.16	Water Management	365
		17.1.17	Chemical and Reagents Handling	366
	17.2	Project Permitting Requirements		366
		17.2.1	Nebraska Underground Injection Control (UIC)	369
		17.2.2	DHHS Radioactive Materials Program and Licensing	369
		17.2.3	Nebraska Air Quality Permitting	370
		17.2.4	Nebraska Dam Permitting	371
		17.2.5	Greenhouse Gas Permitting	371
		17.2.6	Permitting Status	372
		17.2.7	Post-Performance and Reclamation Bonding	372
	17.3	Community Relations and Social Responsibilities		373
		17.3.1	Safety and Health	373

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	17.4	Reclamation and Closure		374
		17.4.1	Surface Disturbance	374
		17.4.2	Buildings and Equipment	375
		17.4.3	Tailings Disposal Facility	375
		17.4.4	Closure Cost Estimate	375
	17.5	International Standards and Guidelines		376
	17.6	Qualified Person’s Opinion		376
18.	CAPITAL AND OPERATING COSTS			377
	18.1	Capital Cost Estimate		377
		18.1.1	Basis of Estimate	377
		18.1.1.1	Mining, Process, and Infrastructure Capital Costs	377
		18.1.1.2	Tailings and Tailings Water Management Capital Costs	377
	18.2	Capital Cost Summary		377
		18.2.1	Capitalized Pre-production Costs	378
		18.2.2	Mining Capital Costs	379
		18.2.3	Processing Plant Capital Costs	381
		18.2.3.1	Processing Indirects	382
		18.2.3.2	Process Commissioning	383
		18.2.4	Tailings Water Management and Salt Management Cells	383
		18.2.4.1	Temporary Waste Rock Storage Facility	385
		18.2.5	Water Management and Infrastructure	385
		18.2.6	Site Preparation and Infrastructure Capital Costs	386
		18.2.6.1	Site Wide Indirects	387
		18.2.7	Owner’s Costs	388
		18.2.8	Closure and Reclamation	390
		18.2.9	Sustaining Capital Costs	391
		18.2.10	Contingency	394
	18.3	Operating Cost Estimate		395
		18.3.1	Basis of Estimate	395
		18.3.1.1	Mining Operating Costs	395
		18.3.1.2	Process Plants Operating Costs	395
		18.3.1.3	Tailings and Tailings Water Management Operating Costs	395
		18.3.1.4	Site G&A Operating Costs	396
		18.3.1.5	Owner’s Costs Capital Costs	396
		18.3.1.6	Water Supply Operating Costs	396

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		18.3.1.7	Closure and Reclamation	396
		18.3.2	Operating Cost Summary	396
		18.3.2.1	Mining Operating Costs	397
		18.3.2.2	Process Plant Operating Costs	397
		18.3.2.3	Tailings, Salt and Tailings Water Management Operating Costs	399
		18.3.2.4	Site G&A Operating Costs	401
19.	ECONOMIC ANALYSIS			407
	19.1	Methodology Used		407
	19.2	Financial Model Parameters and Assumptions		407
		19.2.1	Physicals	408
		19.2.2	Revenue	410
		19.2.3	Operating Costs	412
		19.2.4	Capital Costs	412
	19.3	Cashflow Forecasts and Annual Production Forecasts		414
	19.4	Sensitivity Analysis		418
20.	ADJACENT PROPERTIES			425
21.	OTHER RELEVANT DATA AND INFORMATION			426
	21.1	Project Implementation Plan		426
		21.1.1	Project Cost Objectives	426
		21.1.2	Project Schedule Objectives	426
		21.1.3	Early Works	428
		21.1.4	Project Team	428
		21.1.5	Project and Document Control	429
		21.1.6	Engineering	429
		21.1.7	Supply Chain and Procurement	430
		21.1.8	Construction Management	430
		21.1.9	Commissioning, Operational Readiness, and Early Operations	431
22.	INTERPRETATIONS AND CONCLUSIONS			433
	22.1	Introduction		433
	22.2	Geology & Mineral Resource		433
	22.3	Exploration, Drilling and Analytical Data Collection in Support of Mineral Resource Estimation		433
	22.4	Processing and Metallurgical Testing		434
	22.5	Mining & Mineral Reserve		435
	22.6	Recovery Methods		437
	22.7	Infrastructure		437

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		22.7.1	Tailings	438
	22.8	Environmental, Permitting and Social or Community Considerations		438
	22.9	Market Studies and Contracts		439
	22.10	Capital and Operating Costs		440
	22.11	Economic Analysis		440
	22.12	Opportunities and Risk Assessment		440
		22.12.1	Opportunities	441
		22.12.2	Risks	442
23.	RECOMMENDATIONS			446
	23.1	Recommended Work Programs		446
		23.1.1	Geology and Resources	446
		23.1.1.1	Quality Assurance/Quality Control	448
		23.1.2	Hydrometallurgical Plant	448
		23.1.3	Geotechnical	450
		23.1.4	Mining and Reserves	450
		23.1.5	Recovery Methods	451
		23.1.6	Infrastructure	451
		23.1.7	Environmental and Social	452
		23.1.8	Hoisting Plants	453
		23.1.9	Summary of Costs for Recommended Work	454
24.	REFERENCES			456
	24.1	References		456
	24.2	Glossary		460
		24.2.1	Mineral Resource	461
		24.2.2	Mineral Reserve	461
		24.2.3	Definition of Terms	462
		24.2.4	Abbreviations	463
25.	RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT			469
26.	DATE AND SIGNATURE PAGE			473
27.	QP RESPONSIBILITY MATRIX			476

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List of Figures

Figure 3-1: Project Location Map	19
Figure 3-2: Land Tenure Map*	20
Figure 3-3: Net Smelter Return (NSR) Map	21
Figure 4-1: Project Location Showing Main Access Routes	23
Figure 5-1: 1964 USGS Aeromagnetic Survey Area Showing Surveys 526A, 526B and 530 Respectively	29
Figure 5-2: 1964 USGS Aeromagnetic Results (Merged 526A, 526B, and 530 Surveys)	30
Figure 5-3: Comparison of the 1970 Magnetic and Gravity Geophysical Surveys	31
Figure 5-4: Cross-section A-A’ of the 1970 Gravity and Magnetic Geophysical Surveys	31
Figure 6-1: Regional Geology Map	35
Figure 6-2: Merged Aeromagnetic Anomaly Map of Nebraska, Kansas and Oklahoma States	36
Figure 6-3: Generalized Stratigraphy of Elk Creek Area (see Table 6-2 for details)	38
Figure 6-4: (alt): Cross-Section A-A’ (NW to SE) and B-B’ (SW to NE) from Figure 7-1	39
Figure 6-5: Core Photographs Showing Microstructures	40
Figure 6-6: Schematic of Drill Hole Showing the Typical Transition from Pennsylvanian Sediments to Carbonatite Units	41
Figure 6-7: Drill Hole NEC14-022, ~ 2m Interval of the Mudstone Contact Between the Pennsylvanian Sediments and the Carbonatite Units	42
Figure 6-8: Drill Hole NEC14-022, Relatively Massive Dolomitic Carbonatite ~ 3m Below the Contact with the Pennsylvanian Sediments in Figure 6-7	42
Figure 6-9: Plan View of the Location of the Mineralized Carbonatite (Mcarb)	47
Figure 6-10: Basic Statistics of Nb2O5 Mineralization	48
Figure 6-11: Correlation Statistics of Nb2O5 and TiO2 and Fe2O3	48
Figure 6-12: Basic Statistics of Sc Mineralization	49
Figure 6-13: Schematic diagram of St. Honoré Carbonatite	52
Figure 7-1: Geology of the Elk Creek Carbonatite as Expressed in Drill Holes at an Elevation of 120 m Above Sea Level (Roughly 230 m Below Ground Surface)	53
Figure 7-2: Elk Creek Drill Hole Location Map	55
Figure 7-3: Elk Creek Drill Hole Location Map by Operator	57
Figure 7-4: Drill Hole Traces Used in the 2022 Mineral Resource	60
Figure 7-5: Collar Location of NEC14-MET-01           Figure 7-6: Collar Location of NEC14-009	61
Figure 7-7: 3D View of Elk Creek Deposit Showing Modelled Base of Till and Unconformity between Pennsylvanian Sediments and the Elk Creek Carbonatite	63
Figure 7-8: Plan View Location of 2014-2015 Geotechnical Drill Holes	65
Figure 7-9: Vertical View of the Location of 2014-2015 Geotechnical Drill Holes Looking Towards the North	66
Figure 7-10: Vertical View of the Location of 2014-2015 Geotechnical Drill Holes Looking Towards the North with High Grade Niobium Wireframe (>1% Nb2O5)	67

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Technical Report Summary xiv

Figure 7-11: Geotechnical Model, Vertical Cross Section (N40°E Section)	69
Figure 7-12: Plan View of Geologic Structures (Green) on +60 m Elevation Level in Block 1	70
Figure 7-13: Orientation of Stress Measurements Relative to Faults and Fracture Orientations	73
Figure 7-14: Empirical Stope Design Chart for Moderately Weathered Rock Mass	75
Figure 7-15: Empirical Stope Design Chart for Fresh and Slightly Weathered Rock Mass	76
Figure 7-16: Empirical ELOS Estimate – Moderately Weathered and Fresh Rock Mass	78
Figure 7-17: Hydrogeologic Plan	82
Figure 7-18: Hydrogeologic Section	83
Figure 7-19: Permeability Testing of the Elk Creek Rock Mass	84
Figure 7-20: Response to Injection in Carbonatite - End of Test	86
Figure 8-1: Storage Location of Drill Core and Pulps	99
Figure 8-2: Sample Process Flow Chart (2014 Drill Program)	99
Figure 8-3: Resource Area Assay Distribution Showing REE Assays (Red) and REE Assay Gaps (Blue)	103
Figure 8-4: Summary of Blank Control Charts for Nb2O5 , Sc, TiO2 Submission to Actlabs 2011 and 2014	111
Figure 8-5: Summary of Blank Control Charts for REEs, La, Ce, Nd, and Pr, Dy Submission to Actlabs	112
Figure 8-6: Summary of SX18-01, SX18-02, SX18-04, SX18-05 Nb2O5 Control Chart	114
Figure 8-7: Summary of GRE-04 Nb2O5 Control Chart	115
Figure 8-8: Summary of Oreas 460 and 464 Nb2O5 Control Chart	115
Figure 8-9: Summary of GRE-03 and GRE-04 Sc Control Chart	116
Figure 8-10: Summary of OREAS 460 and OREAS 464 Sc Control Chart	116
Figure 8-11: Summary of SX18-01, SX18-02, SX18-04, SX18-05 TiO2 Control Chart	117
Figure 8-12: Summary of GRE-03 and GRE-04 TiO2 Control Chart	117
Figure 8-13: Summary of OREAS 460 and OREAS 464 TiO2 Control Chart	118
Figure 8-14: Summary of SX18-01, SX18-02, SX18-04, SX18-05 La Control Chart	119
Figure 8-15: Summary of SX18-01, SX18-02, SX18-04, SX18-05 Ce Control Chart	120
Figure 8-16: Summary (2011 – 2016 results) of SX18-01, SX18-02, SX18-04, SX18-05 Nd Control Chart	120
Figure 8-17: CRM AMIS0815 Control Charts for La, Ce, Pr, and Nd, and Dy (Provisional)	121
Figure 8-18: CRM GRE-03 and GRE-04 Control Charts for La and Ce	123
Figure 8-19: CRM GRE-03 and GRE-04 Control Charts for Pr and Nd	123
Figure 8-20: CRM GRE-03 and GRE-04 Control Charts for Dy	124
Figure 8-21: CRM OREAS 460 and OREAS 464 La, Ce, Pr, and Nd, and Dy Control Chart	125
Figure 8-22: CRM OREAS 460 and OREAS 464 Dy Control Chart	126
Figure 8-23: Paired Relative Difference and an XY Scatter Comparison of Original Versus Pulp Duplicate (Riffle Split) Analysis for Nb2O5 (%), Sc (ppm), and TiO2 (%)	127
Figure 8-24: Paired Relative Difference and an XY Scatter Comparison of Original Versus Pulp Duplicate (Riffle Split) Analysis for LREE, La (ppm), Ce (ppm), Pr (ppm), and Nd (ppm)	128
Figure 8-25: Paired Relative Difference and an XY Scatter Comparison of Original Versus Pulp Duplicate (Riffle Split) Analysis for HREE Representative, Dy (ppm)	129

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Technical Report Summary xv

Figure 8-26: Paired Relative Difference and an XY Scatter Comparison of Original Versus Coarse-Reject Duplicate (Riffle Split) Analysis for Nb2O5 (%), Sc (ppm), and TiO2 (%)	130
Figure 8-27: Paired Relative Difference and an XY Scatter Comparison of Original Versus Coarse-Reject Duplicate (Riffle Split) Analysis for LREE, La (ppm), Ce (ppm), Pr (ppm), and Nd (ppm)	131
Figure 8-28: Paired Relative Difference and an XY Scatter Comparison of Original Versus Coarse-Reject Duplicate (Riffle Split) Analysis for HREE Representative, Dy (ppm)	132
Figure 8-29: Paired Relative Difference and an XY Scatter Comparison of Original Versus Quarter (1/4-Core) Core Duplicate (Riffle Split) Analysis for Nb2O5 (%), Sc (ppm), and TiO2 (%)	133
Figure 8-30: Paired Relative Difference and an XY Scatter Comparison of Original Versus Quarter (1/4-Core) Core Duplicate (Riffle Split) Analysis for LREE, La (ppm), Ce (ppm), Pr (ppm), and Nd (ppm)	134
Figure 8-31: Paired Relative Difference and an XY Scatter Comparison of Original Versus Quarter (1/4-Core) Core Duplicate (Riffle Split) Analysis for HREE Representative, Dy (ppm)	135
Figure 8-32: Inspectorate Labs (2011) Paired Relative Difference and an XY Scatter Comparison of Original Versus External Pulp Duplicate (Riffle Split) Analysis for Nb2O5 (%).	135
Figure 8-33: 2015 Duplicate Sample Grade Range Selection Charts for duplicate re-submission, Targeting Nb2O5, Sc, and TiO2.	136
Figure 8-34: SGS External Lab SX18-01, SX18-02, SX18-04, and SX18-05 Nb2O5, Control Charts	137
Figure 8-35: SGS External Lab SX18-01, SX18-02, SX18-04, and SX18-05 TiO2, Control Charts	138
Figure 8-36: SGS External Lab GRE-04 Nb2O5, Sc, and TiO2 Control Charts	138
Figure 8-37: SGS (2014-2015) Labs Paired Relative Difference and an XY Scatter Comparison of Original Versus External Pulp Duplicate (Riffle Split) Analysis for Nb2O5 (%), Sc (ppm), and TiO2 (%)	139
Figure 8-38 : Comparison of Density Measurements Using Volume	141
Figure 9-1: (left) 2022 core review at the Elk Creek Project. (right) Split Core from NEC11-002	142
Figure 9-2: Example of visited physical drill collars: NEC14-009 (top left), NEC-14-013 (top right), NEC14-016 (bottom left), and NEC15-005 (bottom right)	143
Figure 10-1: Average Estimated Precipitation Versus Dilution Ratio	151
Figure 11-1: Niobium concentration box plot by logged lithology	168
Figure 11-2: Titanium concentration box plot by logged lithology	169
Figure 11-3: Scandium concentration box plot by logged lithology	169
Figure 11-4: Plan view of the MCarb domain (upper image) and cross section looking northwest of the MCarb domain and the modelled overlying sediments (lower image). Both diagrams contain the informing drill holes displaying lithology logs	170
Figure 11-5: Oblique view looking northwest drill holes displaying TREO (%) assay results and outline of the MCarb domain	171
Figure 11-6: Plan view of the MCarb, SW, and NE domains (upper image) and cross section looking northwest of the MCarb, SW, and NE domains (lower image). Both diagrams contain drill holes traces displaying TREO assay results	172
Figure 11-7: Histogram of assay lengths within the MCarb, SW, and NE domains	173
Figure 11-8: Probability plot and histogram of uncapped and capped Sc (ppm) composite distributions in the MCarb domain	175

NioCorp Developments Ltd.

Elk Creek Project S-K 1300

Technical Report Summary xvi

Figure 11-9: Correlation matrices of the homotopic observations of capped oxides and scandium distributions within the MCarb (Bound 1), SW (Bound 2), and NE (Bound 3) domains	177
Figure 11-10: Bivariate plots of the homotopic composite observations of niobium, titanium, scandium, and density distributions within the MCarb (Bound 1) domain	178
Figure 11-11: Box plot and histogram of the specific gravity grouped by logged lithology	179
Figure 11-12: Box plot and histogram of the specific gravity grouped by estimation domains	179
Figure 11-13: Fan variogram of niobium	180
Figure 11-14: Omni-direction experimental and model variogram for scandium in the MCarb domain	181
Figure 11-15: Diagram demonstrating the change of support principle (Harding & Deutsch, 2019)	184
Figure 11-16: OK model of TiO2 (MCarb domain) using different number of composites per estimate compared against the target mean and variance	184
Figure 11-17: Contribution percentage of scandium revenue to NSR (Diluted) from block model	185
Figure 11-18: Cross section looking northwest of the unsmoothed and smoothed OK TiO2 models within the MCarb Domain	186
Figure 11-19: Cross section looking northwest of the DDH constrained and unconstrained OK TiO2 models within the MCarb Domain	187
Figure 11-20: The summed estimates of the individual REOs that constitute the LREO, MREO, HREO, and TREO variables versus the estimated LREO, MREO, HREO, and TREO values	189
Figure 11-21: Cross section looking northwest of LREO composites and blocks within the SW domain showing the constrained high-grade blocks	191
Figure 11-22: Histogram comparison of blocks relative to DGM target distributions	192
Figure 11-23: Plan view of the Elk Creek domains and long sections looking southwest of the niobium, titanium, scandium, and TREO block model grades with informing composite grades. Blocks shown are restricted to Indicated and Inferred Mineral Resource material	193
Figure 11-24: Niobium, titanium, scandium, TREO, and density swath plots in the MCarb domain	195
Figure 11-25: Bivariate plots of block model estimate of niobium, titanium, scandium, and density distributions within the MCarb domain	195
Figure 11-26: Long-section looking southwest of the 2022 Inferred and Indicated Domains underlain with niobium composites	196
Figure 11-27: Plan section (elevation of 25) of the 2022 Inferred and Indicated domains with niobium composites displayed as spheres	197
Figure 11-28: Grade-tonnage curve of individual REOs for Indicated Mineral Resources	204
Figure 11-29: Grade-tonnage curve of individual REOs for Inferred Mineral Resources	205
Figure 12-1: Sources of Mining Dilution for Typical Stope Layout (Not to Scale)	207
Figure 12-2: NioCorp Grade (Nb2O5)-Tonne Curves Based on NSR Cut-Off	210
Figure 12-3: NioCorp Grade/Tonne Curves Based on NSR Cut-Off (TiO2)	210
Figure 12-4: NioCorp Grade (Sc ppm) - Tonne Curves Based on NSR Cut-Off	211
Figure 12-5: Completed Mine Design	212
Figure 13-1: Undiluted Stope Optimization Results for Varying NSR Cut-Offs	216
Figure 13-2: Stopes and Cross-Cut Accesses (Cross Section View)	218

NioCorp Developments Ltd.

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Technical Report Summary xvii

Figure 13-3: Level Layout with Stopes and Footwall Accesses (Rotated View Looking North)	218
Figure 13-4: Completed Mine Design (Plan View)	219
Figure 13-5: Completed Mine Design - Main Infrastructure (Looking South)	220
Figure 13-6: Mine Design Coloured by Nb2O5 Grade	221
Figure 13-7: Mine Design Coloured by NSR	222
Figure 13-8: Underground Mine Access Via Twin Concrete Lined Shafts	225
Figure 13-9: Production Shaft Layout	226
Figure 13-10: Ventilation Shaft Layout	228
Figure 13-11: Mine Production Schedule - Coloured By Year	234
Figure 13-12: Haulage Distance — One-Way Length	239
Figure 13-13: Haulage Cycle Time – Roundtrip	239
Figure 13-14: Backfill Production during LOM.	240
Figure 13-15: The Critical Areas Requiring Further Infill Definition Drilling	243
Figure 13-16: Elaboration of the Reconciliation Process Defining Additional Steps	245
Figure 13-17: Garage Area Ventilation Controls (Looking Northwest)	251
Figure 13-18: Level Ventilation Controls (Plan View)	252
Figure 13-19: Typical Pumping System (-15.4m and -335.4 m)	256
Figure 13-20: Production Shaft Hoisting Plant	261
Figure 13-21: Ventilation Shaft Hoisting Plant	262
Figure 14-1: HPGR Conceptual Block Flow Diagram	273
Figure 14-2: Simplified Sheet	275
Figure 14-3: Pyrometallurgical Processing Simplified Flowsheet	281
Figure 14-4: Process Water Treatment Plant Block Flow Diagram	299
Figure 15-1: Elk Creek Project Site Plan Layout	302
Figure 15-2: Process Water Treatment Plant Block Flow Diagram	310
Figure 15-3: Tailings Storage Facility Layout Showing Plant Site Cells 1, 2 and 3 and Area 7 Cell 1	322
Figure 15-4: Tailings and Waste Rock Storage Area Embankment Cross-Section	323
Figure 15-5: Tailings Storage Facility Central and Toe Drain Details	324
Figure 15-6: Leachate Collection Pond Embankment Cross-Section	325
Figure 15-7: Leachate Collection Pond LCRS System	325
Figure 15-8: Paste Backfill Plant process flow diagram	331
Figure 15-9: Typical Freeze Plant Configuration (with gangs of compressors and cooling coils in series to make up the total capacity of the plant)	333
Figure 15-10: Typical Layout of a Freezewall Borehole System	334
Figure 16-1: CMBB Niobium Sales Versus Steel Demand and Niobium Intensity of Use	337
Figure 16-2: Ferroniobium (65% - EU) Price Trends Previous Quarter	338
Figure 16-3: Global Titanium Dioxide Market Value and Volume 2014-2025	339

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Technical Report Summary xviii

Figure 16-4: High, Expected, and Low Case Forecasts for Scandium Oxide Potential Supply 2019 – 2030, Tonnes per Year	342
Figure 16-5: High, Expected, and Low Case Forecasts for Scandium Oxide Potential Supply 2019 – 2030, Tonnes per Year, Excluding Russia and China	343
Figure 16-6: Current/Potential Scandium Market	344
Figure 16-7: Supply-Demand Forecast for Scandium Oxide to 2032, Tonnes, Base Case	346
Figure 16-8: Scandium Oxide Pricing Outlook, US$/kg, 2019 – 2030	347
Figure 16-9: Global Scandium Supply/Demand and Price Projections Summary	348
Figure 16-10: REE uses by volume and by value	350
Figure 16-11: Historical global consumption and forecasted demand for NdFeB magnets by end-use category	350
Figure 16-12: NdFeB magnet demand forecast for passenger EV traction motors	351
Figure 16-13: NdFeB magnet demand forecast for passenger EV traction motors	352
Figure 16-14: Rare earth oxide production by region	353
Figure 16-15: Relative price forecast for individual REEs	354
Figure 16-16: Pricing forecasts out to 2030 for the magnet feed rare earth oxides	355
Figure 18-1: Elk Creek Project LOM organizational chart	404
Figure 19-1: Annual Project Metrics Summary (Pre-Tax)	417
Figure 19-2: Annual Project Metrics Summary (After-Tax)	417
Figure 19-3: Pre-Tax NPV 8% Sensitivity Graph	418
Figure 19-4: After-Tax NPV 8% Sensitivity Graph	419
Figure 19-5: Pre-Tax IRR Sensitivity Graph	419
Figure 19-6: After-Tax IRR Sensitivity Graph	420
Figure 19-7: Pre-Tax NPV 8% Sensitivity Graph	420
Figure 19-8: After-Tax NPV 8% Sensitivity Graph	421
Figure 19-9: Pre-Tax IRR Sensitivity Graph	421
Figure 19-10: After-Tax IRR Sensitivity Graph	422
Figure 19-11: Before-Tax NPV Profile	423
Figure 19-12: After-Tax NPV Profile	424
Figure 21-1: Summary Level – Owner’s Project Team	429
Figure 22-1: Likelihood and Consequence Matrix	441
Figure 23-1: Resource Area Drillhole Intervals Showing Assay Coverage (Red) and Assay Gaps (Blue)	447

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Elk Creek Project S-K 1300

Technical Report Summary xix

List of Tables

Table 1-1: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) excluding reserves	5
Table 1-2:Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) excluding reserves	5
Table 1-3: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) including reserve material	7
Table 1-4: Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) including reserve material	7
Table 1-5: Underground In Situ Mineral Reserves Estimate for Elk Creek, Effective Date June 30, 2022	9
Table 1-6: Capital Costs Summary (US$ 000’s)	13
Table 1-7: LOM Operating Cost Unit Rate Summary	14
Table 2-1: Site Visit Participants	16
Table 3-1: Active Option to Purchase Agreements Covering the Project	20
Table 4-1: Summary of the Project Precipitation Data (4) (5)	24
Table 4-2: Summary of Hydrometallurgical Process Water Requirement	27
Table 4-3: Pyrometallurgical Water Requirements	27
Table 6-1: Project Rock Types as Defined by Molycorp and Dahrouge (2011)	43
Table 6-2: Stratigraphy Overlying the Elk Creek Carbonatite	45
Table 6-3: List of Elements and Oxides Associated REE Mineralization	50
Table 7-1: Summary of Drilling Database within the Geological Complex	55
Table 7-2: Summary of Drilling Database within Elk Creek Deposit Area	56
Table 7-3: Summary of Drilling Database used in the Current Resource Estimation	56
Table 7-4: Summary of the 2011 Drill Program	58
Table 7-5: NioCorp 2014-2015 Drill Hole Locations	59
Table 7-6: Drill Hole Orientation and Data Collection Methods	68
Table 7-7: Summary of Rock Mass Characterization by Domain	71
Table 7-8: Discontinuity Orientation Data for 2014 Geotechnical Investigation	72
Table 7-9: Barton Parameters for Different Excavations	79
Table 7-10: Preliminary Support According to Barton Method	79
Table 8-1: Core Inventory of Drillholes within the Resource Area at the Mead Facility	95
Table 8-2: Summary of Major Rock Unit Codes	97
Table 8-3 Summary of 2010, ALS Labs, Re-Sampling Program Submissions	100
Table 8-4: Summary of 2014-2015, Targeted Sc Re-Sampling Program Submission to SGS Labs	101
Table 8-5: Summary of 2016, Re-Sampling Program Submissions to Act Labs	101
Table 8-6: Pre-2021 Missing REE and Sc Assays that have Nb2O5 Database Results	102
Table 8-7: Pre-2021 Missing REE and Sc Assays that have Nb2O5 Database Results	103
Table 8-8: Detection Limits for Primary Laboratory (Actlabs)	105

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Technical Report Summary xx

Table 8-9: Summary of Actual Submissions per Sample Type within the 2015-2021 Re-Assay Program	107
Table 8-10: Summary of Actual Submissions per Sample Type within the 2015-2021 Re-Sample Program	107
Table 8-11: Summary of Designed Level of Insertion of QC Submissions (2011 and 2014 Drill Program)	108
Table 8-12: Summary of Sample and Control Submissions for Nb2O5, Sc TiO2, REE’s (2011 Drill Program)	108
Table 8-13: Summary of Sample and Control Submissions for Nb2O5, Sc TiO2, REE’s (2014 Drill Program)	109
Table 8-14: Summary of 2011 and 2014 Drill Program Nb2O5 Blank Insertion	110
Table 8-15: Summary of CRM & SRMs Controls Used on the Project	113
Table 8-16: Summary of Nb2O5 CRM (Primary Assays - Actlabs)	114
Table 8-17: Summary of Sc CRMs (Primary Assays – Actlabs + SGS)	115
Table 8-18: Summary of Sc CRMs (Primary Assays – Actlabs)	117
Table 8-19: Summary of SX18-01, SX18-02, SX18-04, SX18-05 REE Results for SRM’s Inserted During 2011-2014	118
Table 8-20: AMIS0815 summary of REE Results of SRMs Inserted During 2011-2014	121
Table 8-21: GRE-03 summary of REE Results Inserted During 2011-2021	122
Table 8-22: GRE-04 summary of REE Results Inserted During 2011-2014	122
Table 8-23: Oreas 460 REE Results (2021) Summary of Results	124
Table 8-24: Oreas 464 REE Results (2021) Summary of Results	124
Table 8-25: SGS (2014-2015) External Lab Sample Summary, for Nb2O5, Sc, and TiO2.	137
Table 9-1: Confirmed collar locations	143
Table 9-2: Drill hole Assay Intervals checked against Laboratory Certificates	144
Table 10-1: Whole Ore Sample Head Assays	147
Table 10-2: PP1-013 Extraction Summary	148
Table 10-3: PP2-013 Extraction Summary	148
Table 10-4: HCI Leach - Summary Design & Extraction Extent	148
Table 10-5: PP1-015 Acid Bake and Water Leach Extractions	149
Table 10-6: Acid Bake and Water Leach - Extraction Results	150
Table 10-7: Iron Reduction	150
Table 10-8: Niobium Precipitation - Elemental Extractions	151
Table 10-9: Phosphate Removal - Summary Extractions	152
Table 10-10: Titanium Precipitation - Elemental Extractions	153
Table 10-11: Scandium Precipitation Summary	154
Table 10-12: Scandium Precipitation - Elemental Extractions	155
Table 10-13: Sulphate Calcining Results Summary	156
Table 10-14: PP2 Overall Metal Distribution	158
Table 10-15: PP3 Overall Metal Distribution	158
Table 10-16: Scandium Refining - Impurity Extractions	159
Table 10-17: PP1 Composite Discharge Solids Assays	160

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Technical Report Summary xxi

Table 10-18: Recovery Summary	160
Table 11-1: Dahrouge lithology codes	165
Table 11-2: Assay variables	166
Table 11-3: Elemental percentage conversion ratios to oxide percentage	166
Table 11-4: Capping levels	174
Table 11-5: Summary statistics of the variables per domain	176
Table 11-6: Variogram models for select variables within the MCarb domain	181
Table 11-7: Block model variables	182
Table 11-8: Comparison of block grades to representative distribution	190
Table 11-9: Block to wireframe volume comparison	191
Table 11-10: Mining cost assumptions	197
Table 11-11: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) excluding reserves	198
Table 11-12: Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) excluding reserves	199
Table 11-13: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) including reserve material	200
Table 11-14: Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) including reserve material	201
Table 11-15: Grade/tonnage by diluted NSR cut-off	203
Table 12-1: Sources of Mining Dilution for Typical Geometry by Stope Type	208
Table 12-2: Example of an NSR Block Calculation	209
Table 12-3: Operating Costs Used for Mine Design NSR Cut-off	211
Table 12-4: Underground In Situ Mineral Reserves Estimate for Elk Creek, Effective Date June 30, 2022	213
Table 13-1: Undiluted Stope Optimization Results for Varying NSR Cut-offs	217
Table 13-2: Mine Design Summary - by Activity Type	222
Table 13-3: Productivity Rates	229
Table 13-4: Dimensions by Heading Types	229
Table 13-5: Workforce Schedule Parameters for Underground Mining	230
Table 13-6: Ground Support Requirements	230
Table 13-7: Mine Production Schedule	232
Table 13-8: LHD Hauling parameters upper mining block	236
Table 13-9: LHD Hauling parameters lower mining block	237
Table 13-10:Truck Hauling parameters lower mining block	237
Table 13-11: Backfill Volume Summary - By Type	240
Table 13-12: Extra Support Assumptions by Heading Type	241
Table 13-13: Typical Mining Labour	246
Table 13-14: Mine Mobile Equipment	247
Table 13-15: Mine Fixed Equipment	248

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Technical Report Summary xxii

Table 13-16: Maximum Airflow Velocities (m/s)	250
Table 13-17: Duties of Main Surface Fans	252
Table 13-18: Total Mine Airflow	253
Table 13-19: Surface Temperatures Near the Elk Creek Mine	254
Table 14-1: Process Design Criteria	265
Table 14-2: Hydrometallurgical Processing Design Criteria	266
Table 14-3: Pyrometallurgical Processing Design Criteria	269
Table 14-4: Primary Equipment List	283
Table 14-5: Ancillary Equipment List	283
Table 14-6: Summarized List of Equipment	284
Table 14-7: Pyrometallurgical Processing Major Equipment List	290
Table 14-8: Acid Plant Equipment List	291
Table 14-9: Hydromet Power Requirements	292
Table 14-10: FeNb Furnace Power Requirements	293
Table 14-11: Design Requirements	294
Table 14-12: Summary of Hydrometallurgical Process Water Requirement	300
Table 14-13: Pyrometallurgical Water Requirements	301
Table 15-1: Design Requirements	304
Table 15-2: Summary of Hydrometallurgical Process Water Requirement	311
Table 15-3: Pyrometallurgical Water Requirements	312
Table 15-4: TSF Area, Storage and Time Characteristics	316
Table 15-5: LCP Area, Storage and Time Characteristics	316
Table 15-6: Tailings Storage Facility Stage-Area-Capacity Data	319
Table 15-7: Mean Monthly Average Precipitation	320
Table 15-8: SMC Footprint Areas and Storage Capacities	326
Table 15-9: Paste Backfill Formulations for Phase 2 Testwork	329
Table 15-10: Results of Phase 2 UCS Testing of paste backfill samples after 7 days	329
Table 15-11: Paste Backfill Formulations for Phase 3 Testwork.	329
Table 15-12: Results of Phase 3 UCS Testing of paste backfill samples after 7 days	330
Table 16-1: Comparison of Project Versus Selected Niobium Producers	336
Table 16-2: Titanium Mineral Concentrates Pricing History (Rutile Concentrate FOB Australia)	340
Table 16-3: Scandium Supply, Demand and Price Forecast Summary	348
Table 17-1: Project Permits	367
Table 18-1: Capital Costs Summary (US$ 000’s)	378
Table 18-2: Capitalized Pre-production Cost Summary (US$ 000’s)	379
Table 18-3: Initial Direct Mine Capital Cost Estimate (US$ 000’s)	379
Table 18-4: Initial Direct Mine Capital Cost Contingency Estimate	380

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Technical Report Summary xxiii

Table 18-5: Mining Indirect Cost	381
Table 18-6: Process Plant Costs Summary	381
Table 18-7: Processing Indirect Costs Summary	382
Table 18-8: Pre-production Tailings and Salt Management Facility Construction Cost	384
Table 18-9: Tailings and Salt Placement Equipment Pre-Production Capital	384
Table 18-10: Pre-production Temporary Waste Rock Storage Facility Construction Cost	385
Table 18-11: Water Management and Infrastructure Cost	386
Table 18-12: Site Preparation and Infrastructure Costs Summary	386
Table 18-13: Site Wide Indirect Costs Summary	388
Table 18-14: Owner’s Costs Summary	389
Table 18-15: Sustaining Capital for Mining (US$ 000’s)	392
Table 18-16: Sustaining Capital for Process and Infrastructure (US$)	393
Table 18-17: Tailings and Salt Facility Sustaining Capital Cost	393
Table 18-18: Tailings Placement Equipment LOM Replacement Capital	394
Table 18-19: Initial Capital Contingency Summary	394
Table 18-20: LOM Operating Cost Unit Rate Summary	396
Table 18-21: Steady State Mining Operating Unit Cost (after pre-production)	397
Table 18-22: ROM Processing Operating Cost Unit Rate Breakdown	398
Table 18-23: Tailings Haulage Calculations	399
Table 18-24: Tailings and Salt Mobile Equipment Hourly Operating Cost	400
Table 18-25: Tailings and Salt Mobile Equipment Utilization	400
Table 18-26: Cost for Tailings and Salt Placement	401
Table 18-27: LOM Site G&A Operating Costs	401
Table 18-28: Site G&A Annual Operating Costs	402
Table 18-29: Average Annual G&A Headcount during Operations	403
Table 18-30: Average Annual G&A Fixed Costs During Operations	404
Table 19-1: General Assumptions	407
Table 19-2: Mining Physicals	409
Table 19-3: Processing Physicals	409
Table 19-4: Processing Recovery Summary	410
Table 19-5: Pricing Assumptions	410
Table 19-6: Scandium Trioxide Pricing Assumptions	410
Table 19-7: Operating Cost Summary	412
Table 19-8: Capital Cost Summary (US$ 000’s)	412
Table 19-9: Initial Capital Cost Summary	413
Table 19-10: Indicative Economic Results (US$ 000’s unless otherwise indicated)	414
Table 19-11: Elk Creek Economic Results 2022	415

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Technical Report Summary xxiv

Table 19-12: Indicative Economic Results	416
Table 19-13: Niobium Price Sensitivity (Sc and Ti Prices Remain Constant)	422
Table 19-14: Scandium Price Sensitivity (Nb and Ti Prices Remain Constant)	423
Table 21-1: Key Project Milestones	427
Table 23-1: 2011 and 2014 Drillhole Intervals Not Sampled and their Priority (Low to High)	446
Table 23-2: Summary of Costs for Recommended Work	454
Table 24-1: Definition of Terms	462

NioCorp Developments Ltd.

Elk Creek Project

S-K 1300 Technical Report Summary

1. EXECUTIVE SUMMARY

This is a Technical Report Summary (TRS) for the Elk Creek Project that has been prepared in accordance with item 601(b)(96) and subpart 1300 of Regulation S-K (SK 1300) promulgated by the Securities and Exchange Commission (SEC). The purpose of this TRS is to report mineral resources and mineral reserves for the Elk Creek Project. The effective date of this report is June 30, 2022. The report has been prepared by Dahrouge Geological Consulting USA Ltd. (Dahrouge), Understood Mineral Resources Ltd. (Understood), Optimize Group (Optimize), Tetra Tech, Adrian Brown Consultants Inc. (ABC), Metallurgy Concept Solutions (MCS), Magemi Mining Inc. (Magemi), L3 Process Development (L3), A2GC, Scott Honan. M.Sc, SME-RM, NioCorp, Everett Bird, P.E., Cementation, Matt Hales, P.E., Cementation, Mahmood Khwaja, P.E., CDM Smith, Martin Lepage, P.Eng, Ing., Cementation and Wynand Marx, M.Eng, BBE Consulting. A full list of Qualified Persons and the sections and sub-sections that they are responsible for is presented in Section 27.

NioCorp is a U.S.-based mineral development company focused on developing several critical minerals from the proposed Elk Creek, Nebraska Critical Minerals Mine. NioCorp plans to produce three commercial mineral products — niobium, scandium, and titanium - from a single ore body that is also enriched in all of the rare earth elements. All three of the Project’s proposed superalloy metals have been designed as “Critical Minerals” by the U.S. Government, as have the rare earth elements. NioCorp is a publicly traded company whose common shares are listed on both the Toronto Stock Exchange under the ticker symbol “NB” and are quoted on the U.S.-based OTCQX exchange under the symbol “NIOBF.” NioCorp’s common shares are also traded on the Frankfurt Exchange, under the ticker symbol “BR3.”

The TRS uses Canadian English and metric units unless otherwise indicated. Monetary units are in United States Dollars (US$).

1.1 Principal Outcomes

The TRS is based on an assumption of processing 36,656 kilo tonnes (kt) of ore over a 38-year life of mine (LOM) to produce 171,140 tonnes (t) of niobium (Nb) in the form of ferroniobium (FeNb), 3,676 tonnes of scandium trioxide (Sc₂O₃) and 431,793 tonnes of titanium dioxide (TiO₂). Rare earths have been added to the Mineral Resource, and the Indicated Resource contains 632.9 kt of Total Rare Earth Oxide (TREO). A breakdown by component is provided below:

● 26.9 kt of praseodymium

● 98.9 kt of neodymium

● 2.3 kt of terbium

● 9.1 kt of dysprosium

● 970.3 kt of niobium oxide

● 11,337 t of scandium oxide

● 4,221 kt of titanium oxide

This has been estimated using a ≥ US$180/tonne Net Smelter Return (NSR) cut-off that was calculated using solely the contained niobium, scandium, and titanium in the Mineral Resource.

Initial capital costs are estimated at US$ 1,141 million.

Total capital cost, inclusive of sustaining, closure/reclamation, and contingency costs is US$ 1,606 million.

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Total LOM operating costs are estimated to be US$ 7,182 million. Costs and product pricing are presented using a 2019 basis.

On a pre-tax basis, the NPV (8% discount) is US$ 2,819 million, the IRR is 29.2%, and the assumed payback period is within 2.67 years.

On a post-tax basis, the NPV (8% discount) is US$ 2,350 million, the IRR is 27.6%, and the assumed payback period is within 2.69 years.

1.2 Property Description and Ownership

The Elk Creek Project is a greenfield exploration project located in southeast Nebraska, USA. It is located approximately 75 km (47 miles) southeast of Lincoln, Nebraska (the state capital), and 110 km (68 miles) south of Omaha, Nebraska. The mineralization is centred about 40°16’0.3.5” N latitude and 96°11’08.5” E longitude. The area is well developed with direct access to roads, rail, supply and distribution companies, and a local workforce including heavy equipment operators. Geologists can be sourced from local universities. An experienced mining-related workforce can be found in Weeping Water, Nebraska as well as Denver, Colorado (eight-hour drive west of the Project). The deposit is located within the U.S. Geological Survey (USGS) Tecumseh Quadrangle Nebraska SE (7.5 minute series) mapsheet in Sections 1-6, 9-11; Township 3N; Range 11 and Sections 19-23, 25-36; Township 4N, Range 11.

The Property consists of one 91.5 ha (226 acre) parcel of land owned by NioCorp Developments Ltd. along with 8 Option To Purchase agreements (OTP) covering approximately 565 hectares (ha). OTPs are between NioCorp’s subsidiary Elk Creek Resources Corp. (ECRC) and the individual landowners. The parcel owned by the Company contains the majority of the Mineral Resources and Mineral Reserves associated with the Project. ECRC is a Nebraska corporation and wholly owned subsidiary of NioCorp. NioCorp owns or has the option to purchase 100% of the mineral rights to the Project and is the operator. The OTPs allow ECRC to purchase the mineral rights and/or the surface rights at any time during the term of the agreement. The individual landowners have title to the surface and subsurface rights, and the agreements are primarily concerned with only the mineral and surface interest of each property. The agreements convey to the Company adequate surface rights to access the land and to complete mineral exploration work. The option agreements that the Company currently holds combined with the land owned by the Company include all the Indicated and Inferred Resources and Probable Reserves described in this TRS.

The options covering the Project are 100% owned by NioCorp and, apart from a 2% NSR royalty attached with the OTPs that include the mineral rights, have no other outstanding royalties, agreements, or encumbrances. The 226-acre parcel of land owned by the Company is also subject to a 2% NSR royalty.

1.3 Geological Setting and Mineralization

The Project includes the Elk Creek Carbonatite (the Carbonatite) that intruded older Precambrian granitic and low to medium grade metamorphic basement rocks. Both the Carbonatite and Precambrian rocks are interpreted to be unconformably overlain by approximately 200 meters (m) of Paleozoic marine sedimentary rocks of Pennsylvanian age. As a result of this thick cover, there is no surface outcrop within the Project area of the Carbonatite, which was identified and targeted through magnetic surveys and confirmed through subsequent drilling. The available magnetic data indicates dominant northeast, west-northwest striking lineaments, and secondary

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northwest and north-oriented features that mimic the position of regional faults parallel and/or perpendicular to the Nemaha Uplift.

The Carbonatite hosts significant niobium (reported as Nb₂O₅), titanium (reported as TiO₂) and scandium (reported as Sc) and is composed predominantly of dolomite, calcite and ankerite, with lesser chlorite, barite, phlogopite, pyrochlore, serpentine, fluorite, sulphides and quartz. Niobium is contained primarily within the mineral pyrochlore, and rare earth element (REE) mineralization is reported to occur as bastnäsite, parisite, synchysite and monazite.

1.4 History

Exploration activities at the Project prior to NioCorp ownership were conducted by the following companies: University of Nebraska – Lincoln, Nebraska Conservation and Survey Division, United States Geological Survey (USGS), Cominco American Incorporated (Cominco American), Molybdenum Corporation of America, later Molycorp Inc. (Molycorp) and Quantum Rare Earth Developments Corp. (Quantum). These activities consisted of airborne magnetic and gravity surveys, geochemical sampling, reverse circulation (RC) drilling, core drilling and mineral resource estimates.

Since 2014, NioCorp has completed metallurgical testing, core drilling, mineral resource estimates in 2014, a mineral resource estimates in 2015, two Preliminary Economic Assessments in 2015, and Feasibility Studies in 2017,2019 and 2022. This report is the first TRS for the Project.

1.5 Drilling

Drilling at the Project was conducted in four phases. The first was during the 1970s and 1980s by Molycorp, the second in 2011 by Quantum, the third in 2014 by NioCorp, and the fourth, not sampled for resource estimation, but for hydrogeological and geotechnical studies, by NioCorp in 2015. To date, 143 diamond core holes have been completed for a total of 70,897 m over the entire geological complex. Of these, a total of 54 holes (37,747 m) have been completed to date in the deposit area of which 45 are used in the current Mineral Resource.

An additional five drill holes, totalling 3,353.1 m, were completed in 2015. This drilling was for the purpose of hydrogeological and geotechnical studies. These holes were not used in the geological modelling for the Mineral Resource.

All drilling has been completed using a combination of Tricone, RC or Diamond Drilling (DD) in the upper portion of the hole within the Pennsylvanian sediments. All drilling within the underlying Carbonatite has been completed using DD methods.

Before the core was split for sampling, depth markers were checked, the core was carefully reconstructed, washed, geotechnically and geologically logged for lithologies, alteration, structures, and mineralization, rock quality designation (RQD), photographed, and marked for sampling. A sampling of the holes for assay was guided by the observed geology. Logging and sampling information was entered into a database template on a computer which was integrated into the Project master digital database on a daily basis. The database is summarized in Microsoft Excel^® .csv spreadsheets containing collar locations surveyed in UTM coordinates, downhole deviation surveys, assay intervals with elemental analyses, geologic intervals with rock types, alteration, and key structures. Further details are provided in Section 8.

Core recovery at the Project has allowed for representative samples to be taken and accurate analyses to be performed. All NioCorp drill hole collars have been surveyed using a Sokkia GS2700 IS GPS, which has 10 mm horizontal and 20 mm vertical accuracy. The trajectory of all drill holes

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was determined during drilling with either a Devico DeviFlex survey tool, which is a nonmagnetic, electronic, multi-shot tool or a Reflex Gyro survey tool. Data points were collected at either 3.05 m or 6.1 m intervals. The 2014/15 program also used ATV acoustic Teleview (ATV) downhole equipment to collect various geological features.

Dahrouge is not aware of any drilling, sampling, or recovery factors that could materially impact the accuracy and reliability of the results. In Dahrouge’s opinion, the drilling, core handling, logging, and sampling procedures meet or exceed industry standards and are adequate for the purpose of Mineral Resource Estimation.

1.6 Mineral Resource Estimation

The 2022 Mineral Resource Estimate for the Elk Creek Deposit was completed by Understood Mineral Resources Ltd. The effective date of the enclosed mineral resource is June 30, 2022.

Understood was supplied a drill hole database as individual spreadsheets. The database contains all 138 drill holes drilled on the property, including the 45 drill holes that define the Elk Creek Deposit, which were used to inform the 2022 Mineral Resource Estimate. A further 5 holes drilled in 2015 for hydrogeological and geotechnical testing were not assayed and did not form part of the Resource Estimate.

Three wireframes were created to domain the data for estimation: the magnetite carbonatite domain (MCarb), the domain southwest of the magnetite carbonatite domain (SW), and the domain northeast of the magnetite carbonatite (NE). The MCarb domain is of particular interest, as it hosts 71% of the reported Indicated Mineral Resources and 97% of the reported Reserves.

Samples were composited to 1 m lengths within the domains and high-grade outlier assay values were capped. Omni-directional variograms were constructed on the composited dataset and were used to support search ranges for estimation.

A block model was constructed to encompass the three domains using 5 m by 5 m by 5 m blocks. The variables of the blocks were populated using Ordinary Kriging (OK) as informed by the omni-directional variograms.

Understood validated the block model using swath plots, mean comparison, volumetric comparison, visual inspection, histogram comparison, bivariate plot comparisons, and correlation checks. Understood found grade continuity to be reasonable and confirmed that the block grades were reasonably consistent with local drill hole composite grades.

The 2022 Elk Creek Mineral Resource Estimate contains Indicated and Inferred Mineral Resources. The classification was assigned to regions of the block model based on the Qualified Person’s confidence and professional judgement related to the geological understanding and continuity of mineralization in conjunction with data quality, spatial continuity, block model representativeness, and data density. On average, the Indicated Mineral Resources are informed with a drill hole spacing between 50-75 m and extend approximately 35-50 m laterally beyond the last drill intercept; the Inferred Mineral Resources capture the sparser drilled areas with an average drill hole spacing of 75-125 m and extend approximately 50-75 m laterally beyond the last drill intercept.

The 2022 Mineral Resource Estimate for the Elk Creek Deposit adheres to the S-K 1300 classification system and was reported at a US$ 180 diluted NSR cut-off (NSR cut-off / cut-off grade / CoG). The Rare Earth Oxides (REOs) were evaluated as a potential by-product to the mining of niobium, titanium, and scandium; thus, the reported REOs are coincident with above-cut-off diluted NSR values as derived from the Nb₂O₅, TiO₂, and Sc estimates. Mineral Resources

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are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the Mineral Resource will be converted into a Mineral Reserve.

Understood is not aware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the Mineral Resource Estimate that is not discussed in this TRS.

A variety of factors may affect the 2022 Elk Creek Mineral Resource Estimate, including but not limited to: changes to product pricing assumptions, re-interpretation of geology geometry and continuity of mineralization zones, mining and metallurgical recovery assumptions, and additional infill or step out drilling. In Understood’s opinion, the estimation methodology is consistent with standard industry practice and the Indicated and Inferred Mineral Resource Estimates for Elk Creek are reasonable and acceptable.

Table 1-1: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) excluding reserves

Class	NSR Cutoff	Tonnage (Mt)
Indicated	180	151.7	Nb2O5 (%)	Nb2O5 (kt)
			0.43	649.8
			TiO2 (%)	TiO2 (kt)
			2.02	3,067
			Sc (ppm)	Sc (t)
			56.42	8,558
Inferred	180	108.3	Nb2O5 (%)	Nb2O5 (kt)
			0.39	426.6
			TiO2 (%)	TiO2 (kt)
			1.92	2,082
			Sc (ppm)	Sc (t)
			52.28	5,660

Table 1-2:Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) excluding reserves

Class	NSR Cut-off	Tonnage (Mt)
Indicated	180	151.7	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0766	116.2	0.1320	200.2	0.0140	21.3
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0511	77.5	0.0116	17.6	0.0040	6.0
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.0096	14.6	0.0011	1.6	0.0044	6.7

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			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0006	1.0	0.0015	2.2	0.0002	0.3
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.0010	1.5	0.0001	0.2	0.0187	28.4
			LREO (%)	LREO (kt)	HREO (%)	HREO (kt)	TREO (%)	TREO (kt)
			0.2737	415.2	0.0528	80.0	0.3265	495.2
Inferred	180	108.3	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0943	102.1	0.1576	170.6	0.0163	17.7
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0575	62.2	0.0116	12.6	0.0038	4.1
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.0090	9.8	0.0010	1.1	0.0042	4.6
			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0006	0.7	0.0014	1.5	0.0002	0.2
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.0010	1.1	0.0001	0.1	0.0182	19.7
			LREO (%)	LREO (kt)	HREO (%)	HREO (kt)	TREO (%)	TREO (kt)
			0.3257	352.6	0.0512	55.5	0.3769	408.1

Notes:

a. Classification of Mineral Resources in the above tables is in accordance with the S-K 1300 classification system. Mineral Resources in this table are reported exclusive of Mineral Reserves

b. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

c. The Mineral Resources are reported at a Diluted Net Smelter Return (NSR) Cut-off of US $180/tonne.

d. The diluted NSR is defined as:

●

● The diluted revenue from Nb₂O₅, TiO₂, and Sc per block used the following factors:

● Nb₂O₅ Revenue: a 94% grade recovery, a 0.696 factor to convert Nb₂O₅ to Nb, 82.36% assumption for plant recovery, and a US$ 39.60 selling price per kg of ferroniobium as of June 30, 2022.

● TiO₂ Revenue: a 94% grade recovery, a 40.31% assumption for plant recovery, and a US$ 0.88 selling price per kg of titanium oxide as of June 30, 2022.

● Sc Revenue: a 94% grade recovery, a 1.534 factor to convert Sc to Sc₂O₃, 93.14% assumption for plant recovery, and a US$ 3,675 kg selling price per kg of scandium oxide as of June 30, 2022.

● The diluted tonnes are a 6% increase in the total tonnes of the block.

e. Price assumptions for FeNb, Sc₂O₃, and TiO₂ are based upon independent market analyses for each product.

f. Numbers may not sum due to rounding. The rounding is not considered to be material.

g. Rare Earth Oxides (REO) were evaluated as a potential by-product to the mining of niobium, titanium, and scandium; thus the estimated values of the REOs are reported using the previously determined diluted NSR as derived from the Nb₂O₅, TiO₂, and Sc Mineral Resources and are assigned a price of $0.

h. The stated Light Rare Earth Oxides (LREO) grade (%) is the summation of La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), and Nd₂O₃ (%) estimates.

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i. The stated Heavy Rare Earth Oxides (HREO) grade (%) is the summation of Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), and Y₂O₃ (%) estimates.

j. The stated Total Rare Earth Oxide (TREO) grade (%) is the summation of LREO (%) and HREO (%).

k. The effective date of the Mineral Resource, including by-products, is June 30, 2022

Table 1-3: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) including reserve material

Class	NSR Cutoff	Tonnage (Mt)
Indicated	180	188.8	Nb2O5 (%)	Nb2O5 (kt)
			0.51	970.3
			TiO2 (%)	TiO2 (kt)
			2.24	4,221
			Sc (ppm)	Sc (t)
			60.06	11,337
Inferred	180	108.3	Nb2O5 (%)	Nb2O5 (kt)
			0.39	426.6
			TiO2 (%)	TiO2 (kt)
			1.92	2,082
			Sc (ppm)	Sc (t)
			52.28	5,660.20

Table 1-4: Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) including reserve material

Class	NSR Cut-off	Tonnage (Mt)
Indicated	180	188.8	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0773	145.8	0.1335	251.9	0.0143	26.9
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0524	98.9	0.0129	24.3	0.0046	8.6
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.0110	20.8	0.0012	2.3	0.0048	9.1
			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0007	1.3	0.0015	2.9	0.0002	0.3
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.0010	1.9	0.0001	0.3	0.0199	37.6
			LREO (%)	LREO (kt)	HREO (%)	HREO (kt)	TREO (%)	TREO (kt)
			0.2774	523.6	0.0579	109.3	0.3353	632.9

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Inferred	180	108.3	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0943	102.1	0.1576	170.6	0.0163	17.7
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0575	62.2	0.0116	12.6	0.0038	4.1
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.009	9.8	0.0010	1.1	0.0042	4.6
			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0006	0.7	0.0014	1.5	0.0002	0.2
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.001	1.1	0.0001	0.1	0.0182	19.7
			LREO (%)	LREO (kt)	HREO (%)	HREO (kt)	TREO (%)	TREO (kt)
			0.3257	352.6	0.0512	55.5	0.3769	408.1

Notes:

a. Classification of Mineral Resources in the above tables is in accordance with the S-K 1300 classification system. Mineral Resources in this table are reported inclusive of Mineral Reserves

b. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

c. The Mineral Resources are reported at a Diluted Net Smelter Return (NSR) Cut-off of US $180/tonne.

d. The diluted NSR is defined as:

●

● The diluted revenue from Nb₂O₅, TiO₂, and Sc per block used the following factors:

● Nb₂O₅ Revenue: a 94% grade recovery, a 0.696 factor to convert Nb₂O₅ to Nb, 82.36% assumption for plant recovery, and a US$ 39.60 selling price per kg of niobium as of June 30, 2022.

● TiO₂ Revenue: a 94% grade recovery, a 40.31% assumption for plant recovery, and a US$ 0.88 selling price per kg of titanium oxide as of June 30, 2022.

● Sc Revenue: a 94% grade recovery, a 1.534 factor to convert Sc to Sc₂O₃, 93.14% assumption for plant recovery, and a US$ 3,675 kg selling price per kg of scandium oxide as of June 30, 2022.

● The diluted tonnes are a 6% increase in the total tonnes of the block.

e. Price assumptions for FeNb, Sc₂O₃, and TiO₂ are based upon independent market analyses for each product.

f. Numbers may not sum due to rounding. The rounding is not considered to be material.

g. Rare Earth Oxides (REO) were evaluated as a potential by-product to the mining of niobium, titanium, and scandium; thus the estimated values of the REOs are reported using the previously determined diluted NSR as derived from the Nb₂O₅, TiO₂, and Sc Mineral Resources and are assigned a price of $0.

h. The stated Light Rare Earth Oxides (LREO) grade (%) is the summation of La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), and Nd₂O₃ (%) estimates.

i. The stated Heavy Rare Earth Oxides (HREO) grade (%) is the summation of Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), and Y₂O₃ (%) estimates.

j. The stated Total Rare Earth Oxide (TREO) grade (%) is the summation of LREO (%) and HREO (%).

k. The effective date of the Mineral Resource, including by-products, is June 30, 2022

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1.7 Mineral Reserve Estimation

The Project is currently in the exploration phase and has not been developed. Based on geotechnical information and mineralization geometry, an underground longhole stoping method (LHS) has been determined to be suitable for the deposit. Paste backfill will be used to allow for a high recovery of ore material.

The stopes’ dimensions are 15 m wide, and stope length varies based on Nb₂O₅ mineralization grade to a maximum of 25 m per panel with a level spacing of 40 m. The variation on stope length allowed for optimization of the Nb₂O₅ grade with a minimal increase to operating costs. The level spacing of 40 m was beneficial to operating and sustaining capital costs. Each block is mined with a bottom-up sequence. A partial sill pillar level is designed to be left between these two mining fronts/blocks. The extraction of ore from the partial sill pillar level is expected to be 62.5% using production up-holes through 25 m of the 40 m thick sill pillar and is accounted for within the reserves. This methodology will allow partial mining of ore on the sill pillar level, while at the same time allowing the development of the lower mining block and establishing an early start to the mining of the upper mining block. Using this approach minimizes the impact on initial capital investment. The backfill was designed to have an adequate strength to allow for mining adjacent to filled stopes, thus eliminating the need for rib pillars.

There will be two shafts, which will minimize the amount of development through water-bearing horizons located in the first 200 m from surface. Both shafts will be excavated at the same time using conventional shaft sinking methods in conjunction with a freezing process through the first 200 m from the surface. The production shaft will facilitate main access and egress, material hoisting, fresh air intake, and material logistics. The ventilation shaft will serve as the mine exhaust system as well as a second means of mechanical egress. Mined ore will be transported from the stopes to the main production shaft hoisting system by underground LHD’s, trucks, ore passes, crusher and conveyor circuit.

A 3D mine design has been created representing the reserve areas. The underground mine design process results in mine plan resources of 36.656 Mt (diluted) with an average grade of 0.81% Nb₂O₅, 2.92% TiO₂, and 70.2 ppm Sc. This estimate is based on a mine design using elevated CoGs and applying the US$ 180/t NSR CoG to capture all potential mineral reserves within the design. These numbers are calculated by applying modifying factors which include a 95% mining ore recovery to the designed wireframes in addition to applying approximately 6% unplanned dilution.

Mineral Reserves were classified using the SK 1300 standards. Table 1-5 summarizes the underground Mineral Reserves. This Mineral Reserve Estimate is as of June 30, 2022.

Table 1- 5: Underground In Situ Mineral Reserves Estimate for Elk Creek, Effective Date June 30, 2022

Classifi-cation	Tonnage (kt)	Nb2O5 Grade (%)	Contained Nb2O5 (t)	Payable Nb (t)	TiO2 Grade (%)	Contained TiO2 (t)	Payable TiO2 (t)	Sc Grade (ppm)	Contained Sc (t)	Payable Sc2O3 (t)
Proven	-	-	-	-	-	-	-	-	-	-
Probable	36,656	0.81	297,278	170,409	2.92	1,071,182	431,793	70.2	2,573	3,677
Total	36,656	0.81	297,278	170,409	2,92	1,071,182	431,793	70,2	2,573	3,677

Source: Optimize Group, 2022. All figures are rounded to reflect the relative accuracy of the estimates. Totals may not sum due to rounding.

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● The Qualified Person for the Mineral Reserve estimate is Optimize Group Inc. The estimate has an effective date of June 30, 2022

● The Mineral Reserve is based on the mine design, mine plan, and cash-flow model utilizing an average cut-off grade of 0.679% Nb₂O₅ with an NSR of US$ 180/t.

● The estimate of Mineral Reserves may be materially affected by metal prices, environmental, permitting, legal, title, taxation, socio-political, marketing, infrastructure development, or other relevant issues.

● The economic assumptions used to define Mineral Reserve cut-off grade are as follows:

o Annual life of mine (LOM) production rate of ~7,450 tonnes of FeNb/annum during the years of full production.

o Initial elevated five-year production rate ~ 7,500 tonnes of FeNb/annum when full production is reached.

o Mining dilution of ~6% was applied to all stopes and development, based on 3% for the primary stopes, 9% for the secondary stopes, and 5% for ore development.

o Mining recoveries of 95% were applied in longhole stopes and 62.5% in sill pillar stopes.

Parameter	Value	Unit
Mining Cost	42.38	US$/t mined
Processing	106.70	US$/t mined
Water Management and Infrastructure	16.62	US$/t mined
Tailings Management	2.01	US$/t mined
Other Infrastructure	5.47	US$/t mined
General and Administrative	8.91	US$/t mined
Royalties/Annual Bond Premium	8.34	US$/t mined
Other Costs	6.29	US$/t mined
Total Cost	196.72	US$/t mined
Nb2O5 to Niobium conversion	69.60	%
Niobium Process Recovery	82.36	%
Niobium Price	39.60	US$/kg
TiO2 Process Recovery	40.31	%
TiO2 Price	0.88	US$/kg
Sc Process Recovery	93.14	%
Sc to Sc2O3 conversion	153.40	%
Sc Price	3,675.00	US$/kg

● Price assumptions are as follows: FeNb US$ 39.60/kg Nb, Sc₂O₃ US $3,675/kg, and TiO₂ US $0.88/kg. Price assumptions are based upon independent market analyses for each product as of June 30, 2022

● Price and cost assumptions are based on the pricing of products at the “mine-gate,” with no additional down-stream costs required. The assumed products are ferroniobium (metallic alloy shots consisting of 65%Nb and 35% Fe), a titanium dioxide product in powder form, and scandium trioxide in powder form.

● The Mineral Reserve has an average LOM NSR of US$ 563.06/tonne.

● Optimize Group has provided detailed estimates of the expected costs based on the knowledge of the style of mining (underground) and potential processing methods (by 3rd party Qualified Persons).

● Mineral reserve effective date is June 30, 2022. The financial model was run after the estimate of the NSR above, which reflects a total cost per tonne of US$ 196.72 versus US$ 189.91. This is not considered a material change.

● Price variances for commodities are based on independent market studies versus earlier projected pricing. The independent market studies do not have a negative effect on the reserve.

1.8 Environmental Studies, Permitting and Social or Community Impact

NioCorp has developed information and conducted a number of environmental studies related to baseline characterization for the Project. These include:

● Soils

● Climate/Meteorology/Air Quality

● Cultural and Archeological Resources

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● Vegetation

● Wildlife

● Threatened, Endangered, and Special Status Species

● Land Use

● Hydrogeology (Groundwater)

● Hydrology (Surface Water)

● Wetlands/Riparian Zones

● Geochemistry

The geochemistry and characterization/classification of the ore and waste materials (including the final process waste streams making up the bulk of the tailings mass and the crystallized reverse osmosis (RO) water treatment salts), directly influences the management of these materials given the presence of naturally occurring radioactive materials (NORMs) (i.e., uranium and thorium) and the potential for limited reaction to contact with water. These materials currently classify as non-hazardous based on regulatory testing. Site-wide management of non-contact and contact stormwater will be essential to the Project compliance. Given the presence of low levels of NORMs and this potential reactivity, NioCorp will take the conservative approach of placing this material in a double-lined containment facility from which any surface water runoff or seepage can be controlled and managed. It is not anticipated that any of the mine development or waste rock would remain exposed on the surface post closure.

Characterization of the various tailings materials has included both the Toxicity Characteristic Leaching Procedure (TCLP) and the Synthetic Precipitation Leach Procedure (SPLP), which are designed to determine the mobility of both organic and inorganic analytes present in the liquid, solid, and multiphasic wastes, and assist in the proper classification of waste materials. The most recent tailings material testing showed negligible mobility of regulated constituents (indicating a non-hazardous classification), although the pH of the TCLP/SPLP extracts remained high. While the calcined tailings are likely to produce heat when exposed to atmospheric moisture and precipitation (i.e., exothermic hydration), this reaction is not “violent” as defined under 40 CFR § 261.23(2) Characteristic of reactivity [for hazardous wastes] (adopted by the State of Nebraska under Title 128 - Nebraska Hazardous Waste Regulations). Given the limited quantities of ore available for testing, further characterization of these materials is recommended in order to establish representativeness of the deposit as a whole with respect to waste classification.

There are currently no known environmental issues that could materially impact NioCorp’s ability to extract the Mineral Resources or Mineral Reserves at Elk Creek. However, there are several key permitting risks and uncertainties that could affect the Project financing and schedule. These are outlined below.

Overburden developed during mine construction will be excavated, crushed and used as a construction material. Small quantities of waste rock will be temporarily stored on the surface (on a lined pad) prior to final disposal underground or within the lined tailings impoundment. The water leach residue tailings will be returned to the underground mine as backfill, along with a portion of the oxide tailings. The remaining oxide tailings and slag will be deposited within double-lined, surface storage facilities.

During the life of the mine there will be an inflow into the mine of up to 63 L/s (1,000 US gpm), which will require removal to the surface via an in-mine pumping system. During operations, clean

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water produced from the RO water treatment system will be used in the process plant, and the reject concentrate will be evaporated and crystallized for disposal as solid waste material.

Upon cessation of mining, the facilities will be closed in accordance with state requirements and best industry practice. Until such time that the final TSF closure cover can be constructed, and any residual water or seepage eliminated, the TSF contact water will require active management.

Engagement of local, state, and federal regulators has commenced. Initiation of the formal operational permitting program for the Project is dependent upon the completion of the mine plan and surface facilities being developed as part of this technical document, as well as additional characterization of the waste materials and potential worker exposures under the jurisdiction of the Nebraska Department of Health and Human Services (DHHS) and U.S. Department of Labor — Mine Safety and Health Administration (MSHA), both of whom will have primary oversight of worker safety and monitoring programs with respect to the presence of NORMs in the ore and waste rock. A comprehensive table of requisite permits and authorizations is presented in Section 17.2.

Stakeholder engagement has been undertaken in parallel with field operations in Nebraska and has included town hall discussions and individual meetings. Some early communications have occurred between NioCorp and Johnson, Pawnee, Nemaha and Richardson County representatives (including the county commissioners) as well as the Southeast Nebraska Development District.

Without specific hardrock mining regulations, there are limited obligatory requirements for reclamation and closure of mining properties in Nebraska. There are provisions, however, within the applicable regulatory framework that is likely to be applied to the Project during the permit and licensing processes, specifically those associated with the TSF and mineral processing facilities. This will include the provision of financial surety for proper closure and reclamation of the site. The estimated direct costs for closure and reclamation of the Project, plus financial assurance premiums, is US$ 50.2 million using a 2019 cost basis.

Overall, the Project appears to be sufficiently advanced to initiate the submission of the remaining formal operational permit applications which will govern the construction, operation, and closure of the mine.

At the time of writing, NioCorp had obtained two permits necessary to allow for the construction of the project: (1) a construction air permit from the state of Nebraska; and (2) a special use permit from Johnson County, Nebraska.

1.9 Capital Cost Estimate

The estimate has an intended accuracy of ± 15% and an overall contingency of 10%. The estimate is reported in Q1 2019 U.S. constant dollars.

Costs have for the most part been retained from the previous 2019 Feasibility Study. The capital cost estimate reflects a detailed bottom-up approach that is based on key engineering deliverables that define the project scope. This scope was described and quantified within material take-offs (MTO’s) in a series of line items. Capital costs are divided among the areas of underground mining, processing, infrastructure, water management, tailings management, mining indirects (indirect costs), mining and processing commissioning and contingency. Sustaining capital costs are related to underground mining fixed equipment and development, process plant, infrastructure maintenance, tailings management, mine closure and contingency.

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The mining, processing and infrastructure capital costs were developed based on a combination of vendor and contractor quotations, first principles buildup, allowances, and historical database costs. The estimates include labour, materials, equipment purchase and operation cost, rental equipment, supplies, freight, and energy. The costs developed include direct and indirect costs and included separate contingencies on both. Equipment purchase includes freight, an allowance for transporting underground, initial training and commissioning. The tailings and water management capital costs were based on contractor estimates for earthworks and liner installation. The estimates were developed from recent and relevant costs on other projects or developed from first principles.

Table 1-6 shows the breakout in LOM initial and sustaining capital estimates, which total US$ 1,609 million. An overall 9.79% contingency factor has been applied to the initial capital estimate, while a smaller 2.06% contingency was applied to the sustaining capital estimate. The pre-production period is defined from April 2022 to the end of construction in June 2025 plus a six-month ramp-up period through the end of December 2025. Commercial production is expected to be declared on January 1, 2026.

Table 1- 6: Capital Costs Summary (US$ 000’s)

Description	Initial	Sustaining	Total
Capitalized Preproduction Expenses	$77,053		$77,053
Site Preparation and Infrastructure	$40,569	$15,007	$55,576
Processing Plant	$367,439	$96,448	$463,886
Water Management & Treatment	$73,756	$23,613	$97,369
Mining Infrastructure	$256,981	$198,482	$455,463
Tailings Management	$21,423	$78,855	$100,277
Site Wide Indirects	$7,368		$7,368
Processing Indirects	$96,028		$96,028
Mining Indirects	$41,130		$41,130
Process Commissioning	$13,350		$13,350
Mining Commissioning	$1,578		$1,578
Owner’s Costs	$33,619		$33,619
Mine Water Management Indirects	$8,520		$8,520
Closure and Reclamation		$44,267	$44,267
Contingency	$101,730	$9,385	$111,116
Total Capital Costs	$1,140,544	$466,058	$1,606,601
Preproduction Revenue Credit	($256,910)		($256,910)
Net Project Total	$883,634	$466,058	$1,349,692

Source: NioCorp, 2022

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1.10 Operating Cost Estimate

Operating cost estimates were developed to show monthly and annual costs for production. All unit costs are expressed as US$/tonne processed and are based on Q1 2019 US$. The estimate has an intended accuracy of ± 15% and an overall contingency of 10%.

Operating cost metrics in the technical economic model are reported on a LOM (life of mine) basis meaning that all of these unit rates are stated on a LOM basis where the costs are estimated from the beginning of construction to the end of mine life. LOM operating costs include the pre-production and first/last years of production.

The total LOM operating cost unit rate of US$ 195.94/t processed is summarized in Table 1-7.

Table 1- 7: LOM Operating Cost Unit Rate Summary

Description	LOM US$/t ore
Mining Cost	42.38
Process Cost	106.70
Water Management Cost	16.62
Tailings Management Cost	2.01
Other Infrastructure	5.47
Site G&A Cost	8.20
Other Expenses	6.22
Subtotal	187.59
Royalties/Annual Bond Premium	8.35
Total LOM Operating Costs	195.94

Source: NioCorp, 2022

1.11 QP Conclusions and Recommendations

Under the assumptions presented in this TRS, the Project shows positive economics. The QPs note the interpretations and conclusions presented in Section 22, in their respective areas of expertise, based on the review of data available for this TRS.

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2. INTRODUCTION

2.1 Terms of Reference and Purpose of the Technical Report Summary

This Technical Report Summary (TRS or Technical Report) conforms to the United States Securities and Exchange Commission’s (SEC) Modernized Property Disclosure Requirements for Mining Registrants as described in Subpart 1300 of Regulation S-K, Disclosure by Registrants Engaged in Mining Operations (S-K 1300), and Item 601(b)(96) of Regulation S-K, Technical Report Summary.

The TRS was prepared by the firms and individuals listed in Section 1 for the Elk Creek Project (Elk Creek or the Project) located in southeast Nebraska. NioCorp (the Company) has determined that these firms and/or individuals meet the qualifications specified under the definition of qualified person in § 229.1300. References to the Qualified Person or QP in this report are references to these firms and individuals. The firms and individuals who are QPs are consultants and engineers that are independent of the Company, with the exception of Scott Honan, who serves as the Company’s Chief Operating Officer.

This TRS presents Mineral Resource and Mineral Reserve Estimates for the Elk Creek Project.

2.2 Sources of Information

The parties responsible for generating this TRS are Dahrouge Geological Consulting USA Ltd. (Dahrouge), Understood Mineral Resources Ltd. (Understood), Optimize Group (Optimize), Tetra Tech, Adrian Brown Consultants Inc. (ABC), Metallurgy Concept Solutions (MCS), Magemi Mining Inc. (Magemi), L3 Process Development (L3), A2GC, Scott Honan, M.Sc, SME-RM, NioCorp, Everett Bird, P.E., Cementation, Matt Hales, P.E., Cementation, Mahmood Khwaja, P.E., CDM Smith, Martin Lepage, P.Eng, Ing., Cementation and Wynand Marx, M.Eng., BBE Consulting.

NioCorp provided the following information to the QPs to support the preparation of this report:

● Information on land ownership and land agreements in the project area

● All assay, water quality, environmental, metallurgical, geochemical and geotechnical data analyzed by external labs

● Information on permitting requirements for the project and the status of the Company’s permitting efforts

● Information related to NioCorp’s relationships with the local community and community groups

● Market reports and market data related to niobium, scandium, titanium and the rare earth elements

● The economic model and associated calculations that define the prospective economic performance of the project.

External sources of information used to prepare the TRS are listed in Section 25.

2.3 Details of Inspection

A summary of site visit inspections by the Consultants is provided in Table 2-1.

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Table 2-1: Site Visit Participants

Personnel	Company	Expertise	Date(s) of Visit	Details of Inspection
Matt Batty	Understood Mineral resources	Geology, Resources	April 27, 2022	Review of drill core, review and verification of the geological setting / environment, logging, sampling, analytical, QA/QC, site facilities, drill collar locations
Trevor Mills	Dahrouge	Elk Creek geology expert	April 27, 2022	Review of drill core, review and verification of the geological setting / environment, logging, sampling, analytical, QA/QC, site facilities, drill collar locations
Bladen Allen	Optimize Group	Reserves/Mining	April 27, 2022	Review of drill core, review, verification of the geological setting / environment, logging, sampling, analytical, QA/QC, site facilities, drill collar locations
Eric Larochelle	SMH	Hydrometallurgy and Process Engineering	October 22, 2014	Visited the greenfield site
Brian Osborn	Olsson	Environmental and Permitting	October 6, 2021	Site visit and meeting with local residents and investors
Guy Cinq-Mars	Tetra Tech	Mech Design	Nov 21,22, 2016	Visited the greenfield site and surrounding region.
Sylvain Turcotte	Tetra Tech	Mech Design	Nov 21, 22, 2016	Visited the greenfield site and surrounding region.
Pierre Mathieu	Tetra Tech	Project Management	Nov 21, 22, 2016	Visited the greenfield site and surrounding region.
John Gorham	Dahrouge	Geology	Site visits 2010 (2)	Visited the greenfield site, reviewed drill core, local ground conditions, local geology, geotechnical, surrounding region.
Scott Honan	NioCorp	Environmental, metallurgy, tailings	April 27, 2022	Review of drill core, review and verification of the geological setting / environment, logging, sampling, analytical, QA/QC, site facilities, drill collar locations

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2.4 History

Exploration activities at the Project prior to NioCorp ownership were conducted by the following companies: University of Nebraska – Lincoln, Nebraska Conservation and Survey Division, United States Geological Survey (USGS), Cominco American Incorporated (Cominco American), Molybdenum Corporation of America, later Molycorp Inc. (Molycorp) and Quantum Rare Earth Developments Corp. (Quantum). These activities consisted of airborne magnetic and gravity surveys, geochemical sampling, Reverse Circulation (RC) drilling, core drilling and Mineral Resource Estimates.

Since 2014, NioCorp has completed metallurgical testing, core drilling, mineral resource and mineral reserve estimates.

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3. PROPERTY DESCRIPTION

3.1 Property Location

The Project is located in southeast Nebraska, USA. The Property is situated, as shown in Figure 3-1, and is located as follows:

● Within United States Geological Survey (USGS) Tecumseh Quadrangle Nebraska SE (7.5-minute series) mapsheet in Sections 1-6, 9-11. Township 3N. Range 11 and Sections 19-23, 25-36. Township 4N, Range 11.

● At approximately 40°16’ north and 96°11’ west in the State of Nebraska, in the central USA.

● On the border of Johnson and Pawnee counties.

● Approximately 75 km southeast of Lincoln, Nebraska, the state capital of Nebraska.

● Approximately 110 km south of Omaha, Nebraska.

● Approximately 183 km northwest of Kansas City, Kansas and Missouri.

● Approximately 5 km southwest of the town of Elk Creek, Nebraska. the closest municipality to the Deposit.

● Approximately 11 km south of Tecumseh, Nebraska

● Approximately 53 km west of the state border with Missouri.

● Approximately 55 km southwest of the state border with Iowa.

● Approximately 29 km north of the state border with Kansas.

● Approximately 53 km west of the Missouri River, which forms the state border with Missouri and Iowa.

● Approximately 5 km southeast of the Nemaha River, a tributary of the Missouri River.

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Source: Nordmin, 2019

Figure 3-1: Project Location Map 

3.2 Property Description and Land Tenure

The Project is a niobium, scandium, titanium and rare earth-bearing carbonatite deposit located in Johnson County, southeast Nebraska.

The Property consists of one 91.5 ha (226 acre) parcel of land owned by the Company along with 8 option-to purchase agreements covering approximately 565 ha. Option agreements are between NioCorp’s subsidiary Elk Creek Resources Corp. (ECRC) and the individual landowners (see Figure 3-2). The parcel owned by the Company contains the majority of the Mineral Resources and Mineral Reserves associated with the project. ECRC is a Nebraska-based wholly owned subsidiary of NioCorp. NioCorp retains 100% of the mineral rights to the Project and is the operator. The option agreements are in the form of pre-paid Exploration Lease Agreements (ELA), with an Option to Purchase (OTP) the mineral rights and/or the surface rights at any time during the term of the agreement. The individual landowners have title to the surface and subsurface rights, and the agreements are primarily concerned with only the mineral and surface interest of each property. The agreements convey to the Company adequate surface rights to access the land and to complete mineral exploration work. The option agreements that the Company currently holds include all the Indicated and Inferred resources and Probable reserves described in this TRS. Active OTP agreements are listed in Table 3-1.

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Source: NioCorp, 2022 *Infilled blue polygons are indicating option agreements for minerals only. For the 32.37 hectare parcel north of Beethe008, NioCorp has an option to purchase the surface rights, and negotiations to secure the mineral rights are underway.

Figure 3-2: Land Tenure Map*

Table 3-1: Active Option to Purchase Agreements Covering the Project

Agreement Identifier	Hectares	Acres	Agreement Expiry
Beethe007	66.27	163.75	20-Jan-26
Heidemann005	79.55	196.57	16-Mar-25
Nielsen001	100.91	249.32	25-Jun-25
Nielsen002	11.91	29.43	14-Jun-25
Woltemath80S	32.37	80.00	4-Dec-24
Woltemath002	152.49	376.81	4-Dec-24
Woltemath003J	89.03	220.00	25-Mar-25
Shuey001	32.37	80.00	27-May-40

Source: NioCorp, 2022

The current (2022) Mineral Resource and the Mineral Reserve is wholly contained within parcels Woltemath003J and ECRC on Figure 3-2.

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3.2.1 Nature and Extent of Issuer’s Interest

As part of the exploration OTP agreements the Company has secured surface rights, which allow for access to the land for drilling activities and associated mineral exploration and project development work.

The agreements that involve mineral rights include a 2% Net Smelter Return (NSR) royalty attached to the OTP. The agreements grant the operator an exclusive right to explore and evaluate the property with an option to purchase the mineral rights, the surface rights or a combination of the mineral and surface rights at any time during the term of the agreement.

3.3 Royalties, Agreements and Encumbrances

The leases covering the Project are 100% owned by NioCorp and, apart from a 2% NSR royalty attached with the OTPs that include the mineral rights, have no other outstanding royalties, agreements, or encumbrances (see Figure 3-3).

Source: NioCorp, 2022

Figure 3-3: Net Smelter Return (NSR) Map

3.3.1 Required Permits and Status

The exploration work conducted to date on the Project has been completed under Exploration Permit NE0211001 issued by the Nebraska Department of Environment and Energy (NDEE). The permit provides the Company with the right to have ten open boreholes active at the Project at any given time. The Company has received a Construction Air Permit from the state of Nebraska along with a Special Use Permit from Johnson County, Nebraska which will allow it to begin construction.

The proposed Project will be required to obtain a series of permits for operations from federal, state and local agencies. The majority of these permits are ministerial in nature and present

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minimal risk to the applicant, and typically involve the completion of an application and the payment of a nominal fee. Three permits from the state of Nebraska are discretionary in nature, where an application and fee are provided to the state and the state must make a decision as to whether or not the permit will be granted. While the risk involved in such permits is low, such discretionary permits require more processing time by the state and do require the state agency to make a decision in favor of issuance of the permit. These three permits include the following:

● Solid Waste Permit

● Air Operating Permit

● Radioactive Materials License

The cost and schedule for obtaining both the discretionary and ministerial permits is included in the overall execution plan for the project. Additional details on the project’s permitting requirements can be found in Section 17 of the TRS.

NioCorp has not received any notices of violations or fines with respect to its permits to this point in the Project.

3.4 Other Significant Factors and Risks

There are no known other significant factors or risks which could have a material impact on the ability to affect access, titles, or the right to perform exploration work on the property.

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4. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

4.1 Physiography

The local topography of eastern Nebraska is relatively low-relief with shallow rolling hills intersected by shallow river valleys. Elevation varies from about 325 to 390 metres above mean sea level (masl). Bedrock outcrop exposure is nonexistent in the Project area.

Much of the Project area is used for cultivation of corn and soybeans, along with use as grazing land. Native vegetation typical of eastern Nebraska is upland tall-grass, prairie, and upland deciduous forests.

4.2 Accessibility and Transportation to the Property

The Project is easily accessible year-round as it is situated approximately 75 km (47 miles) southeast of Lincoln (State Capital), Nebraska and approximately 110 km (68 miles) south of Omaha, Nebraska. Access to the site can be achieved via road or from one of the regional airports. There are several regular flights to both Lincoln and Omaha; however, the Project is most easily accessible from Lincoln (see Figure 4-1).

 

Source: Nordmin, 2019

Figure 4-1: Project Location Showing Main Access Routes

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From Lincoln Municipal Airport, the Property is accessed via paved roads on the main network and a secondary network of gravel roads by following:

● Interstate Highway 80 south for approximately 3.5 km to the Beatrice exit;

● then join Highway 77 south for approximately 41 km;

● then join Highway 41 east for approximately 47 km; and,

● then join Highway 50 south for approximately 16 km through Tecumseh to the approximate center of the Elk Creek Deposit.

The drive from the Lincoln Municipal Airport to the property is typically 1 hour and 15 minutes, and from Omaha’s Eppley Airport the drive is approximately 1 hour and 45 minutes.

4.3 Climate and Length of Operating Season

Southeast Nebraska is situated in a Humid Continental Climate (Dfa) on the Köppen climate classification system. In eastern Nebraska, this climate is generally characterized by hot, humid summers and cold winters. Average winter temperatures vary between -10.4°C to 1.6°C. Average summer temperatures vary between 18°C to 32°C. Exploration and mining related activities may be conducted all year round.

Average monthly precipitation (rain and snowfall) varies between 22 and 127 mm. Average yearly precipitation is between 800 and 850 mm with an average yearly snowfall of approximately 72 cm (see Table 4-1). Nebraska is located within an area known for tornados which run through the central U.S. where thunderstorms are common in the spring and summer months. Tornadoes primarily occur during the spring and summer and may occur into the autumn months.

Table 4-1: Summary of the Project Precipitation Data ^{(4) (5)}

Station	Mean Monthly Precipitation	Mean Monthly Pan Evaporation	Mean Monthly Lake Evaporation (5)	Annual Evapotranspiration
	Tecumseh Station (1) (mm)	Sabetha Lake Station (2) (mm)	Sabetha Lake Station (2) (mm)	Rainwater Basin Station (3)
January	21		-	30
February	28	-	-	32
March	49	-	-	66
April	72	131	98	84
May	111	167	126	98
June	117	186	139	98
July	99	210	158	102
August	97	190	142	87
September	89	138	103	86
October	58	103	77	81
November	39	57	43	58
December	26	-	-	29

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Annual	805	1,182	887	851
7 Year Wet-Cycle Total	6,662
7 Year Dry-Cycle Total	4,318

(1) Tecumseh station data (WRCC, DRI) is considered the most representative based on elevation and proximity to the Project site.

(2) Data from Southwest Climate and Environmental Information Collaborative (WRCC, DRI); Sabetha Lake station data is considered the most representative based on elevation and proximity to the Project site.

(3) RAWS Network (DRI), ASCE Standardized Reference ET Calculations.

(4) 5-year average from 2009 through 2013.

(5) Based on Lake Evaporation as 75% of Pan Evaporation.

4.4 Infrastructure

4.4.1 Personnel and Supplies

Technical and trades personnel can be sourced from local colleges and universities. An underground-experienced mining-related workforce can be found around Weeping Water, Nebraska as well as in neighbouring states such as Salt Lake City, Utah, South Dakota and Denver, Colorado (eight hours drive west of the Project).

The Project is located immediately adjacent to Nebraska state highway 50, providing direct road access to surrounding population centers including the cities of Lincoln and Omaha, Nebraska. Both Omaha and Lincoln are able to supply the project with personnel and supplies for construction and operations if such cannot be supplied by the communities in the immediate vicinity of the project site.

4.4.2 Electrical Power

The local electric utility is the Omaha Public Power District (“OPPD”) and that entity maintains electric power distribution infrastructure in the project area and currently serves the project site.

4.4.2.1   Electrical Power Line & Substation

The local power utility (Omaha Public Power District) will provide power to the site. This will require approximately 29 km (18 miles) of new transmission line be installed by the utility to provide power to the Project site main sub-station to meet the required power demand, which is estimated at a peak of 30 MW. The local power utility will also design and install the main substation that will be owned and maintained by the utility. This infrastructure will be paid back through rate charges on electrical usage.

4.4.2.2   Electrical Power Distribution - Plant and Facilities

The main substation will feed the site distribution substation with 44 kV. A 44 kV pole line will be constructed on the Project site to supply main power throughout the site and to the mine sub yard. In addition, this substation will include two 20/25 MVA transformers to provide 13.8 kV for distribution through the above ground facilities with approximately 1,100 m (3,610 ft) of power cables in vaults, and approximately 1,600 m (5,250 ft) of overhead lines.

4.4.2.3   Electrical Power Distribution - Underground

The underground electrical distribution will be fed from both the production and ventilation shafts, at 13.8 kV. Duplex fused disconnect switches will be present at several levels to allow

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power to be selected from either 13.8 kV feeder, providing redundancy. Power for utilization is accomplished through portable mine power centers, located at each production level. The duplex fused switches are not on every level but are distributed to adjacent levels through medium voltage junction boxes and boreholes.

4.4.2.4   Emergency Power Generation

Independent emergency power generation at the hoist house and ventilation shaft switchgear will be provided for back-up generation for surface infrastructure. Ventilation and hoisting are all powered from the surface, and thus, no emergency power is fed to the underground electrical distribution. Emergency power generation for the hoisting and ventilation systems will be supplied with from two diesel-powered generators, one at the hoist house and one at the ventilation shaft.

4.4.3   Water

Water is available from local surface water bodies and groundwater wells and has been adequate to support operations at the site to this point. Additional sources of water are available from the Tecumseh Board of Public Works and Johnson County Rural Water, both of which maintain water supply infrastructure in the immediate project vicinity.

4.4.3.1   Process Water

Process water will be produced at the Water Treatment Plant. Plant process water will be required in the Hydromet Plant, Acid Plant, HCI Regeneration Plant, Paste Backfill Plant and the Pyromet Plant. Additional treated water will be required for both the Mine, as well as for site potable and fire water systems.

The vast majority of process water will be required in the Hydromet Plant. The Paste Backfill Plant will utilize RO permeate for backfill, as will the mine for underground operations. The remaining plants identified will require small quantities of make-up water primarily for cooling and chilling purposes.

Mineral Processing Plant

The water requirements for the Mineral Processing plant are minimal. Plant water will be available for use during the cleanup.

Hydrometallurgical Plant

The water requirement for the Hydrometallurgical Plant is to provide dilution, make up and wash water to various sections of the plant. Table 4-2 provides a summary of the water requirement for the Hydrometallurgical Plant.

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Table 4-2: Summary of Hydrometallurgical Process Water Requirement

HCI Leach	542.5 t/d
Water Leach	2,914 t/d
Nb Precipitation	1,713.7 t/d
Nb Caustic Leach	63.5 t/d
Sc Precipitation and Refining	24.0 t/d
Ti Precipitation	20.9 t/d
Total	5,272.9 t/d
gpm	967

Source: Tetra Tech, 2017

Pyrometallurgical Plant

The water requirements (Table 4-3) for the Pyrometallurgical Plant provide make-up water to supply the FeNb Furnace cooling systems and the FeNb Furnace Pelletizer basin for cooling and pelletizing the FeNb product.

Table 4-3: Pyrometallurgical Water Requirements

Water Item	Units	Value
Furnace Cooling Water Flowrate	m3/h	64.7
Cooling Water Flowrate Addition for Pelletizing	m3/h	15.1
Water Volume Required	m3/tap	68.5
Steam Produced	%	20
Steam Flowrate	m3/min	0.05
Make-up water Required	m3/min	0.05

Source: Tetra Tech, 2017

4.4.3.2   Fire Water

The firewater system will be comprised of two 225,000 gal insulated fire water tanks and two independent fire water pumps capable of delivering 2,000 gpm for a minimum period of four hours. The primary pump will be electrically driven while the backup pump will be diesel powered. A fire water distribution system will be installed throughout the site. Dry and wet sprinkler systems, hydrants, hose reels and fire extinguishers will be utilized per the design.

All infrastructure facilities on the surface, except for the gate house, will include fire suppression systems. Process building fire suppression systems will include wet sprinklers in all office spaces and control rooms. Dry sprinkler systems will be utilized in the hydrometallurgical buildings within specified high hazard areas. The remaining open process/factory areas of these two process facilities, as well as the open areas of the mineral processing building, will utilize fire hose protection from outside hydrants, as well as interior located fire hose reels.

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4.4.3.3   Potable Water

Potable water will be supplied from three possible available sources at an operational flow rate of 3500 gpm to dedicated potable water tankage. These possible sources with their expected flow rates include; a supply line furnished by the Tecumseh Board of Public Works (2,000 gpm), a well and supply line from the Landowner 1 property (500 gpm), and two (2) wells and a supply line from the Landowner 2 property (1,500 gpm). Potable water will be distributed to all site facilities via a dedicated pumping system at 50 psig pressure. The nominal flow rate will be 100 gpm for the entire facility, with a peak flow rate of 750 gpm during shower usage.

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5. HISTORY

5.1 Exploration History

5.1.1 USGS, 1964

Between November 1963 and January 1964, the USGS flew three airborne magnetic surveys southeast Nebraska. A total of 6,590 line-km was flown (836, 209, and 5,544 line-km, respectively) along an east-west direction at a flight line spacing of 3.2 km (2 miles) and an altitude of 305 m above ground (USGS website: OFR 99-0557). Figure 5-1 shows the area covered by the airborne surveys.

 

Source: Tetra Tech, 2012 – Modified from USGS, 1964

Figure 5-1: 1964 USGS Aeromagnetic Survey Area Showing Surveys 526A, 526B and 530 Respectively

This wide spacing of the flight lines illustrates only regional features and does not locate local anomalies (i.e., Elk Creek Nb-REE anomaly). Details of the aeromagnetic survey may be found in USGS Publication 73-297, which was unavailable at the time of writing. Results of the aeromagnetic survey are shown in Figure 5-2.

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Source: Tetra Tech, 2012

Figure 5-2: 1964 USGS Aeromagnetic Results (Merged 526A, 526B, and 530 Surveys)

5.1.2 Discovery, 1970-1971

Further investigation of the Project was not completed until 1970 when the Elk Creek gravity anomaly was initially identified during a reconnaissance gravity geophysical survey of southeast Nebraska by the Conservation and Survey Division (CSD) of the University of Nebraska-Lincoln (UNL). During the same period the UNL geology department (operating independently), was mapping the magnetic expression of the Nemaha Arch and the Humboldt Fault.

A comparison of the two geophysical survey results showed a positive magnetic anomaly that was coincident with a positive gravity anomaly over the area now defined as the Elk Creek gravity anomaly (Anzman, 1976). The geophysical gravity survey outlined a near-circular anomaly, along with a concurrent magnetic anomaly, approximately 7 km in diameter. Analysis of the geophysical data provided a model of a cylindrical mass of indefinite length with a radius of 1,676 m (5,500 ft) (Carlson et al., 2005). Figure 5-3 and Figure 5-4 illustrate the results of the two surveys.

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In 1971, the Nebraska Geological Survey (NGS) commissioned a test drill hole 2-B-71 to determine the source of the near circular gravity anomaly. With some support from the United States Bureau of Mines (USBM), the test hole was deepened. The test hole 2-B-71, later renamed NN-1 by Molycorp, encountered 191 m (628 ft) of marine sediments, followed by a carbonate-rich rock (carbonatite) to the end of the hole at 290 m (952 ft) (Brookins et al., 1975) in what is now referred to as the Elk Creek Carbonatite.

 

Source: Tetra Tech, 2012

Figure 5-3: Comparison of the 1970 Magnetic and Gravity Geophysical Surveys

 

Source: Tetra Tech, 2012

Figure 5-4: Cross-section A-A’ of the 1970 Gravity and Magnetic Geophysical Surveys

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5.1.3 Cominco American, 1974

The earliest known reference to Cominco American operating within the Elk Creek gravity anomaly area is from 1974. It is unclear precisely when Cominco American first acquired the mineral rights in the Elk Creek anomaly area. It is believed that between 1971 and 1973 both Cominco American and Molycorp held mineral rights over selected portions of the Elk Creek gravity anomaly.

In 1974, Cominco American completed five drill holes (CA-1 to CA-5) within the Elk Creek gravity anomaly. Details of the Cominco American drill holes and exploration activities within the property were not available. The information on drilling activities stated here was taken from the Molycorp database. Dahrouge has not reviewed or included any information from Cominco American as part of the current study.

5.1.4 Molycorp, 1973-1986

The earliest known reference to Molycorp operating within the Elk Creek gravity anomaly area is from 1973. It is unclear precisely when Molycorp first acquired the mineral rights in the Elk Creek anomaly area. Molycorp completed a number of phases of exploration on the Project during this period including more detailed geophysical surveys, regional drilling (mineralization limits) and focused drilling on the Project area. The exploration program focused on understanding the potential for rare earth elements of economic significance at the Project, with results showing a niobium anomaly at Elk Creek.

Between 1986 and 2011, no further exploration was recorded on the property.

5.1.5 Geophysical Surveys

In 1973, a detailed aeromagnetic survey was flown by Olympus Aerial Surveys Inc. (Olympic Aerial Surveys), of Salt Lake City, Utah, USA, for Molycorp, with the aim of locating drill sites. Flight-lines within the Elk Creek anomaly area were spaced at 200 m, and outside the anomaly at 400 m. A total of 50,764 ha was covered by 2,090 line-km (Anzman, 1976). The altitude of the survey was not stated in Anzman (1976).

In 1980, an extensive regional geophysical program was made in southeastern Nebraska including the Elk Creek anomaly. The program consisted of 6,437 line-km of aeromagnetics and approximately 4,000 gravity station readings. The aeromagnetic survey was contracted by Olympus Aerial Surveys.

The gravity geophysical survey was conducted by the CSD-UNL, which undertook approximately a quarter of the station readings, and by Molycorp’s in-house Geophysical Services Group, which undertook the remaining three-quarters of the gravity station readings.

5.1.6 Drilling

Between 1973 and 1974, Molycorp completed six drill holes: EC-1 to EC-4, targeting the Elk Creek anomaly and two other holes outside the Elk Creek anomaly area (Anzman, 1976). Drill holes were typically carried out by RC drilling through the overlying sedimentary rocks and diamond drilling through the Ordovician-Cambrian basement rocks.

Molycorp continued their drill program from 1977 and, in May 1978, Molycorp made its discovery of the current Project with drill hole EC-11. EC-11 is located on Section 33, Township 4N, and Range 11. The carbonatite hosting the Project was intersected at a vertical depth of 203.61 m (668 ft).

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Molycorp continued its drilling program through to 1984, which mainly centred on the Project within a radius of roughly 2 km. By 1984, Molycorp had completed 57 drill holes within the Elk Creek gravity anomaly area, which included 25 drill holes over the Project area.

From 1984 to 1986, drilling was focused on the Elk Creek gravity anomaly area. The anomaly area is roughly 7 km in diameter and drilling was conducted on a grid pattern of approximately 610 m by 610 m (roughly 2,000 ft by 2,000 ft.) with some closely-spaced drill holes in selected areas.

By 1986, a total of 106 drill holes were completed for a total of approximately 46,797 m (153,532 ft). The deepest hole reached a depth of 1,038 m (3,406 ft) and bottomed in carbonatite.

5.1.7 Molycorp Data Verification, 1973-1986

Verification work on the historical database has been completed by Dahrouge, who were contracted by Quantum to compile and verify the historical database between 2010 and 2011. Work included data capture from historical drilling logs, verification drilling and re-analysis of historical samples.

The following excerpt was taken from the Technical Report on the Elk Creek Property, 2010 (McCallum and Cathro, 2010).

“In some of the analytical log sheets available to the Authors, it appears that Molycorp analyzed niobium through their exploration division laboratory at Louviers, Colorado. They also analyzed the same interval at another, unspecified, commercial laboratory. It is unclear to the Authors what material the duplicate analyses were derived from (coarse reject duplicate, pulp duplicate, or % core duplicate).

Molycorp utilized the commercial laboratory, Skyline Labs Inc., of Wheat Ridge, Colorado between 1980 and 1986, with analysis by ICP spectrographic methods and unknown preparation methods. According to analytical reports and certificates available at UNL, values of lanthanum, cerium, neodymium, barium, sodium, thorium, lead, thorium, uranium, potassium, titanium, zinc, vanadium, niobium, phosphorous, beryllium, zircon, strontium, lithium, yttrium, silver, chromium, copper, iron, manganese, nickel and cobalt were tested. The intervals tested are comprised of commonly 100 ft intervals, presumably composited from the pulverized material of the 10 ft intervals.

In the “Niobium Analytical Standardization” report, dated June 1983, by Sisneros and Yernberg, it was noted that the routine XRF analysis performed by Molycorp’s exploration division laboratory at Louviers generated niobium values that were higher than other analytical techniques. This difference in niobium values was concluded not to be a product of preparation techniques, but a result of the standardization errors in the XRF analytical technique. A set of fifteen composites was prepared from Elk Creek drill core samples and analyzed with varying methods including XRF, ICP emission spectrometry and DC plasma emission spectrometry at ten laboratories. It was concluded that the difference was caused by high barium and iron within the matrix of the sample, with the largest deviations found in the coarse-grained material. The deviation of Molycorp’s routine analytical method compared to the recommended value ranges from 20% to just below 50% (except for one sample deviating 1%). The recommended value was based on a statistical analysis of the round-robin results.

The correction for the effect of barium and iron on the given Louviers niobium value was calibrated with the XRF instrument at Molycorp’s Louviers, Colorado exploration laboratory, and many of the previously analyzed samples were re-tested with the new calibration. The samples that have received the Ba+Fe correction have been noted on the historic Molycorp analytical logs;

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however, in the later series of holes, it is not identified on the assay log. It is expected that all holes drilled after 1983 were analyzed with the newer calibration.”

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6. GEOLOGICAL SETTING, MINERALISATION AND DEPOSIT

6.1 Regional Geology

The Nebraska Precambrian basement is comprised mainly of granite, diorite, basalt, anorthosite, gneiss, schist and clastic sediments. A series of island arcs sutured onto the Archean continent created the basic framework of the area. This suture left a north-trending intervening boundary zone ancestral to the Nemaha Uplift, providing a pre-existing tectonic framework which controlled the trend of the later Midcontinent Rift System (1.0 to 1.2 Ga) (Carlson & Treves, 2005). The Elk Creek Carbonatite is located at the northeast extremity of the Nemaha Uplift.

The Midcontinent Rift System, or Keweenawan Rift, comprises mafic igneous rocks and forms a belt over 2,000 km long and 55 km wide that is exposed at the surface in the Lake Superior Region and extends southwards through the states of Michigan, Wisconsin, Minnesota, Iowa, Nebraska and into Kansas (Carlson, 1992). Both basalt and associated red clastic sedimentary rocks are found in the Precambrian basement of southeastern Nebraska. These rocks are very similar to those found in the Lake Superior region and are thus considered to be a product of the Keweenawan rifting (Burchett and Reed, 1967; Treves et al., 1983). Figure 6-1 illustrates the major rock types of the Midcontinental Rift system.

 

Source: Modified from Palacas et al., 1990

Figure 6-1: Regional Geology Map

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The Nemaha Uplift (300 Ma) extends southward as a narrow belt from around Omaha, Nebraska across Kansas to around Oklahoma City, along the midcontinent rift system (King, 1969) (see Figures 6-1 and 6-2). Along the northern and eastern margins are complex fault zones and steeply dipping units. Regional north-northeast to northeast striking faults are locally transected by northwest trending ones, including the Central Plains mega-shear (Central Missouri Fault) to the north and the Oklahoma mega shear to the south (McBee, 2003). The Elk Creek Carbonatite body intruded near to the axis of the Nemaha uplift and has similar age dates to a cluster of carbonatites north of Lake Superior that are in the range of 560 to 580 Ma. (Woolley, 1989; Erdosh, 1979). Temporally, the carbonatite occurs near the boundary between the Penokean Orogen (approximately 1,840 Ma) and the Dawes terrane (1,780 Ma) of the Central Plains Orogen (Carlson and Treves, 2005).

Figure 6-2 shows a merged airborne magnetic anomaly map of Nebraska, Kansas, and Oklahoma states (USGS, 2004) showing the Midcontinent Rift and Nemaha Uplift systems.

 

Source: Modified from USGS 2004, showing the Midcontinental Rift and Nemaha Uplift.

Figure 6-2: Merged Aeromagnetic Anomaly Map of Nebraska, Kansas and Oklahoma States

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Regional geophysical data and drilling have confirmed the presence of kimberlitic intrusive bodies in northern Kansas to the southwest of the Elk Creek Carbonatite. These kimberlites were emplaced along the rift system during the Cretaceous time (Berendsen and Weis, 2001).

The approximately 200 m of Paleozoic rocks overlying the carbonatite region are dominantly essentially flat-lying Pennsylvanian marine strata consisting of carbonates, sandstones, and shales. The eastern portion of Nebraska was glaciated several times throughout the early Pleistocene (Wayne, 1981), resulting in the deposition of up to 50 m of unconsolidated till.

6.2 Property Geology

The property includes the carbonatite that has intruded older Precambrian granitic and low- to medium-grade metamorphic basement rocks. The carbonatite and Precambrian rocks are unconformably overlain by approximately 200 m of Paleozoic marine sedimentary rocks of Pennsylvanian age ranging from ca. 299 to 318 Ma (Figure 6-3).

As a result of this thick cover, there is no surface outcrop within the Project area of the carbonatite, which was identified and targeted through magnetic surveys and confirmed through subsequent drilling. The available magnetic data indicates dominant northeast, west-northwest striking lineaments and secondary northwest and north-oriented features that mimic the position of regional faults parallel and/or perpendicular to the Nemaha Uplift (Figure 6-2).

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Source: Modified from KGS O-F Report 91-52

Figure 6-3: Generalized Stratigraphy of Elk Creek Area (see Table 6-2 for details)

6.3 Elk Creek Carbonatite

The Elk Creek Carbonatite is an elliptical magmatic body with a northwest-trending long axis perpendicular to the strike of the 1.1 Ga Midcontinent Rift System (Figure 7-1, Figure 6-4), near the northern part of the Nemaha uplift (Burchett, 1982; Carlson, 1992). The definitive confirmation of carbonatite was completed using Rare Earth Element (REE), P₂O₅ and Sr⁸⁷/Sr⁸⁶ isotope analysis (Brookins et al., 1975). The carbonatite has also been compared to the Iron Hill carbonatite stock in Gunnison County, Colorado, based on similar mineralogy (Xu, 1996).

The carbonatite consists predominantly of dolomite, calcite and ankerite, with lesser chlorite, barite, phlogopite, pyrochlore, serpentine, fluorite, sulphides and quartz (Xu, 1996). The stratigraphic reconstruction based on drill core observation in the area suggests that the carbonatite is unconformably overlain by approximately 200 m of essentially flat-lying Palaeozoic marine sedimentary rocks, including carbonates, sandstones, and shales of Pennsylvanian age (ca. 299 to 318 Ma).

Current studies suggest that the carbonatite was emplaced ca. 500 Ma (Xu, 1996) in response to stress along the Nemaha Uplift boundary predating deposition of the Pennsylvanian sedimentary sequence (ca. 299 to 318 Ma). Observations on drill core from the Project site show that the contact between the carbonatite body and the Pennsylvanian sediments is a sheared, but oxidized

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contact suggesting that the carbonatite is intrusive in the Pennsylvanian sequence (see Figure 6-3 and Figure 6-4). Furthermore, both rock types appear to have been affected by at least one main brittle-ductile deformation event resulting in the formation of fault structures. Microstructures including sub-vertical and sub-horizontal tension veins, together with related sheared veins and fault planes displaying sub-vertical and sub-horizontal slickensides along drill cores are indications for the presence of extensional and oblique to strike-slip faults (see Figure 6-3 and Figure 6-4). These faults may correspond to the magnetic lineaments present in the area. Investigations aiming to define the location, as well as the orientation and kinematics of these structures are discussed in more detail in Section 6.6.

Microstructures presented in Figure 6-5 suggest the presence of extensional and strike-slip to oblique faults in the area as follows: (A) Spaced foliation and breccia in the contact zone between the carbonatite and the Pennsylvanian sequence; Subvertical (B) and subhorizontal (C) tension veins and associated sheared veins in the carbonatite; Fault planes showing subvertical (D) and oblique (E) slickensides in the carbonatite. Note that observations were made on cores from subvertical holes (about 70° plunge).

 

Source: Drenth,2014

Figure 6-4: (alt): Cross-Section A-A’ (NW to SE) and B-B’ (SW to NE) from Figure 7-1

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Source: SRK, 2014a

Figure 6-5: Core Photographs Showing Microstructures

Figure 6-6 presents a schematic diagram of microstructures along a composite subvertical drill core, suggesting that the carbonatite is intrusive within the Pennsylvanian rock sequence.

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Source: SRK, 2014a. Illustration not to scale.

Figure 6-6: Schematic of Drill Hole Showing the Typical Transition from Pennsylvanian Sediments to Carbonatite Units

Figure 6-7 and Figure 6-8 outline specific intervals within drill hole NEC14-022. These intervals include the contact between the Pennsylvanian sequence and the corresponding next approximately 5 m of carbonatite below the contact with the sediments. NEC14-022 is located on the southeastern extent of the mineralized carbonatite and within approximately 250m of the proposed production and ventilation shaft locations. The faulted and/or fractured mudstone contact is approximately 2 m in thickness and is bound by massive limestone and barite dolomite carbonatite units. The contact has had little to no water movement along or near the contact for the mudstone is relatively fresh and not weathered or significantly stained due to water flow.

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Source: Nordmin, 2019

Figure 6-7: Drill Hole NEC14-022, ~ 2m Interval of the Mudstone Contact Between the Pennsylvanian Sediments and the Carbonatite Units

 

Source: Nordmin, 2019

Figure 6-8: Drill Hole NEC14-022, Relatively Massive Dolomitic Carbonatite ~ 3m Below the Contact with the Pennsylvanian Sediments in Figure 6-7

6.3.1 Age Dating

An original hypothesis suggested that the Elk Creek Carbonatite was of Keweenawan age (Treves et al., 1983) or ca. 1,100 Ma. In 1985, Paterman, of the USGS Isotope Laboratory, provided a K-Ar

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age of 544 (±7) Ma (Cambrian) from biotite within the carbonatite. Two more K-Ar dates were provided by Georgia State University (M. Ghazi (date unknown)) which also provided dates from biotite samples. The ages of 464 (±5) Ma and 484 (±5) Ma, respectively, are Ordovician and thus much younger than the Midcontinent Rift System. While these radiometric dates provide a generalized time range for the carbonatite intrusion, additional age dating is required to establish a more precise date.

6.4 Carbonatite Lithological Unit

The lithological units present in the carbonatite complex were defined by Molycorp during their drill programs and simplified by Dahrouge for interpretation purposes during each stage of the Project (2011 and 2014). The units in Table 6-1 (youngest at the top) represent the data captured during the 2011 field program. The information was compiled from the drill logs and the corresponding geology reports for each drill hole.

Table 6-1: Project Rock Types as Defined by Molycorp and Dahrouge (2011)

Name (Molycorp)	Code	Name (Dahrouge)	Code
Overlying Lithologies
Quaternary sediments	Qt	Overburden	Ovb
Pennsylvanian Sediments	Pu	Pennsylvanian Sediments	sed
Elk Creek Complex
Younger Mafic Rock	ym		mafBc
Barite Beforsite III Barite Beforsite II	bb III bb II	Barite Dolomite Carbonatite	dolCarb
Beforsite Breccia	bbx	Dolomite Carbonatite Breccia	dolCarbBc
Barite Beforsite I	bb I	Barite Dolomite Carbonatite	dolCarb
Apatite Beforsite II Apatite Beforsite I	ab II ab I	Apatite Dolomite Carbonatite Breccia	dolCarb
Older Mafic Rock	om	Mafic dyke, vein or fragment	maf
Magnetite Beforsite	mb	Magnetite Dolomite Carbonatite	mdolCarb
Syenite II Syenite I	sy II sy I	Syenite	sy
Host Rocks
Granite/Gneiss	pCgg	Granite/Gneiss	gn
Amphibole Biotite — Gneiss	pCbg	Amphibole Biotite — Gneiss	gn

A study of six Molycorp drill holes by Xu (1996) identified two main phases within the area, a carbonate phase, and a silicate phase. The study was based on drill holes 2-B-71 (also known as “NN-1”), EC-40, EC-42, EC-50, EC-70, and EC- 82.

The carbonatite phase was classified into two main units (defined by texture, massive or brecciated) and several sub-units (defined by mineralogy as presented below).

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Massive Carbonatite

● Dolomite carbonatite

● Apatite bearing dolomite carbonatite and pyrochlore-bearing carbonatite

● Apatite dolomite carbonatite

● Hematite dolomite carbonatite

● Magnetite dolomite carbonatite

Brecciated Carbonatite

The silicate phase was also classified into several units as follows:

● Altered basalt

● Altered lamprophyre

● Altered syenite

In the 2014 drilling, the Dahrouge geologists split the dolCarb units down into a number of key units using the information of the different phases of carbonatite. The main carbonatite lithologies used are:

● Dolomite Carbonatite — dolCarb

● Dolomite Carbonatite Breccia — dolCarbBc

● Hematite dolomite Carbonatite — hemdolCarb

● Magnetite dolomite Carbonatite — mdolCarb

● Magnetite dolomite Carbonatite Breccia — mdolCarbBc

Dahrouge considers the more detailed split of the carbonatite units to be relevant to determining the distribution of different grade populations as supported by statistics (discussed in Section 11.3). The most significant difference is the change in the logging codes between dolCarb and mdolCarb, in terms of the major rock types.

6.5 Marine Sedimentary Rocks

The state-wide Nebraska test hole database contains information for about 5,500 test holes drilled since 1930 by the CSD (Conservation and Survey Division of the University of Nebraska-Lincoln (UNL), School of Natural resources (SNR), (UNL-CSD/SNR), and cooperating agencies. Test hole location data, as well as lithological descriptions, stratigraphic interpretations, and geophysical log records, are included in the database. In addition, UNL-CSD/SNR maintains an extensive collection of geologic samples obtained from the drilling process (UNL-CSD/SNR website).

The overlying sedimentary units on the Project are of Pennsylvanian age. The CSD’s 1971 test hole 2-B-71, also labelled NN-1 by Molycorp, intersected several formations of overlying Pennsylvanian strata (see Table 6-2).

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Table 6-2: Stratigraphy Overlying the Elk Creek Carbonatite

System	Series	Group	Formation	Member	Depth From (ft)	Depth To (ft)
Quaternary	-	-	-	-	0.00	43.90
Pennsylvanian	Virgilian	Wabaunsee	Zeandale	Wamego	43.90	82.50
Pennsylvanian	Virgilian	Wabaunsee	Emporta	Elmont	82.50	95.00
Pennsylvanian	Virgilian	Wabaunsee	Auburn	-	95.00	113.50
Pennsylvanian	Virgilian	Wabaunsee	Bern	Wakarusa	113.50	138.60
Pennsylvanian	Virgilian	Wabaunsee	Scranton	-	138.60	238.80
Pennsylvanian	Virgilian	Wabaunsee	Howard	-	238.80	243.10
Pennsylvanian	Virgilian	Wabaunsee	Severy	-	243.10	265.50
Pennsylvanian	Virgilian	Shawnee	Topeka	Coal Creek	265.50	292.00
Pennsylvanian	Virgilian	Shawnee	Calhoun	-	292.00	292.80
Pennsylvanian	Virgilian	Shawnee	Deer Creek	Ervine Creek	292.80	331.00
Pennsylvanian	Virgilian	Shawnee	Tecumseh	-	331.00	341.50
Pennsylvanian	Virgilian	Shawnee	Lecompton	Avoca	341.50	369.00
Pennsylvanian	Virgilian	Shawnee	Kanawaka	-	369.00	370.00
Pennsylvanian	Virgilian	Shawnee	Oread	Kereford	370.00	422.30
Pennsylvanian	Virgilian	Douglas	-	-	422.30	478.40
Pennsylvanian	Missourian	Lansing	Stanton	South Bend	478.40	494.70
Pennsylvania	Missourian	Lansing	Stanton	Rock Lake	494.70	500.00
Pennsylvanian	Missourian	Lansing	Stanton	Stoner	500.00	515.10
Pennsylvanian	Missourian	Lansing	Vilas	-	515.10	516.40
Pennsylvanian	Missourian	Lansing	Plattsburgh	-	516.40	523.40
Pennsylvanian	Missourian	Kansas City	Bonner Springs	-	523.40	526.50
Pennsylvanian	Missourian	Kansas City	Wyandotte	Farley	526.50	565.00
Pennsylvanian	Missourian	Kansas City	Lane	-	565.00	567.40
Pennsylvanian	Missourian	Kansas City	Iola	-	567.40	590.00

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Pennsylvanian	Missourian	Kansas City	Chanute	-	590.00	594.40
Pennsylvanian	Missourian	Kansas City	Drum	-	594.40	602.50
Pennsylvanian	Missourian	Kansas City	-	-	602.50	628.30
Cambrian	Undifferentiated	-	Elk Creek Carbonatite	-	628.30	952.00

Test Hole 2-B-71 or NN-1

Source: McCallum and Cathro, 2010

There are active limestone quarries, and underground mines within approximately 70 km of the project site that create road materials, lime, back fill, and construction materials. These quarries are actively mining approximately 2.2 million tons/year from within the Pennsylvanian limestone units. The Pennsylvanian limestone unit is the same as is currently located above the carbonatite unit at the project site.

6.6 Structural Geology

Based on data provided to carry out the structural study, the Project contains five main sets of brittle faults variably cutting through the Pennsylvanian rocks and the carbonatite boundary which appears to be tectonic. The orientations of the faults were determined by comparing Acoustic Televiewer (ATV) logs with specific customized structural core logging data, and by undertaking a preliminary interpretation of the provided geophysics images.

This data has been used to model the fault pattern in 3D for use in further resource estimation and geotechnical studies. The overall fault model included approximately 28 structures with the vicinity of the Project with varying levels of confidence. Based on a review within the mineralization, at least three key northeast-trending faults have been identified and used during the geological modeling process.

The joints and veins define orientation sets comparable to the fault trends. Hematite veins, which may be up to a metre thick, represent the weakest fault- and joint-infilling material which may be problematic for mining and should, therefore, be given more attention during any future geotechnical studies.

6.7 Mineralization

The property hosts niobium, titanium, and scandium mineralization as well as REE and barium mineralization that occur within the Elk Creek Carbonatite.

The current extent of modelled mineralization is 750 m along strike, 500 m wide, and 750 m in dip extent below the unconformity. Figure 6-9 demonstrates that the mineralization is open in all directions. For this TRS, niobium, titanium and scandium are considered the main elements of interest, with niobium and REE mineralization contributing to the resource estimate, discussed below.

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Source: Nordmin, 2019

Figure 6-9: Plan View of the Location of the Mineralized Carbonatite (Mcarb)

The initial Molycorp drill hole database contained a separate geological report summarizing rock types, assay results and associated petrographic descriptions identifying niobium and/or REE minerals. Niobium was reported to be hosted in pyrochlore, and REE mineralization was reported to occur as bastnäsite, parisite, synchysite and monazite. SRK subsequently highlighted that the level of detail shown in the geological reports had not been transferred to the resource model. Since 2014, the database has been improved by geologists familiar with the current logging codes, conducting a review of the historical logs, reports, and available drill core to provide an rigorous geological database.

6.7.1 Niobium and Titanium Mineralization

The deposit contains significant concentrations of niobium. Based on the metallurgical test work completed to date at a number of laboratories using QEMSCAN® analysis, the niobium mineralization is known to be fine-grained, and that 77% of the niobium occurs in the mineral pyrochlore, while the balance occurs in an iron-titanium-niobium oxide mineral of varying composition. Distribution and statistical review of Nb₂O₅ within the mineralized carbonatite, are shown in Figure 6-10 and discussed in Section 8 and Section 11.

Figure 6-11 demonstrates that there is a fairly high correlation between increasing Nb₂O₅ grade and Fe₂O₃ and TiO₂ grades.

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Source: Nordmin, 2019

Figure 6-10: Basic Statistics of Nb2O5 Mineralization

Source: Nordmin, 2019

Figure 6-11: Correlation Statistics of Nb₂O₅ and TiO₂ and Fe₂O₃

6.7.2 Scandium Mineralization

Within the Elk Creek Carbonatite, a host of other elements exist with varying degrees of concentration. The Company has completed both whole rock analysis and multi-element analysis

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on all samples for the 2014 program, plus re-sampling programs of selected historical core and/or pulps between 2011 and 2021.

As the metallurgical test work advanced during 2014 and 2015, the ability to obtain a titanium dioxide (TiO₂) and scandium (Sc) product became apparent. TiO₂ is strongly and positively correlated with niobium grades, whereas the scandium mineralization is spatially related to niobium and titanium mineralization, but with lesser degree of correlation. Basic statistics for Sc mineralization are shown in Figure 6-12. Detailed discussion is presented in Section 11.

Source: Nordmin, 2019

Figure 6-12: Basic Statistics of Sc Mineralization

6.7.3 Rare Earth Element Mineralization

Within the Elk Creek Carbonatite complex, there are several occurrences of REE mineralization, including the Project area. REE mineralization within the carbonatite occurs within the following minerals:

● Bastnäsite ([Ce,La,Y]CO₃F)

● Parisite (Ca[Ce,La]₂[CO₃]3F₂)

● Synchysite (Ca(Ce,La)(CO₃)₂ F)

● Monazite ([Ce,La]PO₄)

Quantum's re-sampling program discovered high grade REE mineralization in EC-93 as noted in this Molycorp drill logs excerpt:

"Barite beforsite is the predominant lithology from 149.4 to 304.8 m. It contains xenoliths of syenite, older mafic rocks, and apatite beforsite I, and is intruded by younger mafic rocks. Intervals 33 m (100 ft) long contain 2.13% to 2.75% LnO from 149.4 to 274.3 ft. An interval 18.3 m long at 179.8 to 198.1 ft contains 3.89% LnO. The highest-grade mineralization intercepted was 3 m at 4.72% LnO at 155.4 to 158.5 m. Lanthanide minerals occur as radial patches and random aggregates of needles, irregular patches and vein-like aggregates. The aggregates occur with and without quartz. The aggregates appear as light-gray patches in reddish-brown, hematite-altered beforsite. Although individual lanthanide mineral grains are

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in the micrometre size range, aggregates of lanthanide minerals range from 0.23 to 8 mm. in maximum dimension. Monazite and bastnäsite have been identified in the aggregates, and EDX spectra show Ce > La."

It should be noted that Molycorp’s term LnO (lanthanide oxides) or rare-earth oxides (REO) incorporates lanthanum, cerium and neodymium along with the other 12 rare earth elements. Present day nomenclature for REE is shown in Table 6-3. Promethium (Pm) is not included as it is very rare in nature. The division into light and heavy rare-earth elements made below is based on differences in processing. Elsewhere in literature, the division has been made between gadolinium and terbium (atomic number 64 and 65) based on the lack of paired electrons in the inner incomplete subshell (4f) (Van Gosen et al., 2017).

Statistical analysis of distribution and correlation of REEs within the deposit are presented in Section 11.

Table 6-3: List of Elements and Oxides Associated REE Mineralization

Element	Element Acronym	Compound
Associated Elements and Oxides Niobium	Nb	Nb2O5
Light Rare Earth Metals and Oxides (LREO) Lanthanum Cerium Praseodymium Neodymium	La Ce Pr Nd	La203 Ce203 Pr203 Nd203
Heavy Rare Earth Metals and Oxides (HREO) Samarium Europium Gadolinium Terbium Dysprosium Holmium Erbium Thulium Ytterbium Lutetium Yttrium	Sm Eu Gd Tb Dy Ho Er Tm Yb Lu Y	Sm2O3 Eu203 Gd203 Tb203 Dy203 Ho203 Er203 Tm203 Yb203 Lu203 Y203

Source: SRK, 2017

6.8 Deposit Types

The Project is hosted within the Elk Creek Carbonatite. By definition, a carbonatite is an igneous rock body with greater than 50% modal carbonate minerals, mainly in the form of calcite, dolomite, ankerite, or sodium- and potassium-bearing carbonates. Carbonatites commonly occur as intrusive bodies, such as isolated sills, dykes, or plugs, although they can rarely occur as

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extrusive rocks (Oldoinyo Lengai, Tanzania). Many carbonatites are associated with alkalic silicate complexes which include syenite, nepheline syenite, ijolite, urtite, and pyroxenite. Carbonatites are generally related to large-scale, intra-plate fractures, grabens, or rifts that correlate with periods of extension, and range from Precambrian to recent in age. They are usually surrounded by an aureole of metasomatically altered rocks called fenites. Carbonatite-associated deposits can be classified as magmatic or metasomatic types (Richardson and Birkett, 1996).

Carbonatites have been classified based on chemical classification into four classes (Woolley and Kempe, 1989; Wyllie and Lee, 1998), and further subdivided based on mineralogical and textural characteristics:

● Calcio-carbonatite coarse-grained: sövite, and finer-grained: alvikite

● Magnesio-carbonatite dolomite-rich: beforsite, and ankerite-rich: rauhaugite

● Ferro-carbonatite (iron-rich carbonates)

● Natro-carbonatite (sodium-potassium-calcium carbonates)

The use of a chemical classification of carbonatites should be used with caution when replacement, or metasomatic, processes have altered the primary composition of the carbonatite rock (Mitchell, 2005).

The majority of carbonatite deposits are located within stable, intra-plate crustal units, although some are linked with orogenic activity or plate separation. It is also important to note that carbonatites tend to occur in clusters, and in many places, there has been a repetition of intrusive activity over time (Woolley, 1989).

Carbonatite-hosted deposits occur almost exclusively in intrusive carbonatite and may be subdivided into magmatic, replacement/veins, and residual sub-types. The Elk Creek Carbonatite can be classified as a magmatic sub-type, similar to the St-Honoré deposit in Quebec, Canada (Niobec niobium mine, Iamgold – Figure 6-13), the Mountain Pass Deposit in California, U.S.A. (REE), and the Palabora Deposit in South Africa (apatite).

The pipe-like carbonatites typically occur as sub-circular or elliptical shapes and can be up to 3-4 km in diameter. Magmatic mineralization within pipe-like carbonatites is commonly found in crescent shaped, steeply-dipping zones. As carbonatite magma is typically volatile rich with low viscosity, it may ascend rapidly through the mantle, fracturing the crust on impact, causing a characteristic alternating ring (crescent) structure of carbonatite and wall rock to be formed. Metasomatic mineralization occurs as irregular forms, breccias, or veins. Carbonatites typically consist of multiple phases of intrusion with different mineralogical and textural characteristics. Early phases tend to consist mainly of calcite with later phases mainly consisting of dolomite, ankerite, or siderite. The later phases are typically more enriched in niobium or tantalum with the latest phases more enriched in rare earth minerals. In general, geochemical zonation of phases begin with calcio-carbonatite intrusion, followed by magnesio-carbonatite and finally ferro-carbonatite. Fenitization (alkali metasomatism) is common around many carbonatite intrusions.

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Source: IAMGOLD, 2011

Figure 6-13: Schematic diagram of St. Honoré Carbonatite

The major mineral constituents are calcite, dolomite, siderite, ferroan calcite, ankerite as carbonates, and hematite, biotite, titanite, olivine, and quartz. Economic minerals include fluorite (F), apatite (P), pyrochlore (Nb), anatase (Ti), columbite (Nb-Ta), monazite (REE), bastnaesite (REE), parasite (REE), zircon (Zr), and magnesite (Mg), among others. Mineralization within carbonatites is typically syn- to post-intrusion. The mineralization is controlled primarily by fractional crystallization within the intrusion, with tectonic and local structures influencing the form of metasomatic mineralization (Woolley and Kempe, 1989; Richardson and Birkett, 1996; Birkett and Simandl, 1999).

Worldwide, carbonatite deposits are mined for niobium, REE, iron, copper, phosphate (apatite), vermiculite and fluorite; with barite, zircon/baddeleyite, tantalum and uranium as common by-products (Richardson and Birkett, 1996).

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7. EXPLORATION

The carbonatite complex is a 6 km to 8 km diameter, alkaline intrusive complex that is buried under approximately 200 m of Pennsylvanian marine clastic sedimentary rocks in southeastern Nebraska. There is no surface expression of the complex, therefore all exploration has been either through geophysics or drilling. The complex is composed of several lithologies. Apatite dolomite is the volumetrically dominant lithology, followed by undifferentiated mafic rocks, syenite, dolomite breccia, barite, and a small body of magnetite dolomite. The magnetite dolomite is the primary host of the niobium mineralization (see Figure 7-1).

Source: Benjamin Drenth, September 2014

Figure 7-1: Geology of the Elk Creek Carbonatite as Expressed in Drill Holes at an Elevation of 120 m Above Sea Level (Roughly 230 m Below Ground Surface)

Note: the term beforsite used in this figure has been superseded by the terms magnesio-carbonatite or dolomite carbonatite. Other rock-type names have been modified subsequently (see Table 6-1).

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7.1 Geophysical Exploration

Several geophysical programs have been conducted on the Property. In 2020, interpretation of a ground gravity survey conducted by Molycorp in 1975 over the Elk Creek Carbonatite was used to plan an airborne gravity gradiometer and airborne magnetic survey over the complex. A FALCON™ airborne gravity gradiometer (gD) and total magnetic intensity (TMI) geophysical survey was flown over the Property by Fugro Airborne Surveys Pty Ltd. (Fugro) from April 5-12, 2011. A total of 1176 line km at a nominal traverse line spacing of 200 m north-south with a nominal tie line spacing of 2000 m east-west using a Cessna C-208B Grand Caravan. The resulting FALCON™ datasets - vertical gravity gradient (GDD) mathematic integral of GDD (gD) equivalent to Bouguer gravity, and the total magnetic intensity (TMI) were interpreted by Condor Consulting of Lakewood CO (Condor)). Both Fugro and Condor are respected independent geophysical contractors. In the opinion of Dahrouge, the sampling methods and quality, including line spacing and data density are consistent industry standards. No likelihood of introduced sample bias was noted.

7.2 Drilling

7.2.1 Type and Extent

Mineral resource definition drilling at the Project was conducted in three phases. The first was during the 1970s and 1980s by Molycorp, the second in 2011 by Quantum, and the third and latest program in 2014 by NioCorp. Re-sampling programs of historical Molycorp core and pulps were conducted between 2010 and 2014 to verify results and QA/QC procedures. In 2015 a re-assay program was conducted on pulps to add scandium and titanium analysis. Further re-sampling was conducted in 2016 and 2021. To date, 143 drill holes have been completed for a total of 70,897 m (see Table 7-1 and Figure 7-2). A further five holes totalling 3,353.1 m were drilled in 2015, for hydrogeological and geotechnical studies but were not used for resource estimation. All drilling has been completed using a combination of tricone, reverse circulation (RC) or diamond drilling (DDH) core in the upper portion of the hole within the Pennsylvanian sediments. A portion of the 2014 drill holes used RC drilling within the Pennsylvanian sediments, to increase the efficiency in drilling through the cover material, within areas of strong geological confidence. All drilling within carbonatite has been completed using diamond coring methods.

To date, local labour has been used by drilling contractors when preparing the drill hole pads. All drilling has been completed using standardized procedures which are in line with international standards of best practice. The drilling by Molycorp was completed using company-owned equipment and sampling procedures. The drilling companies used by the Company during the 2011 - 2014 resource drilling, and 2015 geotechnical drilling programs are detailed below:

● 2011: Black Rock Drilling, LLC (BRD Personnel and Leasing Corp.), 17525 E Euclid Ave, Spokane Valley, WA 99216

● 2014: Envirotech Drilling LLC, 900 East 4th Street, Winnemucca, NV 89445

● 2014: West-Core Drilling, LLC, 561 W Main Elko, NV 89801 USA; and

● 2014: Idea Drilling, 1997 9th Avenue North, Virginia, MN 55792

● 2015: Idea Drilling, LLC, 1997 9th Avenue North, Virginia, MN 55792

● 2015: Envirotech Drilling LLC, 900 East 4th Street, Winnemucca, NV 89445

The drilling has been completed using conventional techniques and experienced drilling contractors.

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The following sections provide a summary of the resource drilling completed by Molycorp, Quantum and NioCorp (as shown in Figure 7-2).

Table 7-1: Summary of Drilling Database within the Geological Complex

Year	Company	Number of Holes	Average Depth (m)	Sum Length (m)
1970-1980	Molycorp	114	422	48,156
2011	Quantum	5	684	3,420
2014-2015	NioCorp	24	840	19,322
Subtotal		143	406	70,897

Source: NioCorp, 2022

One hole (NEC14-MET-03) from the 2014 metallurgical testing has been used in the current resource estimation (Table 7-3).

During 2015 five holes (NEC15-001 to 005), for a total length 3,353.1 m, were drilled. for hydrogeological and geotechnical studies. The drilling was carried out by Idea Drilling and Envirotech Drilling LLC with Envirotech Drilling LLC as subcontractor.

Source: NioCorp, 2021

Figure 7-2: Elk Creek Drill Hole Location Map

Not all the drill holes within the Project were used in the 2022 Mineral Resource Estimation, as many do not intersect the Nb₂O₅ anomaly and are located a significant distance away from the

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Deposit (Figure 7-2). There are a total of 143 drill holes within the Project, of these 45 drill holes were used to inform the Elk Creek Deposit Mineral Resource Estimation (Figure 7-4). Note that there are more holes within the project area, but some holes were excluded from the Mineral Resource as they were drilled for other purposes (geotechnical, hydrogeology) and were not sampled:

Table 7-2: Summary of Drilling Database within Elk Creek Deposit Area

Year	Company	Number of Holes	Average Depth (m)	Sum Length (m)
1970-1980	Molycorp	27	597	16,108
2011	Quantum	3	773	2,318
2014-2015	NioCorp	24	805	19,321
Subtotal		54	699	37,747

Source: NioCorp, 2022

Table 7-3: Summary of Drilling Database used in the Current Resource Estimation

Year	Company	Number of Holes	Average Depth (m)	Sum Length (m)
1970-1980	Molycorp	25	597	14,919
2011	Quantum	3	773	2,318
2014	NioCorp	17	836	14,209
Subtotal		45	699	31,445

Source: NioCorp, 2022

7.2.2 Molycorp, 1973-1986

Between 1973 and 1986, Molycorp completed a regional scale drill program over an approximately 7 km by 7 km gravity anomaly that included the Elk Creek Deposit. The total program consisted of 114 drill holes for a total of approximately 48,156 m. Outside the Elk Creek Deposit area, the regional drill program was conducted on a regular grid of 610 m by 610 m (2,000 ft by 2,000 ft) with some closely spaced holes in selected areas within the gravity anomaly (see Figure 7-2).

Included in this total, 27 holes totalling 16,108 m were drilled over the deposit, of which 25 were used in the current estimate (Table 7-3). Drilling orientations varied considerably.

The Molycorp drill hole locations centred over the Elk Creek Deposit are presented in Figure 7-3 (shown in gold).

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Source: NioCorp, 2021

Figure 7-3: Elk Creek Drill Hole Location Map by Operator

7.2.3 Quantum, 2011

In April 2011, Quantum conducted a preliminary drill program (three holes, Figure 7-3) on the Elk Creek Deposit along with two initial holes focused on REE enrichment targets. These holes have been excluded from the current Mineral Resource Estimation, as they do not intersect the Nb₂O₅ anomaly and are located to the east. The objectives of the drill program over the Project were to verify the presence of higher-grade niobium mineralization at depth and to infill drill the known niobium deposit to upgrade the resource category of the previous resource estimate and expand the known resource. The drill program was also established to collect sufficient sample material for metallurgical characterization and process development studies of the niobium mineralization.

The 2011 program consisted of five inclined drill holes, totalling 3,420 m of NQ size diameter core. Inclusive of this total, three drill holes, totalling 2,318 m were drilled into the known Elk Creek Deposit. The summary of the 2011 drill program is listed in Table 7-4.

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Table 7-4: Summary of the 2011 Drill Program

Drill Hole	UTM Easting	UTM Northing	Elevation (m)	Depth (m)	Bearing (°)	Dip (°)
NEC11-001	739299.0	4461052.0	341.49	900.38	28.1	-72.0
NEC11-002	738955.0	4461058.0	343.88	908.61	33.5	-61.0
NEC11-003	739417.0	4461060.0	340.79	508.71	34.3	-55.9
Outside Elk Creek Deposit; REE Exploration Targets
NEC11-004	741997.0	4460790.0	333.65	465.73	80.7	-55.6
NEC11-005	740604.0	4461660.0	337.48	636.42	95.7	-56.0
Total				3,419.85

Source: Tetra Tech, 2012

DDH NEC11-001 targeted the eastern portion of the deposit below the historical drill hole EC-11 and between vertical holes EC-27 and EC-30. DDH NEC11-002 was drilled into the northwestern portion of the deposit. DDH NEC11-003 was drilled into the southeastern portion of the deposit. Drill holes NEC11-004 and 005 drilled into regional REE targets and are not included in the Mineral Resource Estimate presented in this report.

The Quantum 2011 drill hole locations centred over the Elk Creek Deposit are presented in Figure 7-3: Elk Creek Drill Hole Location Map by Operator (shown in red).

Results from the 2011 drilling program provided additional information on areas of the deposit at depth where limited information was previously available. The drill holes confirmed the high-grade potential of the niobium mineralization, as indicated by the previous drilling completed by Molycorp.

7.2.4 NioCorp 2014 Program

The 2014 drilling campaign was conducted as a three-phase program. The campaign was designed to increase resource confidence by moving Inferred to Indicated resource category, based on the 2012 Mineral Resource Estimate. The program was initially designed for 14 drill holes totaling approximately 12,150 m (NioCorp press release, April 29, 2014) but was expanded to 19 drill holes, to give a total of 15,968.3 m (see Figure 7-3, Table 7-2).

The drilling was conducted by private contractors West-Core Drilling and Idea Drilling.

A location map of the drill holes used in the resource is shown in Figure 7-4. The drill holes (except for two), were designed to intersect perpendicular to the strike of the ore body, trending either southwest or northeast. Two drill holes, NEC14-011 and NEC14-012, were oriented southeast and northwest, respectively. Locations and survey information of the 2014 drill program can be found in Table 7-5.

Three of the 19 holes drilled were for metallurgical characterization development studies. Two of these drill holes (NEC14-MET-01 and NEC14-MET-02) were not sampled. The third drill hole (NEC14-MET-03) was quartered, with one quarter being sent for assays and remainder sent for metallurgical testing. No further resource holes have been drilled since the February 20, 2015, Mineral Resource Estimate. Five holes (NEC15-001 to NEC15-005) were drilled in 2015 for a total of 3,353.1 m as hydrogeological and geotechnical studies but were not sampled and do not form part of the current mineral resource estimate. The NioCorp drill hole locations (shown in Table 7-5) are presented in Figure 7-3.

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Table 7-5: NioCorp 2014-2015 Drill Hole Locations

BHID	XCOLLAR	YCOLLAR	ZCOLLAR	LENGTH	AZIMUTH	INCLINATION	COMMENTS
NEC14-006	739166.2	4461224.0	352.0	772.7	29.9	-70.8
NEC14-007	739088.2	4461083.5	344.8	907.4	29.4	-70.6
NEC14-008	739128.1	4461159.3	351.2	886.0	30.8	-69.8
NEC14-009	739390.2	4461466.2	349.3	751.3	208.7	-70.3
NEC14-009a*	739390.2	4461466.2	349.3	897.0	208.7	-70.3	Wedge: NEC14-009
NEC14-010	739209.5	4461149.8	347.8	796.1	30.0	-73.1
NEC14-011	738892.5	4461513.6	359.7	900.4	125.8	-65.3
NEC14-012	739635.1	4461083.4	339.9	843.2	299.8	-68.0
NEC14-013	739169.3	4461354.3	355.2	880.3	149.4	-89.2
NEC14-014	739034.8	4461218.6	346.1	901.0	28.6	-77.6
NEC14-015	739221.0	4461064.7	342.4	827.8	29.1	-72.4
NEC14-016	739509.1	4461574.7	354.7	913.8	210.5	-60.0
NEC14-020	739037.1	4461305.0	348.4	587.6	28.2	-70.6
NEC14-021	739074.3	4461188.5	347.1	865.0	29.5	-69.2
NEC14-022	739292.2	4461055.3	340.3	950.4	31.3	-68.4
NEC14-023	739377.6	4461071.0	341.5	615.1	30.2	-71.1
NEC14-MET-01	739240.4	4461282.7	352.8	894.7	302.6	-89.6	Not Sampled
NEC14-MET-02	739171.1	4461372.4	355.8	865.0	88.1	-89.6	Not Sampled
NEC14-MET-03	739129.9	4461414.5	355.4	913.3	249.8	-89.8
SUBTOTAL				15,968
NEC15-001	739245.3	4461337.6	354.6	832.1	360.00	-90.00	Hydrology
NEC15-002	739046.5	4460708.9	344.4	850.4	303.82	-88.32	Piezometer
NEC15-003	740346.2	4460854.1	341.6	849.5	267.35	-89.50	Piezometer
NEC15-004	739472.2	4461507.0	354.6	413.6	16.27	-89.38	Shaft Geotech
NEC15-005	739514.8	4461418.5	351.2	407.5	297.36	-88.44	Vent  Geotech
SUBTOTAL				3,353
TOTAL				19,321

* Note that NEC14-009a started at a depth of 485.51 m. Including this hole, a total of 19 holes were drilled during the 2014 program

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Source: NioCorp, 2021

Figure 7-4: Drill Hole Traces Used in the 2022 Mineral Resource

7.2.4.1  Procedures (NioCorp 2014 Program)

Detailed descriptions of Molycorp's drilling, sample procedures, analyses and security have not been documented and reviewed by Dahrouge. Given Molycorp's position as a leader in the rare earth industry at the time, it is believed Molycorp applied industry best practice for the time period. The 2011 drilling campaign was managed by Dahrouge and SRK under the same quality and procedures used in the current study. The 2014 drilling program includes a quality control program to ensure the results can be used to verify earlier drilling results and add confidence to the overall understanding of the deposit.

For the 2014 drilling program, planned drill hole collars were initially located using a handheld GarminTM Global Positioning System (GPS) and marked with wooden stakes. A tracked excavator was used to construct the drill pad and collars were then relocated using the GPS, with wooden stakes set after pad construction. A geological compass and an Azimuth Pointing System (APS) were used to sight in the drill to the planned azimuth and inclination.

The 2014 core drilling was conducted by both West-Core Drilling and Idea Drilling, both private contractors. West-Core used both an AVD R40 track-mounted core drill and an Atlas Copco CS-14 track-mounted core drill, while Idea used an Atlas Copco CT-20 truck-mounted core drill. Overburden was cased in all drill holes to depths ranging from 18 m to 37 m. The Pennsylvanian limestones and mudstones overlying the target carbonatite were drilled as PQ-sized core with HQ-

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sized core for drill holes NEC14-020 to NEC14-023. The carbonatite was drilled with the HQ-sized core, except for the three metallurgical holes (NEC14-MET-01, NEC14-MET-02 and NEC-14MET-03), which were drilled entirely using PQ-sized core. Core size reduction took place just beneath the Pennsylvanian-carbonatite contact at depths ranging from 206 m to 238 m. The core drilling rigs operated 24 hours/day and 7 days/week, with the typical progress of 40 m/day.

During the drilling operation, the core was retrieved from the core barrel and laid sequentially into wooden core boxes by the drilling contractor. Interval blocks were placed at all run breaks. Once the box was full, the ends and top of the box were labelled with drill hole identification and the sequential box number. Upon completing a box, it was stacked on a pallet or a truck bed at the drill rig. At the end of each drilling shift, the boxes of the core were transported by the drilling contractor in a pickup truck to the NioCorp field office. At this point, the core was in the custody of Dahrouge. Eight of the 2014 drill holes had piezometers installed in them after drilling was complete. For these drill holes, surface completion consisted of surface casing capped with a locking steel cover, a 1.2 m² cement pad around the surface casing and a steel nameplate attached to the casing (see Figure 7-5).

Surface completion for the drill holes that did not have piezometers installed consisted of a steel marker post and attached nameplate. All nameplates included drill hole number, total depth, and orientation. Abandonment of the drill holes consisted of cementing from total depth to surface in the non-piezometer drill holes and from total depth to the bottom of the piezometers in the other drill holes with piezometer installations (see Figure 7-6).

Source: Nordmin, 2019	Source: Nordmin, 2019
Figure 7-5: Collar Location of NEC14-MET-01	Figure 7-6: Collar Location of NEC14-009

7.2.4.2  Collar Surveys

All drill hole collars were initially surveyed before drilling using a handheld GPS. On completion of the drill hole an external contractor, ESP INC. (Engineering/Surveying/Planning), based in Lincoln, Nebraska, was used to provide a detailed survey of the collar location using a Sokkia GS2700 IS GPS, which has 10 mm horizontal and 20 mm vertical accuracy. Data was provided to SRK in digital format in NAD 1983 UTM Zone 14N grid coordinates.

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The locations of 24 of the 29 Molycorp drill collars were re-excavated as required to confirm the collar coordinates from 2011 drilling over the Elk Creek Carbonatite, by CES Group P.A. Engineers & Surveyors (CES), based in Kansas City, Missouri. All collars were surveyed using the same UTM coordinate system.

7.2.4.3  Downhole Surveys

Initial collar surveys of dip and azimuth were taken using compass measurements for all holes (RC and DD). Downhole surveying was undertaken on historical Molycorp holes drilled below the Pennsylvanian sediments at an interval of 30.48 m (100 ft).

The 2011 drilling program was surveyed at 3.05 m (10 ft) intervals, based on the drilling rod lengths used at the time. All drill holes were surveyed immediately after completion of drilling. Downhole deviations, subsurface azimuth, and dip were mapped using a Devico DeviFlex survey tool, which is a nonmagnetic, electronic, multi-shot tool. The DeviFlex tool consists of two independent measuring systems, with three accelerometers and four strain gauges used to calculate inclination and change in azimuth.

The DeviFlex tool communicates with a handheld PDA (personal digital assistant), and the survey results can be viewed on the PDA immediately after completion of the survey. Dahrouge geologists checked the downloaded data for possible errors, and inconsistencies and some readings were removed for quality control purposes.

The DeviFlex output contains a column for possible tool movement during surveying. In the event, there was a potential tool movement, that specific reading was removed from the dataset.

The DeviFlex tool records changes in azimuth, as opposed to absolute azimuth measurements. Because of this, an initial (surface) survey azimuth is required to calibrate the DeviFlex downhole azimuth readings. CES surveyed all initial drill hole azimuths by surveying the azimuth of the drill rods extending from the ground during drilling. These initial azimuth readings were used to calibrate the DeviFlex downhole change in azimuth readings and calculate absolute azimuth measurements.

The 2014 drilling program was surveyed at 6.1 m (20 ft) intervals using a Reflex GYRO survey tool. Dahrouge geologists operated the GYRO and collected the surveys. Downhole deviations, subsurface azimuth, and dip were mapped with the GYRO, which utilizes a digital MEMS-angular rate sensor non-magnetic assemblage. The GYRO tool was used to mitigate magnetic deviation caused by metal equipment, or naturally occurring minerals such as magnetite and pyrrhotite found in the deposit.

These surveys were synchronized electronically with a receiver at the surface, and recordings were collected every 30 seconds after the tool equilibrated. The Reflex GYRO has an integrated Azimuth Pointing System (APS) that is used to orient the True North azimuth, a GPS position and degree of inclination. Downhole surveys were completed through the drill rods, and location data points were collected every 6.1 m (20 ft).

The methods used for the downhole surveying during the 2011 and 2014 campaigns meet acceptable industry standards. Given the drill hole lengths of over 700 m, the Company has used suitable techniques to provide a continuous (ranging from 3 to 6 m interval) measure of the drill hole trace from the base of the hole. The use of a GYRO has avoided any potential issues due to the magnetic nature of the rocks. The confidence in the drill hole location of the Molycorp drilling is considered slightly lower due to their historic nature and the wider measurement spacing and equipment used to complete those surveys.

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Overall, Dahrouge considers the level of confidence in the downhole surveys to be sufficient and adequate for the use in Mineral Resource Estimation practices.

7.2.5 Interpretation and Relevant Results

No new drilling has been completed since 2015 for the purposes of Mineral Resource definition.

The 2022 estimate is based on the results of additional assays of Molycorp holes held by the University of Nebraska – Lincoln’s Conservation and Survey Division that were obtained in 2021 and added to the assay database ofr the Project. The assays were executed using an analytical package at Actlabs in Ancaster, Ontario that included Nb, Sc, Ti, and the rare earth elements.

Source: SRK, 2015

Figure 7-7: 3D View of Elk Creek Deposit Showing Modelled Base of Till and Unconformity between Pennsylvanian Sediments and the Elk Creek Carbonatite

The only significant change in the database from 2019 to 2022 is the additional assay of 1186 samples (inclusive of quality control samples) from within the previous footprint of the Mineral Resource, to fill in gaps in available rare earth analyses. Those assays were completed at Actlabs in Ancaster, Ontario.

The drilling was conducted by reputable contractors using industry-standard techniques and procedures. This work has confirmed the presence of niobium, titanium, scandium and rare earth mineralization hosted in dolomite-carbonatite and lamprophyre rocks. In general, the lamprophyre is niobium-depleted, but contacts between lamprophyre and carbonatite may be enriched.

The drill holes within the Deposit and Mineral Resource area have variable drill spacing between 25 m and 225 m. The major drilling direction used by NioCorp has been towards the northeast. Two sets of scissor holes were drilled to the southwest on separate drilling lines within the central portion of the deposit, to confirm that there is no directional bias in the selected drill hole

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orientation. The majority of the holes have inclinations in the order of 60° to 70°. The use of scissor holes has confirmed the sub-vertical nature of the southwest contact (see Figure 7-7).

7.3 Geotechnical Design Parameters

From May 21, 2014, to July 30, 2015, SRK completed a geotechnical investigation program on site for the Project. The program was designed to characterize subsurface geotechnical conditions to assist in the development of a Feasibility Study design capable of meeting the requirements for basic engineering design.

The geotechnical database used for analyses includes data from two geotechnical borehole databases: the 2014 program and a previous 2011 program. A geotechnical model was created using the characterization information in the databases to estimate rock mass quality, rock strength, and major discontinuities in the carbonatite hanging wall and footwall and the Pennsylvanian Formation. The 3D geotechnical model was built using the Vulcan software, wherein representative volumes of rock mass quality domains were created.

The following is a summary of the geotechnical parameters used to assess the mine design.

Data Collection

The geotechnical field investigation consisted of 23 drill holes, totalling 20,379 m. The program was designed to examine rock mass fabric and structural features in and around the mineralized zone at different depths and orientations. The drilling was conducted in three phases with incremental data collection designed to fill knowledge gaps on geotechnical conditions. Drill holes were drilled at varying orientations into the hanging wall, footwall, and mineralized rock to capture data on rock mass discontinuity variations.

The drill holes are shown in Figure 7-8 in plan view. Figure 7-9 shows a vertical view looking towards the north with the mining levels and the planned access ramp. Figure 7-10 shows the same vertical view looking towards the north with wireframe shapes of the high- grade niobium
(Nb₂O₅ > 1%). The drill hole location data are summarized in Table 7-6. The field investigation included drilling of the core, geophysical borehole logging of structural features, geotechnical core logging, core sample collection for laboratory strength testing, and in situ stress measurements.

Structural features (discontinuities) encountered during this field investigation consisted of joints, lithological contacts, veins, dikes, foliation, faults, shear zones and fractures. Orientation data was collected using acoustic televiewer (ATV) down-hole equipment. Typical alteration on discontinuity surfaces was recorded in the logs and included mineralized coatings or infillings with occasional hematite staining. Faulting was interpreted using evidence of slickensided surfaces and the presence of gouge infilling and orientations verified using the ATV orientation data.

Samples were collected for laboratory testing. A series of tests (UCS, TCS, BTS, DSS) were conducted on samples from the upper and lower Pennsylvania formation as well as from the hangingwall, footwall and orebody within the Carbonatite formation, The range of samples were assessed to be representative of each rock mass region. All the tests were conducted in accordance with ASTM and ISRM standards. These standards are considered appropriate for feasibility studies and underground mine design. Laboratory test reports were provided to SRK for quality assurance review. Results of each test provide the data sheet for the test as well as before and after photographs of the samples and loading diagrams. Test results for samples failing along nonconformities in a sample (e.g. microfractures, chips, etc.) were disregarded from the analysis of rock properties. Strength properties were evaluated for mean and standard deviation values in the geotechnical report (SRK, 2017). Mean values have been reported in this document.

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Data from the field drilling program was logged as the core was retrieved. Geotechnical logging consisted of gathering all parameters necessary to estimate RMR and Q values. Quality assurance reviews of the data were conducted on a regular basis in accordance with ISRM logging standards. Rock mass classification values evaluated for mean and standard deviation values for each geotechnical domain and have been documented in the geotechnical report (SRK, 2017). Mean values have been reported in this document.

The geotechnical qualified person has interpreted the methodologies and procedures used to gather strength data and rock mass classification data as sufficient and appropriate for feasibility level design as it impacts mining design and costs. It is assumed that additional geotechnical data will be gathered in a sequential manner as the project is advanced to final design and during operations. This assumption is accordance with industry standards for mining practice.

Source: Nordmin, 2019

Figure 7-8: Plan View Location of 2014-2015 Geotechnical Drill Holes

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Source: Optimize Group, 2022

Figure 7-9: Vertical View of the Location of 2014-2015 Geotechnical Drill Holes Looking Towards the North

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Source: Optimize Group, 2022

Figure 7-10: Vertical View of the Location of 2014-2015 Geotechnical Drill Holes Looking Towards the North with High Grade Niobium Wireframe (>1% Nb₂O₅)

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Table 7-6: Drill Hole Orientation and Data Collection Methods

Year	Hole ID	Easting (m)	Northing (m)	Elev. (m)	Az (°)	Dip (°)	Length	Geotechnical Data
2011	NEC11-001	739297.0	4461224.0	343.4	28.1	-72	900.4		Rock Mass Charact-erization
	NEC11-002	738950.0	4461083.5	343.4	33.5	-61	479.7
	NEC11-003	739417.0	4461059.6	340.8	33.5	-61	508.7
2014	NEC14-006	739166.2	4461224.0	352.0	29.9	-71	772.7	Structure Orientation Data (ATV televiewer)
	NEC14-007	739088.2	4461083.5	344.8	29.4	-71	907.4
	NEC14-008	739128.1	4461159.4	351.2	30.8	-70	886.1
	NEC14-009	739390.2	4461466.2	349.3	208.7	-70	751.3
	NEC14-009a	739390.2	4461466.2	349.3	208.7	-70	897.0
	NEC14-010	739209.5	4461149.8	347.8	30.0	-73	796.1
	NEC14-011	738892.5	4461513.6	359.7	125.8	-65	900.4
	NEC14-012	739635.1	4461083.4	339.9	299.8	-68	843.2
	NEC14-013	739169.3	4461354.3	355.2	-	-90	880.3
	NEC14-014	739034.8	4461218.6	346.3	28.6	-78	901.0
	NEC14-015	739221.0	4461064.7	342.4	29.1	-72	827.8
	NEC14-016	739509.1	4461574.7	354.7	210.5	-60	913.8
	NEC14-020	739037.1	4461305.0	348.4	28.2	-71	587.7
	NEC14-021	739074.3	4461188.0	347.1	29.51	-69	865.0
	NEC14-022	739292.2	4461055.0	340.3	31.3	-68	950.4
	NEC14-023	739377.6	4461071.0	341.5	30.2	-71	615.1
	NEC14-MET-01	739240.4	4461282.7	352.8	-	-90	894.7
	NEC14-MET-02	739171.1	4461372.4	355.8	-	-90	865.0
	NEC14-MET-03	739129.9	4461414.5	355.4	-	-90	913.3
2015	NEC15-002	739046.5	4460708.9	344.4	303.8	-88	850.4
	NEC15-003	740346.2	4460854.1	341.6	306.5	-90	850.5
	NEC15-004	739472.2	4461507.0	354.6	-	-90	413.6	ATV
	NEC15-005	739514.8	4461418.5	351.2	-	-90	407.5

Source: SRK, 2017

Geotechnical Domains

Four spatial geotechnical domains were identified based on lithology, weathering, structural conditions and rock mass strength similarities. These geotechnical domains include:

● Pennsylvania Formation in the upper 200 m.

● Hanging wall material to the southwest of the orebody.

● Mineralized carbonatite orebody.

● Footwall material to the northeast of the orebody.

These domains, shown in Figure 7-11, were delimited based on intact rock properties and in situ rock mass quality from characterization logging. Characterization was based on Rock Mass Rating (RMR76) (Bieniawski, 1976) and the Q-system (Barton et al., 1974). These values were then used with empirical design methods to assess the basic inputs for underground mine design.

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Source: Optimize Group, 2022

Figure 7-11: Geotechnical Model, Vertical Cross Section (N40°E Section)

Structural Faulting

Data on the regional structural geology and the borehole ATV logging data was used to identify 21 major structures in the local mine-scale geology.

Figure 7-12 shows a plan view section through the Vulcan™ model with the position of the stopes and footwall accesses relative to the geologic structures on the +60 m elevation level in Block 1 mining stopes.

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Source: Optimize Group, 2022

Figure 7-12: Plan View of Geologic Structures (Green) on +60 m Elevation Level in Block 1

Rock Mass Quality

The laboratory testing program included 71 Unconfined Compressive Strength (UCS) tests of intact samples, 17 Triaxial Compressive Strength (TCS) tests of intact samples, and 38 Direct Shear Strength (DSS) tests of jointed rock samples. Triaxial tests were conducted at different confinement levels (80 tests) to account for anticipated variations in stress conditions. A set of 29 static and dynamic elastic constant measurements were collected to characterize the elastic properties of the rock. A total of 12 Brazilian Tensile Strength (BTS) tests were conducted to characterize the tensile strength of the rock mass. The UCS results were used to calibrate the spatial variation of UCS values interpreted from 2,180 Point Load Tests (PLT) conducted in the field during the core logging program. The laboratory tests were sufficient to develop shear strength parameters (intact and discontinuity) and provide estimates of the statics and dynamics elastic constants.

The majority of the Rock Quality Designation (RQD) values indicate fair to good rock quality (RQD = 80 to 100) throughout the drill holes. Regions with lower RQD (RQD = 10 to 60) were generally associated with weathered or altered rock zones and/or minor geological intrusions. Fracture frequencies ranged from 0.1 to 9 fractures per metre, with an average of 3 fractures per metre.

The core was generally fresh to slightly weathered with weathering limited to the surfaces of the discontinuities of slight rock mass alteration. Field index strength tests indicate that the core was strong on average (R4). The core tended to break along pre-existing planes of weakness such as veins, foliation and healed structural features.

Laboratory UCS tests results indicate that the carbonatite and lamprophyre strength ranges from strong to very strong (UCS = 50 to 250 MPa), whereas the mafic dikes and mudstone and

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limestone of the Pennsylvania Formation are moderately strong to strong (UCS = 25 to 100 MPa). Table 7-7 shows a summary of the rock mass properties by domain.

Table 7-7: Summary of Rock Mass Characterization by Domain 

Geotechnical Domain	Weathering (%)	Density (t/m3)	IRS (MPa)	RQD (%)	Fracture Frequency (FF/m)	RMR76/ GSI	Q’
I Pennsylvania (13%)	Limestone 48%	2.57	21 – 75	95 - 100	0 – 0.3	73 – 94	17 – 88
	Mudstone 43%		14 – 50	96 – 100	0 – 0.7	68 – 92	14 – 88
II Hanging Wall (22%)	Fresh 49%	2.84	43 – 90	96 – 100	0.5 – 2	59 – 77	6 – 25.1
	Moderated Weathered 41%		28 – 59	81 – 100	1 – 4	51 – 67	4 – 12
	Highly Weathered 10%		8 – 44	17 – 90	2 – 28	33 – 55	0.3 – 5
III Mineralized Carbonatite (50%)	Fresh 72%	3.02	33 – 100	94 – 100	0.3 – 3	57 – 79	6 – 27
	Moderated Weathered 22%		26 – 63	80 – 100	1 – 4	51 – 69	4 – 14
	Highly Weathered 6%		16 – 50	64 – 85	2 – 12	40 -58	1 – 9
IV Footwall (15%)	Fresh 69%	2.84	45 – 113	91 - 100	0.3 – 2	61 - 82	6 – 28
	Moderately Weathered 24%		27 – 59	70 - 90	1 – 5	50 – 70	3 – 14
	Highly Weathered 7%		4 – 28	15 – 85	4 – 26	32 – 48	0.4 – 6

Source: SRK, 2017

The ATV data was used to establish the structural domains and for input to the structural model. The oriented core investigation included a total of 16,790 m of ATV scans of which 9,494 m were sufficient for discontinuity interpretation (i.e., 57% of televiewer success). Structural sets were identified for each domain based on orientation clusters and discontinuity type. Table 7-8 is a summary of the ATV data obtained in each drill hole.

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Table 7-8: Discontinuity Orientation Data for 2014 Geotechnical Investigation 

Drill Hole ID	Drill Hole Length (m)	Total Drill Hole Televiewed (m)	Success of Televiewer (%)	Total Discontinuities Logged
NEC14-006	772.67	499	65%	671
NEC14-007	907.39	723.6	80%	1,132
NEC14-008	886.05	758	86%	908
NEC14-009	751.33	287	38%	754
NEC14-009a	897.03
NEC14-010	796.14	564	71%	1,157
NEC14-011	900.38	807	90%	1,748
NEC14-012	843.23	751	89%	1,305
NEC14-013	880.26
NEC14-014	900.99	716	79%	2,129
NEC14-015	827.84	706	85%	1,691
NEC14-016	913.79	758.4	83%	2,048
NEC14-020	587.7
NEC14-021	865
NEC14-022	949.7	897	94%	3,824
NEC14-023	615.1
NEC14-MET-01	894.74	560.2	63%	1,222
NEC14-MET-02	865.02	820	95%	1,280
NEC14-MET-03	913.33
NEC15-004	413.6	383	93%	1,241
NEC15-005	407.5	264	65%	995

Source: SRK, 2017

The RMR₇₆ values ranged from 50 to 70 in the hanging wall rock, with Barton Q’ values ranging from 4 to 20 with the majority of the rock mass being of fair to good quality. In the footwall, RMR₇₆ values ranged from 50 to 80, with Q’ values ranging from 4 to 30 with the majority of the rock mass being of fair to good quality. The mineralized rock had the greatest RMR₇₆ variations, where values ranged from 60 to 80, and Q’ values ranged from 5 to 25, but overall the rocks are fair to good quality.

Pre-Mining Stress Regime

Since no information on in situ stresses was available in the region and this underground mine would be the first in the district, a stress measurement program was undertaken.

In September 2014, Agapito Associates, Inc. (AAI, 2014) performed downhole in situ horizontal stress testing at the site using the Sigra IST tool. The purpose of the testing program was to estimate the in situ horizontal stress field in the two basic rock types present at the site: the un-mineralized Pennsylvania rock (above 200 m bgs depth) and the mineralized carbonatite ore zone (below 200 m bgs). A total of thirteen tests were conducted; eight of which were successful. The results of the study were as follows:

● There is an apparent increase of stress with the depth of about 36 kPa/m for the major horizontal stress (σ_H) and 21 kPa/m for the minor horizontal stress(σ_h);

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● The major horizontal stress is about 20% greater than the vertical stress (σ_H /σ_V =1.2), and the minor horizontal stress is 71% of the vertical stress (σ_h /σ_V =0.7);

● The average orientation of the major stress is N 66° E. However, a calculation using ATV borehole breakout data provides a better estimate; and

● Figure 7-13 shows a summary of the major and minor stress orientation relative to the fracture set orientations and fault structure orientations.

 

Source: Nordmin, 2019

Figure 7-13: Orientation of Stress Measurements Relative to Faults and Fracture Orientations

Seismicity

A high-level assessment of the local seismic earthquake potential from the International Building Code suggests a local Peak Ground Acceleration (PGA) of 0.02 g for a 50-year return earthquake event. The source of the peak ground acceleration is the 2002 USGS Interactive National Seismic Hazard Map (Frankel et al., 2002). It shows the Maximum Design Earthquake (MDE) with an expected 1% probability of having an earthquake of magnitude greater than 5.0 in 100 years.

No additional studies of seismicity were conducted since the region is not particularly known for large earthquake activity.

Mine Layout Parameters

SRK evaluated different mining methods in their geotechnical analysis. Given constraints on limiting the extent of surface disturbance, long-hole open stoping with backfilling was selected as the optimal mining method for the orebody. To maximize ore extraction, SRK selected a primary/secondary stope extraction sequence, whereby primary stopes are mined first on the first pass and backfilled prior to mining secondary stopes.

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The orebody shape is longest in the northwest-to-southeast direction, which is nearly perpendicular to the major principal stress (s₁). Considering the principal stress orientations, the major geologic structures, and the local discontinuity orientations, the stopes have been oriented to be N 60°E creating the most favourable ground conditions during mining. In this way, the major principal stress is redistributed around the smallest dimension of the primary stopes.

To minimize long-term, mining-induced damage to access drifts, the setback distances used in the design of the mine is 25 m for haulage drives and 75 m for the main ramps. These values represent the minimum distance between the drifts and induced stresses from stope mining. These setback distances were verified from the results of 3D numerical modelling of the mining sequence (SRK, 2017).

Stope Dimensions

A stope stability assessment was completed using the stability graph method (Potvin and Milne, 1992 and Nickson, 1992) with open stope dimensions of:

● Width: 15 m

● Height: 40 m

● Length: 40 m in fresh rock mass (70% of time), 25 m in moderately weathered rock mass (30% of time)

The method compares the hydraulic radius (area divided by perimeter) of a stope face to a stability index number. The stability index number accounts for the rock mass quality (primarily Q values) with adjustments for local fracture orientations, potential block failure mode into the stope, and induced mining stresses. The results from the stability assessment indicated that the initial dimensions for stopes are stable. Figure 7-14 and Figure 7-15 show the stability chart of each stope face under slightly weathered and moderately weathered carbonatite conditions, respectively. The range of stability numbers shown for each stope face consider changes in depth and variations in rock mass quality within each geotechnical domain.

The difference between these stope dimension from the 2017 Feasibility Study is that the stope height is taller by 10 m, but the stope lengths are shorter by 10 m. the result is a very similar hydraulic radius for the critical sidewall stability. The narrow stope widths, oriented along the major principal stress, are the same as in the 2017 Feasibility Study, so the taller stopes have minimal stability impact for the hanging wall and footwall.

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Source: SRK, adopted from Potvin, 2001

Figure 7-14: Empirical Stope Design Chart for Moderately Weathered Rock Mass 

(Red=Footwall Face, Blue=Hanging Wall Face, Magenta=Back, Yellow=Sidewall Faces)

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Source: SRK, adopted from Potvin, 2001

Figure 7-15: Empirical Stope Design Chart for Fresh and Slightly Weathered Rock Mass 

(Red=Footwall Face, Blue=Hanging Wall Face, Magenta=Back, Yellow=Sidewall Faces)

A second stability assessment was completed by using larger stope dimensions (40 m length) assuming 15% dilution along the walls. Results of this analysis indicate that these diluted stopes should be stable but may require some support depending on rock mass quality. The actual in situ rock mass conditions need to be incorporated in the final stope designs.

Stope stability was simulated using a 3D numerical model for the 2017 Feasibility Study stoping sequence and dimensions (SRK, 2017). Although the design stope dimensions and mining sequence have been changes since these analyses were conducted, the results are considered generally applicable because the hydraulic radius and net extraction ratios are quite similar. The modelling results confirm that stopes and access drifts are predicted to remain stable during active mining. A2GC recommends that the numerical analyses should be re-assessed as the project design stage is advanced.

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Backfill Strength

An initial backfill strength assessment was completed for the primary/secondary stopes. This assessment identified the minimum strength required for the primary panels to be self-supporting when they are exposed on the east and west walls. Two methods were used to determine the minimum backfill strengths: 2D analytic vertical stope method and numerical modelling using FLAC3D stope-scale model (SKR, 2017), although the current stope sizes are slightly different.

The analytical methods included Li and Aubertin (2014) and Belem and Benzaazoua (2000) for single stope face of backfill exposed during secondary stope mining (exposing two backfill faces simultaneously is not planned). Using a factor of safety of 2.0, a 14-day UCS strength of 1.0 MPa was considered the minimum strength for the primary stope panels. The numerical modelling results confirmed these values (SRK, 2017). The model was also used to verify that secondary stopes would remain stable in that no “sit-down” failure mechanism would occur against two adjacent backfilled stopes during the progressive mining of the secondary stope.

Dilution

Dilution was estimated using the method developed by Clark and Pakalnis (1997) based on an empirical model calibrated to case histories for dilution into open stopes. The method predicts the quantity of unstable wall rock for a given rock mass quality from a given stope size. The parameters plotted on the dilution chart are the stability number, N’, and hydraulic radius.

The thickness of external dilution is estimated as an ELOS. Figure 7-16 shows the range of predicted equivalent linear overbreak/slough (ELOS) for moderately weathered carbonatite, and for fresh to slightly weathered carbonatite. Sidewall and back dilution are not expected to be a problem because in the primary stopes dilution (from secondary stopes) will be at grade, and dilution from secondary stopes is managed by controlling backfill strength. Given the low fracture frequencies (about 3 m / fracture) dilution is anticipated to be variable.

For stope sidewalls, the dilution depends on stope length and depth below ground and is summarized as follows.

● 25 m long stopes in moderately weathered rock mass dilution from HW & FW = 0.25 m (blue)

● 25 m long stopes in moderately weathered rock mass dilution from Sidewalls = 0.25 m in shallow stopes increasing to about 2 m in deepest stopes (green)

● 40 m long stopes in fresh (unweathered) rock mass dilution from HW & FW = 0.10 m (red)

● 40 m long stopes in fresh (unweathered) rock mass dilution from HW & FW = 1 m in shallow stopes increasing to about 2 m in deepest stopes (yellow)

For design purposes, an average of 0.25 m of ELOS dilution has been assumed for endwalls 0.75 m for sidewalls. In backfilled sidewalls, an average of 0.2 m has been assumed.

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Source: SRK, 2017, adapted from Clark & Pakalnis, 1997

Figure 7-16: Empirical ELOS Estimate – Moderately Weathered and Fresh Rock Mass

Ground Support

Ground support requirements have been estimated using empirical support charts developed by Barton (1974). The method relates the rock mass quality (Q) to the equivalent dimension of the excavation (De). De is the ratio of the excavation width (D) to the excavation support ratio (ESR) index. The ESR value accounts for a degree of safety required depending on the use of the excavation. Values range from ESR=1.6 for long-term critical access drifts to 2.5 for short-term temporary accesses into stopes. The recommended ground support consists of three general classes of support levels based on rock mass quality, drift usage and drift dimensions. The three-class system includes:

● Support Type 1 – spot bolting for Q>4 (~69% of drifts);

● Support Type 2 – systematic bolting and 4 to 10 cm of fibre reinforced shotcrete for Q<4 and Q>1 (~24% of drifts); and

● Support Type 3 – systematic bolting, steel mesh, and 4 to 10 cm of plain shotcrete for Q<1 (~7% of drifts).

Table 7-9 summarizes the drift dimensions used to estimate ground support requirements. The support specifications are summarized in Table 7-10.

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Table 7-9: Barton Parameters for Different Excavations

Excavation	Type of Excavation	Opening Dimensions W x H (m)	ESR	D	De
Main Ramps	Long Term (6-10 years)	5.5 x 5.5	1.6	5.5	3.4
Footwall Accesses	Medium Term (1 year)	5.0 x 5.0	2.0	5.0	2.5
Stope Accesses	Short Term (1-2 months)	4.5 x 4.5	2.5	4.5	1.8

Source: SRK, 2017

Table 7-10: Preliminary Support According to Barton Method

Geotechnical Zone	Q	Excavation	Support Categories	Bolt Length	Bolt Spacing	Other Support
Footwall, highly weathered (6%)	0.4 to 6.2 (Very Poor)	Main Ramp	3-Bolts, mesh and shotcrete	2.5 m	1.2 m	Fully grouted rebar, mesh, 5 cm shotcrete
		FW Access	2-Systematic bolting	2.5 m	1.2 m	Split sets and mesh
		Stope Access	2-Systematic bolting	2.5 m	1.6 m	Split sets and mesh
Footwall, moderately weathered (24%)	3.2 to 13.8 (Poor-Fair)	Main Ramp	2-Systematic bolting	2.5 m	1.2 m	Fully grouted rebar, mesh
		FW Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets
		Stope Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets
Footwall, slightly weathered (70%)	5.9 to 28.1 (Fair-Good)	Main Ramp	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Grouted rebar
		FW Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets
		Stope Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets

(%) Amount of Expected Ground 

Source: SRK, 2017

Smaller drifts in the competent ground may use short length bolts (i.e., 1.8 m long versus 2.5 m) in the good quality ground, depending on ground conditions. Also, in short-term accesses in the good quality ground, it may be possible to substitute swellex or split set bolts, depending on ground conditions. The need for cable bolts and/or shotcrete in stope brows will be dependent on ground conditions and the degree of fracturing in the brow area. The final decision on these substitutions will be up to the on-site geotechnical ground support engineer.

Crown Pillar Stability

Crown pillar stability assessment was conducted using the empirical Scaled Span Method (Carter, 1992) and a limit equilibrium analysis in the CPillar software. The results of the empirical analysis indicate that if the rock mass quality in the crown has a Q value greater than 37, then all stopes

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could be mined in Level 1 with a Factor of Safety (FOS) of more than 2.0. No additional analyses are required at this time because the stopes are planned to be tightly backfilled and the FOS will be much higher.

The results of the 3D numerical model (SRK, 2017) confirm that the crown pillar will remain stable during and after mining. The total predicted surface displacement is anticipated to be less than about 1 cm at the end of mining.

7.4 Hydrogeology Design Parameters

The hydrogeology of the deposit was characterized based on three phases of work:

1. Phase I: The first phase of hydrogeological characterization was conducted during phases 1 and 2 of the core drilling program and consisted of packer testing, installation of piezometers, and measurement of water levels. Specifically, the program included:

● 42 downhole packer-isolated injection and airlift tests in drill holes.

● Installation of six 2“ PVC standpipe piezometers isolated in the carbonatite and open to large intervals of the deposit.

● Installation of two nominal 2“ PVC standpipe piezometers isolated in the 180 m thick Pennsylvanian aquitard above the carbonatite.

● Frequent measurement of water levels in open drill holes and piezometers over a period of six months.

2. Phase 2: Following the second phase of resource-related core drilling, a 10-day airlift pumping test was completed using a deep, open, vertical PQ drill hole as a pumping well. Water levels from the surrounding piezometers were recorded over the duration of the test and for several weeks following the test.

3. Phase 3: The third phase of hydrogeological characterization involved installation of two multi-level, distally-located piezometers and a deep 6” diameter injection well completed to depths of 850 m, followed by the performance of a nominal 30-day injection test. The piezometers were completed within the carbonatite at distances of 0.6 km and 1.2 km from the center of the injection well, which was located at the center of the orebody. The injection test was chosen as a test method over a standard pumping test due to the salinity of the groundwater and the expense of handling the discharge water. During the injection test, surface water from Todd Creek was injected at rates of between 22 L/s and 30 L/s (350 to 480 gpm) over a period of 33 days, including downtime. Response to the injection test was monitored over the duration of the test and for more than eight weeks following the test.

Groundwater quality data was collected by NioCorp using sampling procedures developed by SRK and NioCorp. These procedures included prescriptive methods for sample collection, preservation, chain of custody and transport to a local commercial laboratory (Midwest Labs, Omaha, NE). The laboratory maintains ISO/IEC 17025:2017 certification and incorporates appropriate QA/QC procedures in their analysis. ABC’s opinion is that the sampling and analytical procedures for groundwater are adequate to appropriately characterize the water quality of the groundwater systems present at the Elk Creek site.

Groundwater hydrogeological characterization data collected in the three phases described above was completed by SRK, NioCorp and the contract drilling companies present at the project site in 2014 and 2015. The data collection was conducted by professional hydrogeologist and geologists in accordance with well established procedures. No formal QA/QC program was included in data

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collection effort. ABC’s opinion is that the data collection methods were appropriate for the groundwater system present at the site and the data is adequate to support hydrogeologic modelling.

7.4.1 Conceptual Geohydrology

Geology

The Elk Creek Deposit is hosted in the Elk Creek Carbonatite, a volcanic carbonatite plug located in south-east Nebraska. The carbonatite plug is 6 km to 8 km in diameter and contains the orebody at its approximate center. Surrounding the plug is low permeability Precambrian-aged granite bedrock (see Figure 7-17).

The local geology generally consists of a 30 m thick layer of low permeability Pleistocene-aged glacial till overlying a 180 m thick low-permeability Pennsylvanian-aged shale and limestone, which rests on top of a moderate-permeability Phanerozoic-aged carbonatite volcanic plug extending to a depth in excess of 1,000 m (see Figure 7-18).

Glacial Till

Pleistocene-aged glacial till covers the surface of the site, to a depth of approximately 30 m. It is variably permeable, with lenticular glacial outwash features providing potable water to shallow wells that service local agricultural activities. Water levels in these wells are typically within 10 m of the ground surface.

Pennsylvanian Sediments

The Pennsylvanian sediments are made up of interbedded shale and limestone strata. The ensemble displays horizontal hydraulic conductivity on the order of 10^-9 m per second in localized testing from single wells, with the vertical hydraulic conductivity of the horizontally layered strata likely 100 times lower. This unit hydraulically isolates the glacial till groundwater system from the deeper carbonatite and effectively functions as an aquiclude for vertical water infiltration to the carbonatite below.

Water levels in wells completed in this unit are typically 50 m below ground surface, indicating a vertical downward head gradient from the glacial till above to the carbonatite below. However, due to the very low vertical permeability of the sediments, there is essentially no vertical downward groundwater flow through them.

Phanerozoic Carbonatite

The Phanerozoic carbonatite unit is made up of carbonatite (volcanic calcium-magnesium-iron carbonate) with siliceous lamprophyre dikes and sills interspersed throughout. The intact carbonatite and lamprophyre rocks are essentially impermeable, and the rock mass is generally lightly fractured, resulting in generally low hydraulic conductivity. However, the plug is intersected by fractured zones and dissolution structures that are interpreted to be related to faulting that likely occurred at the time of emplacement of the carbonatite. As a result, the total orebody material is of moderate hydraulic conductivity, on the order of 10^-6 m per second (based on the multiple well pumping and injection tests), ranging locally from 10^-9 to 10^-5 m per second (based on the results of single hole testing, shown in Figure 7-19).

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Source: Adrian Brown Consultants, 2019

Figure 7-17: Hydrogeologic Plan

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Source: Adrian Brown Consultants, 2019

Figure 7-18: Hydrogeologic Section

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Source: Adrian Brown Consultants, 2019

Figure 7-19: Permeability Testing of the Elk Creek Rock Mass

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The nature of the fracturing and dissolution of the carbonatite changes with depth in the orebody, which also causes variability of permeability with depth, as follows:

1. Contact Zone: The uppermost 90 m of the carbonatite show heavy fracturing and occasional dissolution cavities, indicating that it has been subject to long-term leaching by fresh groundwater from above, likely occurring prior to the emplacement of the overlying Pennsylvanian sediments. The ubiquitous presence of this zone in all drill holes penetrating the carbonatite and the appearance of this rock in core suggests it forms a thin dissolution caprock on the entire carbonatite plug, similar to caprocks seen at the top of salt domes. The permeability of this zone averages 4 x 10^-7 m per second but is highly variable.

2. Mining Zone: The orebody from approximately 90 m to approximately 600 m below the top of the carbonatite is made up of largely unfractured carbonatite and other volcanics, intersected with rubble zones, probably caused by faulting, with some associated evidence of dissolution on joints and fractures. These broken rock zones impart an elevated hydraulic conductivity to some locations in the orebody, which collectively cause this portion of the rock mass to have an average hydraulic conductivity in the order of 2 x 10^-6 m per second. The local hydraulic conductivity is also highly variable, with values up to 2.5 x 10^-5 m per second.

Within all zones, the permeability of the carbonatite appears to be locally quasi-isotropic and relatively homogenous, based on the relatively uniform head increase cone observable in the water level data during the large-scale injection test (see Figure 7-20).

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Source: Adrian Brown Consultants, 2019

Figure 7-20: Response to Injection in Carbonatite - End of Test

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Water in the carbonatite is saline. When sampled from wells unaffected by injection of fresh water by drilling or testing, the carbonatite groundwater has a total dissolved solids concentration of approximately 19,000 mg/L, and a sodium chloride concentration of approximately 16,500 mg/L, comprising 89% of the total dissolved solids concentration (the balance is mostly calcium, magnesium, and carbonate). The high salinity of the water in the carbonatite cannot have originated from the overlying materials, which are essentially devoid of sodium chloride. Accordingly, the sodium chloride must either have come from dilution of connate seawater (which has a sodium chloride concentration of 35,000 mg/L), or dissolution of evaporates located in the carbonatite rock mass. Given the volcanic origin of the carbonatite, seawater is considered to be the more likely genesis. This indicates that the water in the carbonatite is a captured volume of seawater and has been diluted over geologic time by slow infiltration of fresh water through the overlying Pennsylvanian strata, with the resulting dense water very slowly discharging from the carbonatite at great depth through the surrounding Precambrian granite.

Water levels in wells and piezometers completed in the carbonatite unit are currently approximately 100 m below ground surface. This water level is approximately equal to the water level in the Missouri River, 52 km to the east, which demonstrates that the water level in the carbonatite is not controlled by regional drainage to the river. The reasons that the water level is depressed below the level expected to exist in an open groundwater flow system may include the following:

1. Salinity. The water in the carbonatite is saline, approximately half the sodium chloride concentration of the ocean, and is approximately 0.1% denser than pure water. This column of water in the carbonatite balances the water pressure in the granite outside, which it is reasonable to expect is in general non-saline. As a result, at any point above the point where saline carbonatite water seeps into the surrounding granite, the static (saline) water level in the carbonatite will be less than the static water level in the granite. In the case of the Elk Creek Carbonatite, the water level in the carbonatite would be expected for this reason alone to be approximately 5 m to 10 m lower than the water level in a pure water system, such as the overlying till.

2. Isolation. If the carbonatite is an isolated system, its current water level will not bear any set relation to the water levels in adjacent materials. The injection test performed as part of this project injected a total of 68,588 m³ of (fresh) water into the carbonatite, which produced a long term – apparently permanent – increase in the water level in the entire 40 km² carbonatite plug of approximately 1 m (see Figure 7-20). The storativity displayed by the rock mass is 0.003, which is typical for a thick confined fractured rock aquifer. This permanent head-change behaviour is indicative of an effectively isolated system, in which the water level that is currently observed is the result of all the inputs and outputs of water to the carbonatite over geologic time.

3. Isostatic Uplift. The carbonatite lies within the area of late Winsconinan glaciation and was most recently covered by between 2 km and 4 km of the glacial ice sheet between 10,000 and 25,000 years ago. This ice sheet acted as a load on the carbonatite system, compressing it with an added total stress of approximately 20 to 40 Mega-Pascal (MPa). When the glacial sheet ice melted, this load was removed quite rapidly in geological terms, causing a large reduction in total stress in the carbonatite. As the carbonatite is isolated, this reduction in total stress resulted in a corresponding reduction in the porewater pressure in the carbonatite, reducing the water head in a hypothetical piezometer in the carbonatite by up to between 2,000 m and 4,000 m. The porewater pressure in the carbonatite and the water level along with it has been slowly recovering since, by drawing water from the surrounding

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Precambrian granite and overlying Pennsylvanian sediments into the carbonatite. It appears that when the current investigation program began, the water level had recovered to within 70 m of the equilibrium level, which is represented by the current water level in the glacial till.

4. Investigation Extraction. The investigation of the carbonatite has injected water into and withdrawn water from the carbonatite. As was seen in the injection test, this can permanently influence the water level in the carbonatite, as it is an effectively closed system. In the case of the current program, the drilling has generally used coring technology, which involves the injection of water to keep the bit cool and to remove cuttings. Lost circulation has been a persistent issue, so the drilling has if anything caused a net injection of water into, and an increase in water level in, the carbonatite. The only deliberate withdrawal from the carbonatite occurred during the 10-day extraction test, which extracted 1,900 m³ of water. Based on the injection test result, this would have reduced the water level in the carbonatite by a net 0.03 m (3 cm).

In conclusion, the carbonatite is an isolated, fractured volcanic plug, surrounded and covered by essentially impermeable materials. It has a moderate hydraulic conductivity at the top of the plug, decreasing with depth due to reduced fracturing and dissolution. The head conditions and the large-scale testing of this material show that it is hydraulically isolated, and the water pressure is lithostatically controlled, with effectively no exit for water from the carbonatite to any other material.

Precambrian Granite

Precambrian granite surrounds the carbonatite and acts as a containment structure for the water within it. The granite has a hydraulic conductivity of less than 10^-4 m per day based on the results of modelling the long-term injection test. The granite provides very limited opportunity for groundwater flow laterally to or from the carbonatite, consistent with the above conclusion about the genesis of the brine contained within the carbonatite rock mass, the observed static water level in the carbonatite, and the permanent water level changes induced by injection testing at the site.

7.4.2 Mine Inflow Control

Concept

Mine inflow control will be achieved in the Elk Creek Mine by limiting groundwater inflow to the mine to a maximum flow of 63 L/s (1,000 US gpm). This will be achieved by reducing the permeability of the materials around mine openings, by freezing in the upper glacial and sedimentary strata, and grouting the lower fractured rock carbonatite strata (which contain the orebody).

The resulting mine inflow will be pumped to the surface for treatment, with the filtrate used or discharged, and recovered solids placed in lined storage facilities.

Inflow Control

The material to be mined is located entirely within the Phanerozoic carbonatite unit, and based on the mine design in this report comprises a block of material with the following approximate dimensions:

Length (L) = 500 m 

Width (W) = 200 m

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Depth (D) = 600 m

The block is surrounded by brine with a current average piezometric surface at elevation 250 m amsl, which will exert an average hydraulic head on the sides of the mined block as follows:

Hydraulic head (H) = 445 m amsl

The block will be grouted before mining, and completed with cemented backfill after mining, so inflow can be estimated as flow from the outer surfaces of the mining block to the inside of the block, taken to be the center of the block.

Using Darcy’s Law:

Q = K I A and K = Q / (I A)

where:                Q = inflow from each side of block = 0.0315 m³/s 

K = hydraulic conductivity (to be determined) (m/s) 

I = hydraulic gradient = H / (0.5 W) = (445 m) / (0.5 * 200 m) = 4.45 

A = flow area = D * L = 600 m * 500 m = 300,000 m²

produces the following estimate for the required hydraulic conductivity in the block to limit the overall inflow to 63 L/s:

K = Q / (I A) = (0.0315 m³s^-1) / (4.45 * 300,000 m²) = 2x10^-8 m/s

The average hydraulic conductivity of the carbonatite in the vicinity of the orebody was found by pumping tests to be in the order of 2x10^-6 m/s (SRK, 2017), approximately two orders of magnitude higher than the hydraulic conductivity required to restrict the mine inflow to the desired rate. Further, 35 single hole tests of the carbonatite were performed in the carbonatite, of which 30 (85%) showed hydraulic conductivity greater than 2x10^-8 m/s (Figure 7-19 from this report). By both metrics, each mined opening in the orebody will need to have the permeability of the surrounding rock reduced to below 2x10^-8 m/s before it can be extracted.

The permeability of concrete made up of ~10% cement by weight is approximately 2x10^-8 m/s (Whiting, 1988).¹. It is therefore reasonable to assume that grouted rock will also have this – or a considerably lower – hydraulic conductivity, as intact rock will in this case will constitute around 98-99% of the mass, and the low permeability grout will fill the other 1-2%, creating a very low permeability material.

Groutability

Groutability in fractured rock depends on the ability to deliver grout to the fractures and voids in the rock, and thereby to largely eliminate the ability of those conduits to convey brine to the mine

¹ Whiting, David, 1988. “Permeability of Selected Concretes,” Permeability of Concrete, SP-108, American Concrete Institute, Detroit, Michigan. 

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workings. Grout is a suspension of cementitious materials (typically including some amounts of portland cement, fly ash, blast furnace slag, and sometimes bentonite) in water which requires significant velocity in the delivery pipe to prevent separation, and to allow effective filling of the receiving flow conduits.

For long, small diameter grout holes, experience has shown that a flow rate in the order of 1 L/s is required to emplace grout, particularly for holes that are horizontal or sloping upward from the collar (which will be typically the case for this mine). In this mine application, it would be expected that grouting would be conducted under the following conditions:

Length of grout hole open: ~100 m 

Applied injection pressure: ~100 m (1 MPa) in excess of ambient 

Required flow: >1 L/s at the collar.

Consider the rockmass hydraulic conductivity that is required to achieve that flow. To good approximation, for grouting via long, slender grout holes, the flow equation is (Theim, 1870)²:

Q ≈ L K H and K = Q / (L H)

where:                Q = grout injection flow (0.001 m³/s minimum) 

L = length of borehole open to grout injection (~100 m) 

K = hydraulic conductivity (to be determined) (m/s) 

H = excess injection pressure (~100 m)

Applying the parameters to the equation, assuming the viscosity of the grout is similar to that of water, in this orebody a hole will be groutable if the hydraulic conductivity (K) of the rockmass is greater than:

K > Q / (L H) = (0.001 m³/s) / (100 m * 100 m) = 1x10^-7 m/s

Based on the results of the injection testing performed in carbonatite for this project (Figure 13-12), 60% were groutable (displaying a hydraulic conductivity in excess of 1x10^-7 m/s). Of the rest, 15% did not require grouting because they displayed a hydraulic conductivity of less than 2x10^-8 m/s. The remaining 25% displayed hydraulic conductivity values between 2x10^-8 m/s and 1x10^-7 m/s, and may not be groutable, but may not need to be grouted as they would contribute only a small flow to the mine if ungrouted. In conclusion, by this measure, most or all of the mine openings in the carbonatite can be grouted, or don’t need to be grouted.

A second consideration for groutability is the ability to force the cementitious particles in the grout slurry into and through the flow conduits to plug them. There is currently no data on fracture opening size for this orebody, so this issue will remain for demonstration prior to implementation. In any event, it will be prudent for grouting to be performed using high-shear superfine grout, comprising blast furnace slag, fly ash, and ultrafine cementitious grout. This grout has a D₉₅ size of 10 microns, and a D₅₀ size of 3 microns. These diameters are smaller than the

² Thiem, A., 1870. The yield of artesian wells, dug wells and filter galleries. J. Gas Lighting Water Supply (Munich), vol.14 

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apertures of most significant water-carrying factures in carbonate rocks. Given the likelihood that the major water conduits along fractures and fault material in the carbonatite have been subject to dissolution and karstification, grout should be readily forced into the significant water pathways.

Grout Take

The quantity of grout that will be required to plug the joints and faults in the carbonatite depends on the porosity of the material. This porosity is likely to be present as primary porosity (porosity of the unfractured material) and secondary porosity (porosity of the fractures and faults). In rocks not all porosity is effective; effective porosity is that portion of the porosity that is connected and provides conduits for flow through the rock. The effective porosity of the carbonatite has been estimated based on the storage characteristics displayed by the formation during the long-term injection test (SRK, 2017). Two approaches have been taken.

1. Storage coefficient.

The storage coefficient is a measure of the volume of water that is required to be injected to raise the water level in one square meter of the carbonatite in the test by one meter. The storage coefficient of the carbonatite identified in the test was 0.0128. This could be taken to mean that the effective porosity of the formation is 1.3%. As the top of the carbonatite remained saturated during the test, this may not be a reliable measure of the porosity of carbonatite, but it does seem likely that it is not less than this value.

2. Specific storage.

The specific storage of the carbonatite is the volume of water required to be injected into a cubic meter of the carbonatite to cause the water pressure to increase by one meter. It is effectively a measure of the compressibility of the void spaces in the carbonatite material. In the long-term injection test the Specific Storage (S_S) was determined to be (SRK, 2017):

Specific Storage (S_S) = 1.8x10-5 m^-1

The definition of (rock) compressibility is the change of volume of rock per unit volume of rock per unit change in pressure. Combining this with the definition of specific storage, and noting that porosity is equal to the volume of voids per unit volume of rock, the effective porosity can be approximated as:

 n_e ≈ C_R / S_S 

where:               n_e = effective porosity

C_R = rock compressibility (m^-1)

S_S = specific storage (m^-1)

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The compressibility of intact silica is 3x10^-7 m^-1 (Domenico and Mifflin, 1965)³. Using this value, and the value of specific storage found in the injection test results in the following effective porosity estimate:

n_e = 3x10^-7 m^-1 / 1.8x10^-5 m^-1 = 0.016 or 1.6%

This would seem likely to be equivalent to the amount of water that would be drainable from the rock. The compressibility of intact carbonate rock (e.g. calcium carbonate or dolomite) would be expected to be somewhat higher than at of silica, but we have no data for this carbonatite at this time.

In conclusion, it is reasonable to expect that the effective porosity of the carbonatite is in the order of 1.3% - 2%. As only 60% of the orebody will accept grout, it is concluded that the grout take of the orebody carbonatite will be in the order of 1% by volume. This value has been used in computations of the grout requirements and cost for this project.

Shaft Inflow Control

Groundwater inflow to the shafts and associated breakouts will be controlled by freezing the upper portion of the shafts in the glacial and Pennsylvanian sediments and grouting of the rock in advance of shaft sinking. A sealed concrete pressure liner will be emplaced in the shaft progressively during sinking to control inflow after completion of the shaft sinking and cessation of freezing.

Breakout stations will be grouted from within the shaft liner to control inflow during development and may be lined or dentally grouted to complete flow control once accessed.

Inflow to the shaft and breakout stations is expected to be nominal, with a peak estimated at 10 L/s (150 US gpm). If the flow exceeds this value, additional freezing or dental grouting will be employed.

Development Inflow Control

Inflow to all permanent development drifts, transport haulage ways, ramps, and other non-production underground facilities will be controlled by grouting significant inflow conduits (generally fractured and faulted zones within the orebody that contain rubble and dissolution pathways). The method of grouting will be developed during the initial excavation of the mine, and is expected to be generally as follows:

1. Prior to development of the underground facility, cover holes will be drilled from the access point at a spacing of approximately 5 m (15 ft) and grouted with superfine cement grout to a pressure equal to 150% of the static water pressure originally computed for that location, to create a low-permeability envelope around the facility advance.

³ Domenico, P. A.; Mifflin, M. D. (1965). "Water from low permeability sediments and land subsidence". Water Resources Research 1 (4): 563–576.

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2. Following excavation of the facility or drift, all locations where there is visible inflow to the mine in excess of 1 L/s per 100 m of facility length will be sealed by dental grouting, using shear-activated grout injected at a pressure up to 150% of the observed water pressure in the area to be grouted.

Stope Inflow Control

Inflow to all mined stopes will be controlled during mining by grouting any major inflow conduits (generally faulted zones within the orebody that contain rubble and dissolution pathways). After mining is complete, inflow will be controlled by sealing the stopes using cemented paste backfill. The method of mine inflow control to be used for stopes will be developed during the mining process, and is expected to be generally as follows:

1. Prior to the development of a stope, a drill hole will be advanced along the stope centerline to the distal end of the stope.

2. In the event that the free-flow from the advance drill hole exceeds 3 L/s (50 gpm), the advance hole will be pressure grouted with ultrafine neat cement grout at a pressure up to 200% of the calculated pre-mining static water pressure at that location to control inflow during the mining of the ore in the stope.

3. After mining is completed, the stope will be completely backfilled with cemented paste, creating a low permeability inclusion in the stope to reduce flow from the stope to less than 0.5 L/s (10 gpm). If the outflow from the stope to the mine workings exceeds 0.5 L/s, a concrete bulkhead will be constructed in the access drift to the stope with a pressure rating equal to 150% of the calculated pre-mining static water pressure at that location. Any remaining leakage past the bulkhead will be controlled by dental grouting of the rock surrounding the bulkhead.

Impact of Mine Inflow Control on Groundwater Pressure in the Carbonatite

The groundwater control system proposed for the Elk Creek Mine will result in the extraction of up to 66 L/s from the isolated carbonatite volcanic plug. This will have the effect of reducing the water head pressure in the carbonatite plug at a rate of up to 30 m per year, which if sustained will dewater orebody down to the base of the proposed mine by the end of mining.

This dewatering has an impact on the optimal order of mine development, favouring early development in lower permeability deep portions of the orebody, and later mining of higher-permeability portions of the orebody when the water pressure in them has been reduced by pumping the prior mine inflow. The current mine plan takes advantage of this opportunity, at the same time as targeting the higher-grade ore, which is generally in the deeper portion of the mine, for early extraction.

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8. SAMPLE PREPARATION, ANALYSES, AND SECURITY

The following section summarizes the sampling methodologies used by the Company and Molycorp drilling operations. Workflows described in this section include historical Molycorp procedures (1973-1986); Quantum historical re-sampling and drilling procedures (2010-2011), and NioCorp’s historical re-sampling and drilling procedures (2014-2021). Analytical procedures and evaluations have been expanded from previous reports to include Rare Earth Elements (REE), in addition to the previously present niobium, titanium, and scandium results (Nordmin et al, 2019 & SRK, 2017).

8.1 Sample Preparation and Security

8.1.1 Molycorp, 1973 – 1986

Detailed descriptions of Molycorp’s sample procedures, analyses and security have not been documented and reviewed directly by the QP’s. However, given the detailed nature of the historic drill logs and reports for the individual drillholes, and Molycorp’s position as a leader in the rare earth industry at the time, it is considered likely that Molycorp applied the same standards to their sampling procedures.

A review of previous reports prepared under the Canadian NI 43-101 standard and details collected from discussions with the local Molycorp sampling technician defined the following procedures:

● The drill core was photographed.

● Samples were derived from 1.52 m (5 ft) or 3.05 m (10 ft) intervals of hydraulically split dominantly NQ diameter size core, with minor BQ diameter, that was crushed on site, before sending samples to the lab (the crusher is no longer on site).

○ The core crusher was cleaned between samples by using limestone blank material.

○ Only NQ diameter drill core was identified and relogged in with the defined Resource area.

● Sample homogenization methods were not clearly defined in historical records.

● Drill core samples were sent to Molycorp’s exploration laboratory at Louviers, Colorado for niobium and LnO (lanthanide oxide) analysis, without the individual REEs being reported.

● Molycorp drill information was used to support the geological model and validated historical niobium results were used in the Resource.

● Select Molycorp samples were re-assayed in 2010 (by Quantum), 2014, 2016, and 2021, with database expanded to include elements that were not historically assayed.

Complete details of the sampling procedures and details around any changes of the procedures over the period of the drill programs remain unclear. Photographs of the core were not included with Molycorp’s available historic records.

Molycorp built two insulated, steel buildings, located on the Property of Ms. Elda Beethe, within 100 m of the known deposit. The buildings were ceded to Ms. Beethe when Molycorp abandoned the Project and ownership has been transferred to NioCorp.

Drill core samples collected were sent to Molycorp’s exploration laboratory at Louviers, Colorado for niobium and LnO analysis. The analytical methods are described in an internal memo by Sisneros and Yernberg, 1983, where “…Niobium was analyzed by wavelength dispersive XRF on

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pressed powder pellets, following pulverization to -325 mesh. Molycorp did include some quality control methods. Standardization was provided by using a variety of Elk Creek samples, which had been analyzed by alternative methods at other internal Molycorp laboratory facilities. Over the project duration, the number and/or identification of the standards used changed several times. In 1981, the instrumentation changed from a Philips PW1212 to a PW1400.” (Sisernos and Yernberg, 1983)

The assay tables from some of the holes (EC-27 and EC-30) indicate a ‘tentative test’ (XRF) of niobium value from Louviers laboratory, and a ‘commercial lab test” (XRF) of niobium values. It is unclear which commercial laboratory conducted these tests, although the 1983 Niobium Analytical Standardization report mentions that the Molycorp exploration department occasionally utilized Bondar-Clegg. Notes on the assay tables indicate that the commercial laboratory utilized one standard (from hole EC-11) for its XRF analysis, whereas Louviers utilized 19 standards from hole EC-11.

The drill-core crushed (coarse reject), and pulverized material are currently being stored at a facility managed by the University of Nebraska-Lincoln (UNL). This facility is located approximately 8.5 km south of the town of Mead, Nebraska, and approximately 63 km northeast of Lincoln, Nebraska. The core was stored at two other storage facilities on UNL property, prior to its current location. Prior to the acquisition of the core by UNL in the late 1990s, the core was stored in the steel sheds on the property of Elda Beethe.

Dahrouge and NioCorp completed multiple site visits to the Mead Core facility, monitoring the core, coarse-split and pulverized split sample storage, and organization. These facilities are kept secured and in high quality condition by UNL.

In 2014 a core logging validation was completed by NioCorp which concentrated on the 27 historical drillholes within the resource area. Table 8-1 is an inventory of core at the Mead facility filtered to include only the 26 drillholes within the resource area.

Table 8-1: Core Inventory of Drillholes within the Resource Area at the Mead Facility

Hole ID	Core Box Intervals		Depth		Drill Core review
	Box # From	Box # To	From (m)	To (m)
EC-11	7	41	207.6	310.3	2014 relog Validation
EC-11A	1	180	233.2	769.6	2014 relog Validation
EC-14	13	188	43.9	707.1	2014 relog Validation
EC-15	30	244	215.8	839.7	2014 relog Validation
EC-16	7	218	214.6	817.5	2014 relog Validation
EC-18	9	102	189.6	462.4	Material review & standardization
EC-19	9	178	194.2	664.2	2014 relog Validation
EC-20	6	189	190.5	739	2014 relog Validation
EC-21	6	156	210.3	644.3	2014 relog Validation
EC-22	12	193	207	733.3	2014 relog Validation
EC-24	11	39	191.7	281.9	Material review & standardization
EC-25	9	47	192.9	304.5	Material review & standardization
EC-26	14	191	199.3	733	2014 relog Validation
EC-27 & 27A	13	186	202.4	702	2014 relog Validation
EC-28	16	209	193.5	769.6	Material review & standardization
EC-29	14	182	196.9	726	2014 relog Validation

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Hole ID	Core Box Intervals		Depth		Drill Core review
	Box # From	Box # To	From (m)	To (m)
EC-30	9	201	182.9	757.1	2014 relog Validation
EC-31	16	117	203.3	512.4	2014 relog Validation
EC-32	13	165	196	681.2	Material review & standardization
EC-33	14	83	199	405.4	Material review & standardization
EC-34	16	69	202.4	362.7	Material review & standardization
EC-35	6	27	192	260	Material review & standardization
EC-36	8	95	214	474	Material review & standardization
EC-37	14	87	239.9	457.5	Material review & standardization
EC-51	1	89	220.1	470.6	Material review & standardization
EC-54	7	97	213.7	464.5	Material review & standardization

Source: Dahrouge, 2014

An investigation of the drillholes within the resource area led to the following conclusions:

● Core boxes were generally in good condition and labeled well;

● Not all of the historical core was transferred to the Mead facility, with most drillholes missing between six and 26 of the first core boxes, which contained material from the capping limestone unit that overlays the Carbonate (Table 8-1);

● No drill core of the Pennsylvanian strata exists, hence no information on the strata was gathered;

● Drill core is typically NQ, and some noted as being BQ;

● All drill core had been hydraulically split, removing the option of additional sampling for geotechnical purposes;

● Accurate geotechnical and hydrogeological parameters were difficult to estimate due to the core appearing to have been hydraulically split; and

● Identifying mineralization was difficult due to the fine-grained nature of the rock and a lack of differences between mineralized and non-mineralized rock.

In addition to the drill core, there also exists an unknown inventory of sample pulps and rejects at the Mead facility

8.1.2 NioCorp Drilling Program, 2011 - Current

A detailed core processing and sampling program was implemented during the 2011 Drilling program and continuously improved upon through the 2014 drill and historical core relogging program. The 2014 re-logging of the historical Molycorp drill core, stored at the Mead core library, and the 2011 Quantum (NioCorp) drill core, stored in the NioCorp processing facility, provided a consistent geological database spanning all project generations.

In 2011-2014, the core was boxed at the drill site and delivered each day to the project core processing facility where it was logged and split. The diamond-drilling programs utilized up to three coring drill rigs which were monitored by two qualified professional geologists and a trained

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geological crew. Professional project oversight was provided by geologists and engineers from Dahrouge, SRK, and NioCorp.

Standardized core logging codes and lithology descriptions were created in Datamine’s Fusion drill hole database to ensure consistency among logging geologists. A total of twenty-two detailed rock codes were used during the logging, which was subsequently reduced to ten codes under a simplified logging code defined as "Major Unit" in the database (Table 8-2).

Table 8-2: Summary of Major Rock Unit Codes

Major Unit	Rock Unit	Rock Code
INT	Syenite	INT
MAFIC	Mafic	maf
	Mafic Breccia	mafBc
CARB-LAMP	lamprophyre Dolomite Carbonatite Breccia	dolCarbLamp
MCARB-LAMP	lamprophyre Magnetite-dolomite Carbonatite Breccia	mdolCarbLamp
LAMP	lamprophyre Breccia	LamphBc
	lamprophyre	Lamph
MCARB	Hematite-dolomite Carbonatite	hemdolCarb
	Magnetite-dolomite Carbonatite	mdolCarb
	Magnetite-dolomite Carbonatite Breccia	mdolCarbBc
CARB	Dolomite Carbonatite	dolCarb
	Dolomite Carbonatite Breccia	dolCarbBc

Source: Dahrouge, 2015

The drill core within each core box was marked up and split along orientation marks. Cutting was completed using one of three electric-powered, water-cooled diamond-bladed BD 3003E core saws at the Project sample preparation and storage facility. HQ and minor intervals of PQ core were halved for assay. Drill hole NEC14-MET-03, a PQ-sized hole, was quartered with one quarter being assayed, and the remaining core packaged for metallurgical testing.

Infrequent broken or soft sections of the core (typically the iron oxide altered zones) were sampled by the geologists, and an equal sample split was taken from this material. These intervals account for a significantly small portion of the sampled material. Core not used for assaying or metallurgical testing is stored at the project facility work area at the Project site.

Drill core was digitally photographed under natural outdoor or indoor fluorescent lighting before core cutting. All digital photos are of high resolution and stored in a digital archive format. The geological logs included observations of colour, lithology, texture, structure, mineralization, and alteration. All geological information is collected at a sample-interval scale and recorded into the Fusion Database, which was the digital core logging and sampling storage software program.

SRK was responsible for the geotechnical logging. Rock quality was determined using the Q-system (Q=(RQD/Jn)*(Jr/Ja)*(Jw/SRF), where RQD = Rock quality designation; Jn = Joint set number; Jr = Roughness of the most unfavorable joint or discontinuity; Ja = Degree of alteration or filling along the weakest joint; Jw = Water inflow; SRF = Stress reduction factor. SRK personnel also recorded hardness and weathering to aid in geotechnical parameters for the future mine design. Core recovery and RQD were generally competent for the majority of the drill core. Core recovery was recorded in the database and was measured in the field at the drilling rig by the geologist.

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The borehole name is noted, and the drilling interval was compared to the actual core measured to back-calculate the recovery. The recovery information was loaded into the sample database.

The sampling procedure used to collect core samples entailed:

● The sampling of the entire carbonatite intersection, including the geologically-logged low- grade niobium carbonatite intervals of the footwall or hanging wall, for all holes except NEC14-020 to NEC14-023, where approximately 10 m of the hanging wall was sampled;

● Sample intervals, generally 1 m in length, were marked on the core and recorded in the geological database (Datamine Fusion).

● Sample intervals were assigned a unique sample number.

● Specific gravity measurements were performed at approximately 6 m spacing.

● Hand-held Niton-XRF measurements were collected on the core to assist geological and sample divisions.

● Magnetic susceptibility measurements were performed on the core to assist geological and sample divisions.

● Clearly marked sample intervals were split in half using a wet diamond saw.

● Split intervals were cleaned before bagging, and the cutting equipment was regularly cleaned.

● Sampled intervals were placed in durable barcoded sample bags that were clearly labelled and contained back up sample tags within each bag.

● Sample bags containing original core sections and field inserted control samples were barcode-scanned and secured in five-gallon plastic shipping pails.

● Hard copy and digital detailed shipping logs and preparation requests were sent to the primary analytical laboratory.

● Sampled core sections and blind control samples were shipped for analysis in secured pails and transferred using a bonded trucking company.

● Storage of the unsampled half of the core in labelled wooden core boxes at the Project site for reference or further sampling.

● The core samples and the core library are securely stored in the locked core processing facility directly on the Property (Figure 8-1).

NioCorp employed rigorous security measures to prevent tampering of the core or samples before and during the transport process. These measures included redundant sample identification, appropriate sample bag closures and the shipment of sample bags inside pails with lids. The authors are of the opinion that these measures are consistent with current industry best practices for projects at this scale of exploration.

The sample collection, preparation, and shipment workflow process were standardized and monitored to reduce or eliminate the downstream progression of incorrectly identified samples. The on-site professional geologist managed the drilling QA/QC program, which consisted of the insertion of control samples to monitor each stage of preparation and analysis. These control samples included laboratory-blind certified reference material samples (CRMs), optical-quality quartz blanks, field duplicates, coarse-reject duplicates, pulp duplicates, and external “umpire” lab duplicates. All samples were prepared and analyzed at Activation Laboratories (Actlabs), and select samples were subsequently sent to SGS Labs for a secondary check analysis (Figure 8-2).

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Source: Dahrouge, 2021

Figure 8-1: Storage Location of Drill Core and Pulps

Source: Dahrouge, 2014

Figure 8-2: Sample Process Flow Chart (2014 Drill Program)

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Once the field and control samples were received at Actlabs, they were crushed, pulverized, and split. A pulverization target of 95% passing 200 mesh (-200 mesh) was applied, with a quartz wash between the preparation of each sample. All duplicate sample splits were extracted simultaneously to their parent samples to ensure a sample level of homogenization and handling procedures. The 2011 program did not utilize laboratory blind duplicates.

● Pulps were prepared by Actlabs, which were then subsequently assayed using ICP-MS whole rock analysis, complete elemental packages, and XRF niobium analysis.

● Dahrouge reviewed the initial results of the ICP-MS whole-rock analysis and selected 55 samples that were analyzed for FUS-ICP and ICP-MS reporting niobium and whole rock values. External pulps splits for check analysis were created at the same time as primary pulps.

8.1.3   Historical Re-Sampling Programs

Between 2010 and 2021, NioCorp has undertaken multiple historical re-sampling and QA/QC programs to increase interval, analyte, and QA/QC sample coverage.

8.1.3.1   NioCorp (Quantum 2010) Historical Re-Sampling Program

The 2010 re-sampling program involved sending 1,860 samples of pulverized material to ALS testing facility in North Vancouver, B.C., from the Molycorp drill holes that were originally prepared by the analytical division of Molycorp. Original samples were derived from 1.52 m (5 ft) or 3.05 m (10 ft) intervals of split NQ or HQ diameter core. The samples were selected based on the geological interpretation at the time and in areas of elevated Nb2O5 values. The purpose of this program was to have continuous sample representation down each drill hole. Follow-up sampling programs, defined in the preceding sections, expanded these targeted intervals to cover the full Carbonatite intervals.

This re-sampling exercise also allowed the opportunity to increase the amount of QA/QC data available over the historical data period. A protocol was implemented, which included the routine insertion of field duplicates, laboratory pulp duplicates, blanks and two niobium-certified reference standards (SX18-01 and SX18-05) (Table 8-3). Samples were transported to the ALS Chemex (ALS) facility in Reno, Nevada, where they were prepared for analysis prior to being shipped to the ALS testing facility in North Vancouver, B.C. The ALS testing facility, using method XE-XRF10, whereby samples are prepared by pulverizing to 90% passing -70 pm, then decomposed utilizing a lithium borate flux, for analysis by XRF. A portion of niobium results was checked with Hazen of Golden, Colorado (Quantum news release February 22, 2011).

Table 8-3 Summary of 2010, ALS Labs, Re-Sampling Program Submissions

Sample Category	Sample Type	Material Source	Total	Insertion Rate
Original Re-Sampled	Pulverized Core	Molycorp Drill Core	1860	NA
Duplicates	Field Pulp Duplicate	Molycorp Drill Core	21	1.3%
Control Blank	Quartz Blank	Optical Quartz	54	2.9%
Standard Reference Materials	SX18-01 (Dillinger Hütte Lab)	Carbonatite (Nb2O5, La, Ce, Nd)	46	2.5%
	*SX18-05 (Dillinger Hütte Lab)	Carbonatite (Nb2O5 , La, Ce, Nd)	47	2.5%

Source: Dahrouge, 2022

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Nordmin evaluated the controls samples, SRM SX18-01 and SX18-05 and concluded results ran consistently low using the ME-XRF10 methodology compared to the 2011 and 2014 drill programs (Nordmin, 2019). The 2011 and 2014 sampling or re-assaying programs did not use this methodology.

8.1.3.2  NioCorp (2014-2016) Historical Re-Sampling Program

During the 2014 drilling program and the 2015 Mineral Resource Estimate (SRK, 2015), it was noted that Sc results were not included in the 2010 ALS resampling program and a portion of the database was missing assays grades for TiO₂, and Sc used in the estimate. Two phases of resampling were completed to increase confidence in these intervals and incorporate a more rigorous QA/QC program into the analysis.

The first phase completed in 2014-2015 included resampling of the historical Molycorp pulverized drill core material from the 2010 resampling program, completed at ALS. A total of 1410 samples, including 67 standards (GRE-04) and 8 pulp duplicates, were submitted to SGS Labs, and provided in-fill results for the missing Sc values and the incorporation of a Sc specific standard (Table 8-4). Analysis at SGS was restricted to Sc for this program, since a multi-element analysis was completed during the 2010 ALS sample submission.

Table 8-4: Summary of 2014-2015, Targeted Sc Re-Sampling Program Submission to SGS Labs.

Sample Category	Sample Type	Material Source	Total	Insertion Rate
Original Re-Sampled (Sc only)	Pulverized / fine-Crush splits	Molycorp Drill Core	1410	NA
Duplicates	Pulp Duplicates (Sc)	Molycorp Drill Core	8	0.6%
Certified Reference Materials	GRE-4 (Geostats PTY Ltd)	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	67	4.8%

Source: Dahrouge 2022

A second phase of resampling completed in 2016, focused on infilling missing results, while collecting whole rock and multi-element analysis. The sourced samples were a combination of fine-grained chips (crush) and pulverized samples from the original 1/2 NQ core samples, sourced from the storage facility in Mead. The samples were selected based on absent analyte values from the 2015 Mineral Resource Estimate and were located under the guidance of NioCorp’s geologist. During the re-assay program, NioCorp included QA/QC samples in the form of standards and duplicates. In total, 766 samples were included in the program of which 44 were pulp duplicates, and 55 were standard reference material inserted approximately every 15 samples (Table 8-5). The standards used were GRE-03, GRE-04 and SX18-01. The samples reanalyzed were submitted to Actlabs using the same Code 8-Nb₂O₅ & Ta₂O₅ - XRF option, and ICP/MS methods described in Section 8-2 of this TRS.

Table 8-5: Summary of 2016, Re-Sampling Program Submissions to Act Labs.

Sample Category	Sample Type	Material Source	Total	Insertion Rate
Original Re-Sampled	Pulverized / fine-Crush splits	Molycorp Drilling	667	NA

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Duplicates	Pulp split	Molycorp Drilling	44	6.6%
Certified Reference Materials	GRE-3 (Geostats PTY Ltd)	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	21	3.1%
	GRE-4 (Geostats PTY Ltd)	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	21	3.1%
Standard Reference Materials	SX18-01	Carbonatite SRM (Nb2O5, La, Ce, Nd)	13	1.9%

Source: Dahrouge, 2022

8.1.3.3 NioCorp (2021) Historical Re-Sampling Programs

An internal evaluation on REE potential within the existing Resource Estimates identified analyte gaps, for REE’s and Sc, within the lower-grade Nb₂O₅ outer boundaries of the resource. A total of 1095 samples, containing historical Nb₂O₅ results, lacked REE and Sc results and were selected for resampling and analysis at Act Labs using the methods defined in Section 8.2. A list of the drillholes, sample storage location and number of assay results that were missing are presented in Table 8-6 and represented as blue drillhole intervals in Figure 8-3.

Table 8-6: Pre-2021 Missing REE and Sc Assays that have Nb₂O₅ Database Results.

Resource Area Drillholes	Source / Storage Facility	Missing REE and Sc Assays
EC-011	Molycorp Samples / Mead Core Warehouse	65
EC-014	Molycorp Samples / Mead Core Warehouse	16
EC-015	Molycorp Samples / Mead Core Warehouse	151
EC-016	Molycorp Samples / Mead Core Warehouse	26
EC-018	Molycorp Samples / Mead Core Warehouse	92
EC-019	Molycorp Samples / Mead Core Warehouse	53
EC-020	Molycorp Samples / Mead Core Warehouse	30
EC-021	Molycorp Samples / Mead Core Warehouse	45
EC-022	Molycorp Samples / Mead Core Warehouse	57
EC-024	Molycorp Samples / Mead Core Warehouse	19
EC-026	Molycorp Samples / Mead Core Warehouse	86
EC-027	Molycorp Samples / Mead Core Warehouse	34
EC-029	Molycorp Samples / Mead Core Warehouse	27
EC-030	Molycorp Samples / Mead Core Warehouse	25
EC-031	Molycorp Samples / Mead Core Warehouse	47
EC-032	Molycorp Samples / Mead Core Warehouse	111
EC-034	Molycorp Samples / Mead Core Warehouse	54
EC-037	Molycorp Samples / Mead Core Warehouse	74
EC-054	Molycorp Samples / Mead Core Warehouse	83
	Total	1095

Source: Dahrouge, 2021

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Source: Dahrouge, 2021

Figure 8-3: Resource Area Assay Distribution Showing REE Assays (Red) and REE Assay Gaps (Blue).

The sourced samples were a combination of fine-grained chips (crush) and pulverized samples from the original 1/2 NQ core samples, sourced from the storage facility in Mead. A total of 1180 samples were analyzed, including 1094 interval samples (24 coarse-splits, 23 chip-splits, and 1047 pulverized splits) and 86 standards (AMIS0815, GRE-03, Oreas 460, and Oreas 464), with the details summarized in Table 8-7.

Table 8-7: Pre-2021 Missing REE and Sc Assays that have Nb₂O₅ Database Results.

Sample Category	Sample Type	Material Source	Total	Insertion Rate
Original Re-Sampled	Pulverized / fine-Crush/ Coarse-Crush splits	Molycorp Drilling	1094	NA
Certified Reference Materials	GRE-3 (Geostats PTY Ltd)	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	17	1.6%
	Oreas 460	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	29	2.7%
	Oreas 464	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	25	2.3%
	AMIS0185	Carbonatite CRM (Nb2O5, LREE)	15	1.4%

Source: Dahrouge, 2022

8.2 Sample Analysis Procedures, 2011 – Current

The 2011 and 2014 sawn core samples were shipped to Activation Laboratories Ltd. (Actlabs), 41 Bittern Street (previously 1336 Sandhill Drive), Ancaster, Ontario, Canada. Actlabs was the primary laboratory for sample preparation and analysis of the 2011 and 2014 drill core samples and for the 2016 and 2021 resample programs. Actlabs regularly participates in proficiency testing and maintained formal approval for CAN-P-1578, CAN-P-1579, CAN-P-1585, CAN-P-4E during the 2011-2016 programs and currently maintains ISO/IEC 17025:2017, and ISO 9001:2015 accreditation

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from Standards Council of Canada and Canadian General Standards Board, respectively. Actlabs maintains ISO-17025 standards, which are obtained through experienced peer audits that ensure they conform to recognized analytical standards. Additionally, the accredited method validation verifies several analytical variables designed to ensures that data obtained from these methods are defensible. Actlabs maintain a custom Laboratory Information Management System (LIMS) system for reporting requirements.

NioCorp employed SGS through the 2014 sampling program as their secondary laboratory. SGS is an integrated geochemistry, mineralogy, and metallurgy laboratory in Lakefield, Ontario which has extensive experience with Nb₂O₅ and REE analysis for both exploration and metallurgy projects. SGS Lakefield is IS017025 accredited for the analysis methods used on this project (GO_XRF76V & GE JCP90A).

Core samples were shipped to Actlabs, where they were received, weighed, prepared, and assayed. Sample preparation is completed using Actlabs' RX1 preparation package that has been modified to meet the Project requirements. A summary of the process is detailed below:

● Samples were received and cataloged.

● Collection of as-received sample weight (kg).

● Drying of the whole sample at 60°C for 12 hours, in a customized high air flow drying room.

● Collection of dry sample weight (kg).

● Crushed in a jaw crusher (Boyd crushers) to 90% passing -10 mesh (2 mm), with quartz cleaner between each sample.

● Riffle split (RSD splitters or the option of Jones Riffle split) coarse crushed sample and extract a 250 g sample.

● Pulverization of the 250 g sample using ESSA pulverizers with ring and puck bowls to 95% -200 mesh (75 μm), with quartz cleaner used between each sample.

● Laboratory internal coarse-reject duplicates (1 in 50) and pulp duplicates (1 in 30) are also routinely prepared.

● Quality of the rejects and pulps are routinely monitored to ensure proper preparation procedures are performed.

Core samples were systematically assayed at Actlabs for niobium (Nb₂O₅) and tantalum (Ta₂O₅) by XRF analysis, using a Panalytical Axios-mAX, following a lithium metaborate/tetraborate fusion of a 2 g sample. All XRF analysis followed procedures outlined in Actlabs "8-XRF" package, with selected analytical results provided for Nb₂O₅ and Ta₂O₅. A whole rock and forty-three (43) major elements analyses were completed using ICP and ICP/MS (by a Perkin Elmer Sciex ELAN 6000, 6100, 9000 ICP/MS) finish following a lithium metaborate/tetraborate fusion preparation as defined by analytical Actlabs' "8-REE Major Elements Fusion ICP(WRA)/Trace Elements Fusion ICP/MS(WRA4B2)" package.

Additional analysis was performed for fluoride, using the analytical package "4F-F". Fluoride content is quantified using a fluoride ion electrode to directly measure fluoride-ion activity when a prepared fuseate is dissolved in dilute nitric acid and its ionic strength adjusted in ammonium citrate buffer. Before the analysis, the sample is prepped using a combined fusion with lithium metaborate and lithium tetraborate in an induction furnace. Fluoride analysis was completed for 2014 drill holes, NEC14-006, NEC14-007, and NEC14-008.

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All QC data is registered in the LIMS system, and assay results have been returned to NioCorp in an electronic format. Following the QA/QC review by the qualified geologist, the results are loaded into the Fusion database with the batch number and date of assay recorded.

During the preparation procedure, coarse-reject splits and pulp-splits are extracted from the original core sections for primary laboratory and secondary (external) laboratory check analysis. These samples are then inserted into the sampled sequence.

Pulp samples are routinely extracted with inserted CRM samples which were prepared by Actlabs and shipped to SGS (Lakefield), where they were received, evaluated for sample quality and re-homogenized, and assayed. SGS (Lakefield) prepared and re-homogenized samples before analysis using MISC80. During preparation, SGS completed a 10% sieve check (SCR32 package) to ensure 95% sample pulverization passes 200 mesh (75 μm) preparation requirements. Samples were assayed using an XRF analysis for Nb₂O₅ and thirteen major whole rock oxides, following a borate fusion as defined under SGS package "GO XRF76V - ORE GRADE" (see Table 8-8). Scandium analysis has been completed at SGS laboratory using GE JCP90A package, which has a detection limit of 5 ppm.

Table 8-8: Detection Limits for Primary Laboratory (Actlabs)

XRF (%)		Trace Elements ICP & ICP/MS (ppm)
Oxide	Detection Limit	Element	Detection Limit	Reported By	Element	Detection Limit	Reported By
Nb2O5	0.003	Ag	0.5	ICP/MS	Nb	1	ICP/MS
Ta2O5	0.003	As	5	ICP/MS	Nd	0.1	ICP/MS
4F-F (%)		Ba	3	ICP	Ni	20	ICP/MS
Analysis	Detection Limit	Be	1	ICP	Pb	5	ICP/MS
F	0.01	Bi	0.4	ICP/MS	Pr	0.05	ICP/MS
Fusion ICP (%)		Ce	0.1	ICP/MS	Rb	2	ICP/MS
Oxide	Detection Limit	Co	1	ICP/MS	Sb	0.5	ICP/MS
SiO2	0.01	Cr	20	ICP/MS	Sc	1	ICP
Al2O3	0.01	Cs	0.5	ICP/MS	Sm	0.1	ICP/MS
Fe2O3	0.01	Cu	10	ICP/MS	Sn	1	ICP/MS
MgO	0.01	Dy	0.1	ICP/MS	Sr	2	ICP
MnO	0.001	Er	0.1	ICP/MS	Ta	0.1	ICP/MS
CaO	0.01	Eu	0.05	ICP/MS	Tb	0.1	ICP/MS
TiO2	0.001	Ga	1	ICP/MS	Th	0.1	ICP/MS
Na2O	0.01	Gd	0.1	ICP/MS	T	0.1	ICP/MS
K2O	0.01	Ge	1	ICP/MS	Tm	0.05	ICP/MS
P2O5	0.01	Hf	0.2	ICP/MS	U	0.1	ICP/MS
Loss on Ignition	0.01	Ho	0.1	ICP/MS	V	5	ICP
		In	0.2	ICP/MS	W	1	ICP/MS
		La	0.1	ICP/MS	Y	2	ICP
		Lu	0.04	ICP/MS	Yb	0.1	ICP/MS
		Mo	2	ICP/MS	Zn	30	ICP/MS
					Zr	4	ICP

Source: Dahrouge, 2015

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8.3 Quality Assurance/Quality Control Programs

Quality Control (QC) measures are typically set in place to ensure the reliability and trustworthiness of exploration data. These measures include written field procedures and independent verifications of aspects such as drilling, surveying, sampling, and assaying, data management and database integrity. Appropriate documentation of quality control measures and regular analysis of quality control data are essential as a safeguard for project data and form the basis for the Quality Assurance (QA) program implemented during exploration.

Analytical QC measures typically involve internal and external laboratory procedures implemented to monitor the precision and accuracy of the sample preparation and assay data. They are also important to identify potential sample sequencing errors and to monitor for contamination of samples.

Sampling and analytical QA/QC protocols typically involve taking duplicate samples and inserting quality control samples (CRMs and blanks) to monitor the reliability of the assay results throughout the drill program. Umpire check assays are typically performed to evaluate the primary lab for bias and involve re-assaying a set proportion of sample rejects and pulps at a secondary umpire laboratory.

8.3.1 Re-Sampling/Verification of Historical Assays

Between 2010 and 2021, NioCorp completed extensive re-sampling/verification work programs concerning the historical assays.

● The 2010, a total of 1,860 interval samples were re-sampled during the program and subjected to the current QA/QC protocols. The selection for re-assay was based on available material and proximity to the mineralization wireframe used during that study.

● The 2015/2016 re-assay program was completed which consisted of sending 1,410 pulps to SGS, which previously were not analyzed for titanium and scandium. This included the insertion of additional QA/QC material. The QA/QC program by NioCorp used a CRM sourced from Geostats (GRE-04), which contained a certified value for Sc (ppm) and Ti (%).

● The 2016 re-assay program included a total of 667 pulps were sent to ActLabs for whole rock and multi-element analysis. These samples pulp duplicates, standard reference material Geostats (GRE-03, GRE-04) and Dillinger Hütte (SX18-01).

● A total of 1094 interval samples, in 2021, were selected and re-assayed, with the incorporation of current QA/QC standard protocols. The QA/QC program by NioCorp used CRMs sourced from Geostats (GRE-03), Oreas (Oreas 460 and 464), and African Mineral Standards (AMIS0185), which contained specific certified value for (Nb₂O₅, Sc, TiO₂, REE).

A summary of these programs and the associated control samples, source, and level of insertion for the re-assay program are included in Table 8-9 and Table 8-10. These Samples were selected from a historical database containing 9008 compiled Molycorp assay results covering the resource area and select other areas of the complex.

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Table 8-9: Summary of Actual Submissions per Sample Type within the 2015-2021 Re-Assay Program

Re-Sampling Programs	Purpose	Company	Lab	Material	Interval Samples	Control Samples
2010	Historical Assay Gaps	Quantum (NioCorp)	ALS	Molycorp Pulverized and Course-Splits	1860	168
2014-2015	Sc (2010 re-sample infill)	NioCorp	SGS	Molycorp Pulverized and Course-Splits	1410	75
2016	Historical Assay Gaps	NioCorp	Actlabs	Molycorp Pulverized and Course-Splits	667	99
2021	Historical Assay Gaps	NioCorp	Actlabs	Molycorp Pulverized and Course-Splits	1094	86

Source: Dahrouge, 2022

Table 8-10: Summary of Actual Submissions per Sample Type within the 2015-2021 Re-Sample Program

Sample Category	Sample Type	Material Source	Total	Insertion Rate
Original Re-Sampled	Pulverized / fine-Crush splits	Molycorp Drilling	5031	NA
Control Blank	Quartz Blank	Optical Quartz	54	1.1%
Duplicates	Field Pulp Duplicate	Molycorp Drill Core	21	0.4%
	Pulp split	Molycorp Drilling	52	1.0%
Certified Reference Materials	GRE-3 (Geostats PTY Ltd)	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	38	0.8%
	GRE-4 (Geostats PTY Ltd)	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	88	1.7%
	Oreas 460	Carbonatite CRM (Nb2O5, Sc, TiO2, REE)	29	0.6%
	Oreas 464	Carbonatite CRM (Nb2O5 Sc, TiO2, REE)	25	0.5%
	AMIS0185	Carbonatite CRM (Nb2O5, LREE)	15	0.3%
Standard Reference Materials	SX18-01 (Dillinger Hütte Lab)	Carbonatite SRM Nb2O5 (, La, Ce, Nd)	59	1.2%
	*SX18-05 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, La, Ce, Nd)	47	0.9%

Source: Dahrouge, 2022

8.3.2 NioCorp 2011 - Current

NioCorp integrated a series of routine QA/QC procedures throughout the sampling and analytical analysis for both the 2011 and 2014 drilling programs to ensure the highest level of quality was maintained throughout the process. This included the insertion of duplicate samples taken from various stages of the process, insertion of known control samples (SRMs, CRMs and blanks) and sending third-party pulps to the secondary lab (SGS).

Sample tickets were assigned initially at the core shed using barcodes with duplicate tickets placed inside and on the outside of the bag. Sample identification was confirmed using barcode labelling and visual sample type comparisons before sample shipment. The use of barcoded samples ensured both shipment forms and analytical labs used accurate information. Multiple types of QC samples were inserted at this stage of the process, which includes the following:

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● Field quartz blanks (1 in 20, or 5%) were inserted within or immediately after samples collected from mineralized intervals, targeting zones of elevated visual mineralization, where possible.

● CRMs (1 in 20, or 5%) were inserted in the field with the sample sequence.

● Field quarter-core duplicates (1 in 20, or 5%) were inserted to test mineralization and sampling variability.

These following control measures were used to monitor both the precision and accuracy of sampling, sub-sampling, preparation, and assaying. A summary of the designed type of samples, source and level of insertion is included in Table 8-11 and the actual submissions in Table 8-12 and Table 8-13.

Table 8-11: Summary of Designed Level of Insertion of QC Submissions (2011 and 2014 Drill Program)

Sample Type	Sample Sub-type	Type	Insertion Rate
Blanks	Field Quartz Blanks	Optical Quartz	5.0%
Certified Reference Material	SX18-01 (Dillinger Hütte Lab)	Nb CRM	6.0%
	SX18-02 (Dillinger Hütte Lab)	Nb CRM
	SX18-04 (Dillinger Hütte Lab)	Nb CRM
	SX18-05 (Dillinger Hütte Lab)	Nb CRM
Duplicates	Field quartered core	¼ HQ Core	5.0%
External Lab Checks	Coarse-Rejects	Reject split	3.0%
	Pulp	Pulp split	5.0%
	Field Quartz Blanks	Optical Quartz	(5% of splits)

Source: Dahrouge, 2014

Table 8-12: Summary of Sample and Control Submissions for Nb₂O₅, Sc TiO₂, REE’s (2011 Drill Program)

Actlabs Primary Analysis & Control Samples
Sample Category	Sample Type	Material Type	Total Samples	Insertion Rate (%)
Original Samples	Original Section	NioCorp 1/2 or 1/4 HQ core	1776	NA
Duplicate Samples	Quartered-Core Duplicate	NioCorp 1/4 HQ core	90	5.07%
	Coarse-Reject Duplicate	Reject split	86	4.84%
	Pulp Duplicate	Pulp split	86	4.84%
Standard Reference Material	SX18-01 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	23	1.30%
	SX18-04 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	15	0.84%
	*SX18-05 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	19	1.07%
Certified Reference Material	AMIS0185	Carbonatite CRM Nb2O5 (, LREE)	31	1.75%
Blanks	Field Quartz Blanks	Optical Quartz	90	5.07%
Inspectorate External Check Samples: Pulp-Split
External Duplicate	Pulp Duplicate	Pulp Split	82	5.63%
Blind Duplicate	External Pulp Duplicates	Pulp Split	11	11%

Source: Dahrouge, 2022

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Table 8-13: Summary of Sample and Control Submissions for Nb₂O₅, Sc TiO₂, REE’s (2014 Drill Program)

Actlabs Primary Analysis & Control Samples
Sample Category	Sample Type	Material Source	Total	Insertion Rate
Original Samples	Original Section	NioCorp 1/2 or 1/4 core (HQ & PQ)	9653	NA
Duplicate Samples	Quartered-Core Duplicate	NioCorp 1/4 core (HQ & PQ)	419**	4.34%
	Coarse-Reject Duplicate	Reject split	260	2.69%
	Pulp Duplicate	Pulp split	468	4.85%
Standard Reference Material	SX18-01 (Dillinger Hütte Lab)	Carbonatite SRM Nb2O5 (, TiO2, La, Ce, Nd)	169	1.75%
	SX18-02 (Dillinger Hütte Lab)	Carbonatite SRM Nb2O5 (, TiO2, La, Ce, Nd)	154	1.60%
	SX18-04 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	165	1.71%
	*SX18-05 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	8	0.08%
Blanks	Field Quartz Blanks	Optical Quartz	454	4.70%
SGS External Check Samples: Pulp-Split
External Duplicate	Pulp Split	Original Pulp	462	4.79%
Blind SGS Duplicate	Pulp Split Duplicate	Original Pulp	44	9.52%
Standard Reference Material	SX18-01 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	17	3.68%*
	SX18-02 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	17	3.68%*
	SX18-04 (Dillinger Hütte Lab)	Carbonatite SRM Nb2O5 (, TiO2, La, Ce, Nd)	14	3.03%*
	*SX18-05 (Dillinger Hütte Lab)	Carbonatite SRM (Nb2O5, TiO2, La, Ce, Nd)	1	0.22%*

Source: Dahrouge, 2022

**Total includes 2 miss-matched quarter core duplicates

*Insertion rate is a percentage of total External Check Samples submitted

*Does not include any duplicates for the CRM in 2011

Except for 2016, the QA/QC data was analyzed by the project geologist on a routine basis before entering the data into the central database. When SRMs or CRMs failed, each result was checked

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for possible sample swaps or significant failures. A single failure did not guarantee the failure of the entire batch, but multiple failures warrant re-analysis. Failures were reported directly back to the laboratory, for re-analysis. In the 2011 drill program, no standards were identified as failing with a high enough level of concern to request re-assay. Any revision to the 2011 certificates was either a result of further re-sampling or internal lab revisions. In the 2014 program, when it was determined that a SRM or CRM failure was significant, ten samples on either side of the SRM or CRM were re-assayed. The re-assay was taken as the final assay result, and the original certificate was overwritten when imported into Fusion. Fusion maintains both certificates and can produce an audit trail to detail which assays have been updated. A limited number of original samples were re-sampled throughout the program.

The following section provides details of the types of samples used at each section of the sampling process and followed by a discussion of the QA/QC results.

8.3.3 Quality Assurance & Quality Control Results

8.3.3.1 Field Quartz Blanks

Coarse natural clear quartz blanks (sourced from an optical-quality quartz quarry, in Arkansas, USA) were inserted into the sample sequence to identify potential contamination and to confirm sample sequence consistency.

Methodology:

● Utilization of the same sample preparation system as project samples.

● Field quartz blanks were selected as a hard element-homogenous material that would pick up material contamination from precedding samples.

● Matieral is placed in sequential order and not advertized to the lab.

● Random spiked blanks are inserted into the sequence to ensure all contaminated samples are being recorded propoerly.

The field quartz blanks used a base detection limit of 2 x XRF detection limits and 20x (30x for La and Pr) ICP-MS detection limits. Results falling above this value are reported to the laboratory as having potential contamination and additional cleaning cycles are introduced into the procedures if the issue is consistent.

It is noted that both sets of controlled blanks used in the two analytical procedures exhibited clusters of contamination at the start of the program. These clusters were significantly reduced as the program advanced; the failing blanks were re-analyzed to differentiate between inaccurate analytical results and potential contamination points. The findings were discussed with the lab to allow for corrective measures to be implemented (Table 8-14 and Figure 8-4).

Table 8-14: Summary of 2011 and 2014 Drill Program Nb₂O₅ Blank Insertion

Element Nb2O5 – Actlabs	Drill Program
Year (Lab Package: REE Package + 8-Nb2O5 XRF)	2011	2014
# of Assays sent to Lab	1776	9,653
# of Field Quartz Blanks Sent to Lab	90	454
Insertion Rate of Blanks	5.1%	4.7%
# of Blank Failure (2x XRF Detection Limit)	35	19
Percentage of Blank Failure Rate	39%	4%

Source: Dahrouge, 2022

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Source: Dahrouge, 2022

Figure 8-4: Summary of Blank Control Charts for Nb₂O₅ , Sc, TiO₂ Submission to Actlabs 2011 and 2014 Drill Program

It is noted the TiO₂ QA/QC data is more variable than the Nb₂O₅ data for the 2011 and 2014 program. Overall, the majority of the 2014 samples are less than 0.02% control line, which is the equivalent of 20x the detection limit, above which potential contamination may be identified.

Overall Dahrouge considers that the blank material has acceptable levels of error and there is limited evidence of any major contamination issues at the laboratory since the problem was identified and corrected during the 2014 drill program (Figure 8-5).

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Source: Dahrouge, 2022

Figure 8-5: Summary of Blank Control Charts for REEs, La, Ce, Nd, and Pr, Dy Submission to Actlabs

8.3.3.2   Certified & Standard Reference Material

Different suppliers of Certified Reference Material (CRM) and Standard Reference Material (SRM) were used on the Project during the 2010 re-assay program through the 2021 drill and sampling programs. The selected standards were representative of a carbonatite matrix, but not all standards covered the targeted mineralization of niobium, titanium, and scandium. A total of 4 SRMs and 6 CRMs were used to evaluate the range of material.

The primary standard used through the duration of the programs was the SX-Series SRMs, from Dillinger Hütte (SX18-01, SX-02, SX-04, and SX-05), which provide reference control of the Nb₂O₅ , TiO₂, La, and Ce. Additional CRMs incorporated into the check and reasampling programs included, standards from Geostats (GRE-03 and GRE-04), American Mineral Standards (AMIS0815), Ore Reserch & Exploration (Oreas 460 and Oreas 464), which were added to better represent Sc, TiO2, REE’s and validate Nb₂O₅ ranges (Table 8-15).

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Table 8-15: Summary of CRM & SRMs Controls Used on the Project 

		CRM - Best Values						SRM - Best Values
Analyte	Unit	AMIS0815	GRE-02	GRE-03	GRE-04	OREAS 460	OREAS 464	SX18-01	SX18-02	SX18-04	SX18-05
Nb2O5	%	-	0.055	0.504	0.508	0.100	0.272	0.695	0.199	1.320	0.973
Sc	ppm	-	76.70	49.85	88.36	27.90	141.00	-	-	-	-
TiO2	%	-	0.438	1.633	2.769	2.000	3.260	0.266	0.237	0.287	0.295
La	ppm	29760.0	9786	2224	2736	1369	12000	358.0	350.0	759.0	501.0
Ce	ppm	40740.0	16797	4354	6127	1798	15300	773.0	798.0	1425.0	1042.0
Pr	ppm	3471.0	1883	497	721	244	2597	-	-	-	-
Nd	ppm	9238.0	7048	1836	2702	781	9940	-	-	-	-
Sm	ppm	556.0	769.2	279.0	390.0	107.0	1498.0	-	-	-	-
Eu	ppm	-	139.90	75.00	101.00	22.70	324.00	-	-	-	-
Gd	ppm	-	262.20	191.00	233.00	50.00	676.00	-	-	-	-
Tb	ppm	-	14.62	22.00	24.00	4.84	54.00	-	-	-	-
Dy	ppm	-	28.56	92.00	97.00	19.80	178.00	-	-	-	-
Ho	ppm	-	2.87	14.00	14.00	2.77	21.30	-	-	-	-
Er	ppm	-	7.96	29.00	29.00	6.01	38.20	-	-	-	-
Tm	ppm	-	0.53	3.00	3.00	0.70	3.56	-	-	-	-
Yb	ppm	-	2.96	16.00	15.00	3.91	15.70	-	-	-	-
Lu	ppm	-	0.42	2.00	2.00	0.52	1.69	-	-	-	-
Y	ppm	-	56.0	320.6	319.4	60.0	449.0	134.0	126.0	168.0	232.0

Source: Dahrouge, 2022

8.3.3.2.1  Nb₂O₅ Standard and Certified Reference Material

A summary of the defined limits and results for Nb₂O₅ 2011-2014 drill programs and follow up re-sampling programs are shown in Table 8-16. The results for Nb₂O₅ CRM are summarized in Table 8-16 and charted in Figure 8-6, Figure 8-7 and Figure 8-8. The CRM submissions show an insertion rate of between 3.3% and 5.1%, for the 2011 and 2014 drill programs, respectfully, with greater the 5% insertion rate on re-sampling programs. The results show a relatively low failure less then 3.6 %, which are within acceptable limits.

All 6 of the inserted standard reference materials report a high bias, relative to the certified standard values, ranging from -1.2 to 12.7 % average relative difference, with the lowest average relative difference recorded in the higher-grade samples (SX18-04 and SX18-05) and Low-grade Oreas 460 CRM. The highest recorded in the GRE-03 and GRE-04 standards. Possible reasons for the continued high bias include:

● Certified or Standard Reference Material (CRMs or SRMs) are designed using round- robin analysis, with the best value being determined from results running both above and below the best value. Values that fall within the defined certified ranges are classified as acceptable and within the risk ranges assigned to that standard.

● The SX-Series certificates do not define the method or standard deviations, presenting only 95% confidence Intervals, resulting in potential analytical method mismatching and variations resulting from precision, equipment, or method used by Actlabs.

The continuation of a robust QA/QC program with the introduction of an additional check laboratory or analytical method will help to better quantify risk associated with this bias high. The current bias ranges and relative difference fall within all accepted parameters.

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Table 8-16: Summary of Nb₂O₅ CRM (Primary Assays - Actlabs) 

Standard (Nb2O5)	Count	Certified Value (%)	STD DEV (%)	Mean Assay (%)	Range (%)	Min (%)	Max (%)	N outside 10%
SX18-01	186	0.695	0.0695*	0.708	0.176	0.593	0.769	3	1.6%
SX18-02	138	0.199	0.0199*	0.207	0.025	0.193	0.218	0	0.0%
SX18-04	165	1.32	0.132*	1.4	0.216	1.256	1.472	6	3.6%
SX18-05	27	0.973	0.0973*	0.986	0.063	0.968	1.031	0	0.0%
GRE-03	33	0.504	0.062	0.572	0.069	0.549	0.618	0	0.0%
GRE-04	15	0.51	0.04	0.6	0.025	0.558	0.583	0	0.0%
Oreas 460	29	0.0998	0.0056	0.105	0.006	0.102	0.108	0	0.0%
Oreas 464	26	0.272	0.012	0.267	0.021	0.259	0.28	0	0.0%

Source: Dahrouge, 2022 

* 10% Certified value (no Standard Deviation)

 

Source: Dahrouge, 2022

Figure 8-6: Summary of SX18-01, SX18-02, SX18-04, SX18-05 Nb₂O₅ Control Chart

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Source: Dahrouge, 2022

Figure 8-7: Summary of GRE-04 Nb₂O₅ Control Chart

 

Source: Dahrouge, 2022

Figure 8-8: Summary of Oreas 460 and 464 Nb₂O₅ Control Chart

8.3.3.2.2  Sc Certified Reference Material

Certified Sc standards were not used during the 2011 and 2014 programs but were inserted into all resampling and QA/QC validation programs that followed. Results presented in this section include all CRMs that were included in results used in the primary assay database, including SGS (2014 and 2015) and historical re-sampling programs at Actlabs (2016 and 2021). A summary of the samples, defined limits, and results for Sc during the 2011 and 2014 drill programs are shown in Table 8-17, Figure 8-9 and Figure 8-10. The result fall within the accepted ranges and show no time-based variation trends. The average relative differences recorded were 2.58% (GRE-03), 0.29 % (GRE-04), 6.31% (Oreas 460), and 8.77% (Oreas 464) for the representative Sc CRMs.

Table 8-17: Summary of Sc CRMs (Primary Assays – Actlabs + SGS) 

Element (Sc)	Count	Certified Assay Value (ppm)	STD DEV (ppm)	Mean Assay (ppm)	Range (ppm)	Min (ppm)	Max (ppm)	N outside 3 STD DEV
GRE-03	33	49.9	1.6	51.18	54.43	1.57	56.00	1	3.0%
GRE-04	82	88.36	4.21	88.9	50	79	129	3	3.7%
Oreas 460	29	27.90	1.26	29.7	3	28	31	0	0.0%
Oreas 464	26	141.00	5.80	154.0	160	148	160	1	3.8%

Source: Dahrouge, 2022

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Source: Dahrouge, 2022

Figure 8-9: Summary of GRE-03 and GRE-04 Sc Control Chart

 

Source: Dahrouge, 2022

Figure 8-10: Summary of OREAS 460 and OREAS 464 Sc Control Chart

8.3.3.2.3  TiO₂ Standard and Certified Reference Material

A summary of the defined limits and results for TiO₂ during the 2011-2014 drill programs and follow up resampling programs is shown in Table 8-18. The results for TiO₂ CRM are demonstrated in Figure 8-11 to Figure 8-13. For TiO₂ standards, a significant issue identified was the grade range for a significant number of the TiO₂ in the CRM (SX18-01, SX18-02, SX18-04 and SX18-05), in the order of 0.25% to 0.30%, which is an order of magnitude lower than the typical grade ranges at the Project of 2.0% to 3.5% within the resource model. Given the low-grade nature of the assays in the TiO₂ CRMs, Dahrouge relied on additional support from the duplicate assays and external checks by SGS, as well as the results of the 2016-2021 re-assay program. The result fall within the accepted SRM and CRM ranges and show no time-based variation trends.

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Table 8-18: Summary of Sc CRMs (Primary Assays – Actlabs) 

Standard (TiO2)	Count	Certified Value (%)	STD DEV (%)	Mean Assay (%)	Range (%)	Min (%)	Max (%)	N outside 10%
SX18-01	202	0.027	0.266*	0.254	0.141	0.234	0.375	9	4.5%
SX18-02	154	0.237	0.0237*	0.231	0.086	0.201	0.287	5	3.2%
SX18-04	180	0.287	0.0287*	0.3	0.056	0.235	0.291	34	18.9%
SX18-05	27	0.295	0.0295*	0.3	0.027	0.267	0.294	0	0.0%
GRE-03	33	1.633	0.042	1.6	0.138	1.572	1.71	0	0.0%
GRE-04	15	2.769	0.080	2.6	0.22	2.509	2.729	1	6.7%
Oreas 460	29	2.00	0.05	2.0	0.149	1.913	2.062	0	0.0%
Oreas 464	26	3.26	0.10	3.2	0.232	3.004	3.236	1	3.8%

Source: Dahrouge, 2022 

* 10% Certified value (no Standard Deviation)

 

Source: Dahrouge, 2022

Figure 8-11: Summary of SX18-01, SX18-02, SX18-04, SX18-05 TiO₂ Control Chart 

 

Source: Dahrouge, 2022

Figure 8-12: Summary of GRE-03 and GRE-04 TiO₂ Control Chart

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Source: Dahrouge, 2022

Figure 8-13: Summary of OREAS 460 and OREAS 464 TiO₂ Control Chart

8.3.3.2.4  REE Certified Reference Material

The primary standards used through the duration of the programs were the SX-Series SRMs, which provide reference control of La, Ce, Nd, and Y. The SX-Series, SX18-01, -02 and -04 focused on the evaluation of La and Ce. These standards performed within acceptable ranges with a slight low bias across all grade ranges for both La and Ce. Reported results for both La and Ce were generally below the certified values, at -0.7% to -6.5% and -0.7% -6.4%, respectively. In general, these ranged between the assigned value and the +10% caution line.

A summary of the samples falling outside the failure level of +10% is presented in Table 8-19. There is a 1.9 to 7.8 % sample failure rate for La and a 4.9 to 9.7% sample failure rate for Ce. These standard reference materials are not optimal for REE evaluations but provided the needed confirmation that the results are within an accepted level of accuracy when combined with the AMIS, GRE-Series, and Oreas CRMs. No measures were taken to request batch re-analysis, since REE performance was not reviewed during the 2011 and 2014 drill programs or the 2016 re-sampling program.

The SX-Series SRMs performed poorly for Nd and Y, where more than 94% of the results fell outside the 10% evaluation ranges. Using information obtained from external carbonatite projects that used this standard, Dahrouge attributes this performance to inappropriate reference results for Nd and Y, rather than to laboratory performance. Additional support for the REE results was provided by the AMIS, GRE-series, and Oreas-series CRM results, presented below in Table 8-19 and Figure 8-14 to Figure 8-16.

Table 8-19: Summary of SX18-01, SX18-02, SX18-04, SX18-05 REE Results for SRM’s Inserted During 2011-2014

Element	Count	Certified Value (ppm)	10% Certified value (ppm)	Mean Assay (ppm)	Range (ppm)	Min (ppm)	Max (ppm)	N outside 10%
SX18-01
La	211	358.13	358.13	352.0	140	310	450	8	3.8%
Ce	211	773.00	773.00	783.7	263	700	963	2	0.9%
Nd	211	437.00	437.00	363.3	69	325	394	210	99.5%
Y	211	133.87	133.87	121.7	35	109	144	83	39.3%
SX18-02
La	154	349.60	349.60	347.5	78	303	381	3	1.9%
Ce	154	797.80	79.78	753.9	175	666	841	15	9.7%

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Nd	154	420.10	42.01	353.8	81	312	393	151	98.1%
Y	154	125.99	12.60	117.7	98	36	134	20	13.0%
SX18-04
La	180	758.89	75.89	715.0	773.9	2.1	776	6	3.3%
Ce	180	1424.65	142.46	1343.6	1505.7	4.3	1510	11	6.1%
Nd	180	619.01	61.90	522.1	570.5	1.5	572	178	98.9%
Y	180	167.73	16.77	166.4	45	151	196	1	0.6%
SX18-05
La	36	501.00	501.00	452.2	96	393	489	17	47.2%
Ce	36	1042.00	1042.00	970.9	202	838	1040	4	11.1%
Nd	36	511.00	511.00	430.5	95	376	471	33	91.7%
Y	36	232.30	232.30	235.4	61	201	262	2	5.6%

Source: Dahrouge, 2022

 

Source: Dahrouge, 2022

Figure 8-14: Summary of SX18-01, SX18-02, SX18-04, SX18-05 La Control Chart

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Source: Dahrouge, 2022

Figure 8-15: Summary of SX18-01, SX18-02, SX18-04, SX18-05 Ce Control Chart

 

Source: Dahrouge, 2022

Figure 8-16: Summary (2011 – 2016 results) of SX18-01, SX18-02, SX18-04, SX18-05 Nd Control Chart

During the 2011 drill program the AMIS0815 CRM was used to validate the REE results. During this time, the standards were monitored for LREEs, and no results fell outside the 3-standard deviation target range (see table 8-20). All LREEs, specifically the primary target elements, Nd and Pr, all fell within the 2-standard deviation target and showed less than a 6% average relative difference from the certified value (see Figure 8-17). The Nd and Pr showed a slight low bias with an approximate -

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5% average relative difference compared to the certified value. The performance observed for these standards and their coverage support accurate assay results and the incorporation of these standards in future work.

Table 8-20: AMIS0815 summary of REE Results of SRMs Inserted During 2011-2014

AMIS0815
Element	Count	Certified Assay Value (ppm)	STD DEV (ppm)	Mean Assay (ppm)	Range (ppm)	Min (ppm)	Max (ppm)	N outside 3 STD DEV
La	31	29760.00	1360.00	29358.1	3400	27400	30800	0	0.0%
Ce	31	40740.00	2305.00	41374.2	3200	39200	42400	0	0.0%
Pr	31	3471.00	171.50	3385.5	320	3210	3530	0	0.0%
Nd	31	9238.00	516.50	8971.9	1040	8510	9550	0	0.0%
Sm	31	556.00	24.00	8971.9	1040	8510	9550	0	0.0%
Eu	31	94.20	6.05	84.7	67.7	25.1	92.8	2	6.5%
Dy	31	27.10	2.55	22.9	4.7	20.9	25.6	0	0.0%

Source: Dahrouge, 2022

 

Source: Dahrouge, 2022

Figure 8-17: CRM AMIS0815 Control Charts for La, Ce, Pr, and Nd, and Dy (Provisional)

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A sample set was tested using the GRE-03, GRE-04, Oreas 460 and Orease 464 CRMs during a targeted re-assay program designed to increase sample coverage for scandium, while also assaying for REEs. These results, including 3 Standard deviation failure rates, are presented to show REE coverage ranges and performance (Table 8-21 to Table 8-24). These four CRMs were evaluated for all REEs and show strong reproducability for REEs, except Ce (GRE-03), and Er for both GRE-04 and Oreas 464. The primary target elements, focused on Nd, Pr, and Dy, fell with the 3-standard deviation target and showed less then a 2% average relative difference from the certified value (Figure 8-18 to Figure 8-22). The performance observed for these standards and their coverage support accurate assay results and the incorporation of these standards in future work.

Table 8-21: GRE-03 summary of REE Results Inserted During 2011-2021

Element (GRE-03)	Count	Certified Assay (ppm)	STD DEV (ppm)	Mean Assay (ppm)	Range (ppm)	Min (ppm)	Max (ppm)	N outside 3 STD DEV
La	33	2224.00	104.00	2285.5	360	2080	2440	0	0.0%
Ce	33	4354.10	83.50	4411.2	710	4050	4760	5	15.2%
Pr	33	496.60	25.80	492.5	61	461	522	0	0.0%
Nd	33	1835.90	96.60	1831.8	300	1730	2030	0	0.0%
Sm	33	279.40	14.20	283.9	49	263	312	0	0.0%
Eu	33	75.24	5.68	74.5	12.2	68.2	80.4	0	0.0%
Gd	33	191.00	10.50	173.5	32	159	191	1	3.0%
Tb	33	21.65	1.23	19.8	3.5	18.4	21.9	0	0.0%
Dy	33	92.33	6.36	86.3	13	81.5	94.5	0	0.0%
Ho	33	13.53	0.95	12.5	2	11.6	13.6	0	0.0%
Er	33	28.84	1.51	26.4	5.4	24.4	29.8	0	0.0%
Tm	33	3.08	0.20	2.9	0.58	2.62	3.2	0	0.0%
Yb	33	15.50	0.70	14.6	3	13.3	16.3	2	6.1%
Lu	33	1.81	0.19	1.9	0.35	1.75	2.1	0	0.0%

Source: Dahrouge, 2022

Table 8-22: GRE-04 summary of REE Results Inserted During 2011-2014

Element (GRE-04)	Count	Certified Assay (ppm)	STD DEV	Mean Assay (ppm)	Range (ppm)	Min (ppm)	Max (ppm)	N outside 3 STD DEV
La	15	2735.50	53.20	2714.0	260	2610	2870	0	0.0%
Ce	15	6127.00	146.00	6101.3	360	5910	6270	0	0.0%
Pr	15	721.10	46.00	709.9	44	693	737	0	0.0%
Nd	15	2702.00	133.00	2598.7	170	2520	2690	0	0.0%
Sm	15	390.40	18.70	377.9	24	366	390	0	0.0%
Eu	15	100.56	9.15	93.6	8	89.7	97.7	0	0.0%
Gd	15	232.80	15.10	200.7	24	191	215	0	0.0%
Tb	15	24.45	1.61	20.6	2.1	19.8	21.9	0	0.0%
Dy	15	96.52	6.36	81.9	11.2	76.9	88.1	1	6.7%
Ho	15	13.52	1.03	11.3	1.6	10.7	12.3	0	0.0%
Er	15	28.56	1.81	23.9	4.7	21.6	26.3	4	26.7%
Tm	15	3.01	0.21	2.6	0.43	2.42	2.85	0	0.0%
Yb	15	15.02	0.74	13.6	1.5	13	14.5	0	0.0%
Lu	15	1.76	0.20	1.7	0.22	1.56	1.78	0	0.0%

Source: Dahrouge, 2022

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Source: Dahrouge, 2022

Figure 8-18: CRM GRE-03 and GRE-04 Control Charts for La and Ce

 

Source: Dahrouge, 2022

Figure 8-19: CRM GRE-03 and GRE-04 Control Charts for Pr and Nd

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Figure 8-20: CRM GRE-03 and GRE-04 Control Charts for Dy

Table 8-23: Oreas 460 REE Results (2021) Summary of Results

Element (Oreas 460)	Count	Certified Assay Value (ppm)	STD DEV (ppm)	Mean Assay (ppm)	Range (ppm)	Min (ppm)	Max (ppm)	N outside 3 STD DEV
La	29	1369.00	75.00	1369.0	170	1280	1450	0	0.0%
Ce	29	1798.00	72.00	1831.0	280	1670	1950	0	0.0%
Pr	29	244.00	8.00	243.9	40	221	261	0	0.0%
Nd	29	781.00	47.00	827.3	131	754	885	0	0.0%
Sm	29	107.00	3.00	108.2	17.1	99.9	117	1	3.4%
Eu	29	22.70	0.96	23.1	3.7	20.9	24.6	1	3.4%
Gd	29	50.00	3.00	46.7	8.6	41.9	50.5	0	0.0%
Tb	29	4.84	0.21	4.8	1	4.2	5.2	0	0.0%
Dy	29	19.80	0.75	19.6	3.4	18	21.4	0	0.0%
Ho	29	2.77	0.22	2.7	0.6	2.4	3	0	0.0%
Er	29	6.01	0.35	5.7	1.6	4.8	6.4	2	6.9%
Tm	29	0.70	0.05	0.7	0.16	0.59	0.75	0	0.0%
Yb	29	3.91	0.26	3.6	0.8	3.3	4.1	0	0.0%
Lu	29	1.69	0.10	0.5	0.14	0.46	0.6	0	0.0%

Source: Dahrouge, 2022

Table 8-24: Oreas 464 REE Results (2021) Summary of Results

Element (Oreas 464)	Count	Certified Assay Value (ppm)	STD DEV (ppm)	Mean Assay (ppm)	Range (ppm)	Min (ppm)	Max (ppm)	N outside 3 STD DEV
La	26	11700.00	320.00	11738.5	1800	10800	12600	0	0.0%
Ce	26	15300.00	530.00	15288.5	2200	14000	16200	0	0.0%
Pr	26	2597.00	106.00	2496.2	330	2300	2630	0	0.0%
Nd	26	9940.00	320.00	9470.4	1080	8870	9950	2	7.7%
Sm	26	1498.00	38.00	1484.2	230	1370	1600	1	3.8%
Eu	26	324.00	8.00	322.7	53	294	347	1	3.8%
Gd	26	676.00	30.00	628.7	93	593	686	0	0.0%
Tb	26	54.00	2.70	52.4	9.4	49	58.4	0	0.0%
Dy	26	178.00	8.00	173.3	28	157	185	0	0.0%
Ho	26	21.30	1.34	20.2	2.9	18.5	21.4	0	0.0%
Er	26	38.20	1.30	35.0	6.1	32	38.1	9	34.6%
Tm	26	3.56	0.18	3.3	0.74	2.97	3.71	2	7.7%
Yb	26	15.70	0.99	15.0	4.4	12.8	17.2	0	0.0%
Lu	26	1.69	0.10	2.0	0.53	1.72	2.25	0	0.0%

Source: Dahrouge, 2022

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Source: Dahrouge, 2022

Figure 8-21: CRM OREAS 460 and OREAS 464 La, Ce, Pr, and Nd, and Dy Control Chart

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Source: Dahrouge, 2022

Figure 8-22: CRM OREAS 460 and OREAS 464 Dy Control Chart

8.3.3.2.5  Field Pulp Duplicates

A second riffled sample split of 555 Nb₂O₅ pulp duplicate samples and 555 TiO₂ and Sc pulp duplicate samples comprising, taken after pulverization, were sent to Actlabs as part of the routine sample submission from diamond drilling samples, which represent ~4.9% of total sample submissions from the 2011 and 2014 drilling program. The results are shown in Figure 8-23, Figure 8-24 and Figure 8-25, and indicate a reasonable comparison between the original and duplicate assays. All REE’s were evaluated, charted, and classified as reasonable comparisons, during this review and the targeted REE’s element charts were presented.

The base statistics were compared for the two datasets and the difference in the mean grades found to be -0.03 for Nb₂O₅ (%), -0.37 for TiO₂ (%), 1.08 Sc (ppm), 0.09 La (ppm), 0.63 Ce (ppm), 0.14 Pr (ppm), 0.11 Nd (ppm), and 0.08 Dy (ppm), which indicates an acceptable level of precision at the laboratory. A review of the 555 Pulp duplicate samples, a total of 3 Nb₂O₅, 3 TiO₂, and no Sc results fell outside the targeted 20% relative difference boundary.

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Source: Dahrouge, 2022

Figure 8-23: Paired Relative Difference and an XY Scatter Comparison of Original Versus Pulp Duplicate (Riffle Split) Analysis for Nb₂O₅ (%), Sc (ppm), and TiO₂ (%)

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Source: Dahrouge, 2022

Figure 8-24: Paired Relative Difference and an XY Scatter Comparison of Original Versus Pulp Duplicate (Riffle Split) Analysis for LREE, La (ppm), Ce (ppm), Pr (ppm), and Nd (ppm)

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Source: Dahrouge, 2022

Figure 8-25: Paired Relative Difference and an XY Scatter Comparison of Original Versus Pulp Duplicate (Riffle Split) Analysis for HREE Representative, Dy (ppm)

Following a review of all the data available using primarily Paired Relative Difference and XY Scatter plots, the conclusion is that no significant issues regarding the precision exist from the Actlabs assays in the database. All phases of the sample preparation display strong correlations between the original and duplicate assays.

8.3.3.3    Reject Duplicates

A second riffled split sample of 346 reject duplicate samples, taken after crushing, were sent to Actlabs for analysis (blind) as part of the routine sample submission from diamond drill hole samples, which represent ~3% of the total sample submissions from the 2011 and 2014 drilling program. The results are shown in Figure 8-26, Figure 8-27 and Figure 8-28, and indicate a reasonable comparison between the original and duplicate assays.

The base statistics for the two datasets were compared and the difference in the mean grades were found to be 0.00 for Nb₂O₅ (%), 0.00 for TiO₂ (%), -0.15 Sc (ppm), -5.29 La (ppm), -6.73 Ce (ppm), -0.34 Pr (ppm), -2.64 Nd (ppm), and -0.35 Dy (ppm), which indicates an acceptable level of precision at the laboratory. A review of the 346 reject duplicate samples, a total of 4 Nb₂O₅, 6 TiO₂, and no Sc results fell outside the targeted 20% relative difference boundary.

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Source: Dahrouge, 2022

Figure 8-26: Paired Relative Difference and an XY Scatter Comparison of Original Versus Coarse-Reject Duplicate (Riffle Split) Analysis for Nb₂O₅ (%), Sc (ppm), and TiO₂ (%)

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Source: Dahrouge, 2022

Figure 8-27: Paired Relative Difference and an XY Scatter Comparison of Original Versus Coarse-Reject Duplicate (Riffle Split) Analysis for LREE, La (ppm), Ce (ppm), Pr (ppm), and Nd (ppm)

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Figure 8-28: Paired Relative Difference and an XY Scatter Comparison of Original Versus Coarse-Reject Duplicate (Riffle Split) Analysis for HREE Representative, Dy (ppm)

Following a review of all the data available using primarily Paired Relative Difference and XY Scatter plots, the conclusion is that no significant issues regarding the precision exist from the Actlabs assays in the database. All phases of the sample preparation display strong correlations between the original and duplicate assays.

8.3.3.4    Field 1/4 Core Duplicates

A total of 507 field duplicate samples comprised of 1/4 core were resubmitted to Actlabs as part of the routine sample submission from DDH samples, which represent 4.5% of total sample submissions from the 2011 and 2014 drilling program. The results are shown in Figure 8-29, Figure 8-30, and Figure 8-31, and indicate a reasonable comparison between the original and duplicate assays.

The base statistics for the two datasets were completed and the difference in the mean grades found to be 1.3% for Nb₂O₅, 1.0% for TiO₂, and 0.4% for Sc (ppm), -29.69 La (ppm), -31.57 Ce (ppm), -3.16 Pr (ppm), -9.23 Nd (ppm), and -0.25 Dy (ppm), which indicates an acceptable level of precision at the laboratory.

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Figure 8-29: Paired Relative Difference and an XY Scatter Comparison of Original Versus Quarter (1/4-Core) Core Duplicate (Riffle Split) Analysis for Nb₂O₅ (%), Sc (ppm), and TiO₂ (%)

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Source: Dahrouge, 2022

Figure 8-30: Paired Relative Difference and an XY Scatter Comparison of Original Versus Quarter (1/4-Core) Core Duplicate (Riffle Split) Analysis for LREE, La (ppm), Ce (ppm), Pr (ppm), and Nd (ppm)

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Source: Dahrouge, 2022

Figure 8-31: Paired Relative Difference and an XY Scatter Comparison of Original Versus Quarter (1/4-Core) Core Duplicate (Riffle Split) Analysis for HREE Representative, Dy (ppm)

Following a review of all the data available using primarily Paired Relative Difference and XY Scatter plots, the conclusion is that no significant issues regarding the precision exist from the Actlabs assays in the database. All phases of the sample preparation display strong correlations between the original and duplicate assays expected from quarter core duplicates.

8.3.3.5 Third-Party Duplicate Check Analysis

During the 2011 and 2014 drill programs third-party validations were completed at Inspectorate and SGS Laboratories. Analytes tested at SGS external check programs focused on Nb₂O₅, Sc, TiO₂ and did not include REE elements. The lack of external laboratory REE analysis limited their evaluation to CRM and pulp duplicate review.

In 2011, a second riffled sample split of 82 Nb₂O₅ pulp duplicate samples were generated by Actlabs after pulverization and submitted to Inspectorate Laboratories as part of the routine external laboratory sample submission, which represent ~4.6% of total sample submissions from the 2011 drilling program. The results are shown in Figure 8-32 and indicate a reasonable comparison between the original and duplicate assays. A bias high has been identified in the Actlabs results relative to the Inspectorate results, for Nb₂O₅, and is defined by an average 7.2 % increase in relative difference (Figure 8-32).

Source: Dahrouge, 2022

Figure 8-32: Inspectorate Labs (2011) Paired Relative Difference and an XY Scatter Comparison of Original Versus External Pulp Duplicate (Riffle Split) Analysis for Nb₂O₅ (%).

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A total of 462 pulp duplicate samples comprising a second riffled sample split of pulverized material, taken at the same time of extraction as the primary pulps, were submitted as part of the routine sample submission to a check laboratory (SGS). The total number of samples represents approximately 5% of the original submissions.

The 2014 drill program was followed up with an additional QA/QC program which submitted an additional 150 external duplicate samples, including 101 duplicate samples and 49 standards (SX18-01, SX18-02, SX18-05, and GRE-04), selected to cover all analytical grade ranges from the 2014 drill program (Figure 8-33 and Table 8-25). The samples submitted to SGS, received a whole rock analysis that included TiO₂, with Nb₂O₅ and Sc added to the package. The program was completed to add additional duplicate QA/QC datapoints to the 2014 sample test ranges, for Nb₂O₅, Sc, and TiO₂, while evaluating the Sc specific standard performance for the 2014 drilling results. The results from this work provided confidence validation of the SGS historical Sc re-sampling results and overlapping Nb₂O₅, Sc, and TiO₂ assay results from SGS, for the 2011-2014 drill results, assayed at Actlabs.

Source: Dahrouge, 2022

Figure 8-33: 2015 Duplicate Sample Grade Range Selection Charts for duplicate re-submission, Targeting Nb₂O₅, Sc, and TiO₂.

The CRM material submitted to SGS returned assays within the range of CRM values similar to Actlabs for Nb₂O₅, TiO₂, and limited Sc (Figure 8-34, Figure 8-35, and Figure 8-36). The charts indicate that like at Actlabs, the CRMs returned values within the upper and lower limits of the assigned grades for Nb₂O₅ and TiO₂ and have a slightly improved distribution between the upper and lower limits. Both laboratories provide sufficient accuracy for Indicated Mineral Resources. The SGS dataset remains small relative to the primary Actlabs submissions and confirms the

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previously identified bias high but does not provide a large enough dataset to quantify its consistency. The identified bias falls within the accepted standard and duplicate limits.

Table 8-25: SGS (2014-2015) External Lab Sample Summary, for Nb₂O₅, Sc, and TiO₂.

SGS Check Samples	Sample Source	Nb2O5 Results	Sc Results	TiO2 Results
External Duplicate	Pulp Split	558	403	559
Blind SGS Duplicate	Pulp Split Duplicate	50	30	50
Standard Reference Material	SX18-01	36*	-	39
	SX18-02	29	-	29
	SX18-04	26	-	26
	SX18-05	10	-	10
Certified Reference Materials	GRE-4	13	13	13

Source: Dahrouge, 2022

*Only 36 of 39 Nb2O5 results for SX18-01 available, due to method variation.

Source: Dahrouge, 2022

Figure 8-34: SGS External Lab SX18-01, SX18-02, SX18-04, and SX18-05 Nb₂O_5, Control Charts

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Source: Dahrouge, 2022

Figure 8-35: SGS External Lab SX18-01, SX18-02, SX18-04, and SX18-05 TiO_2, Control Charts

Source: Dahrouge, 2022

Figure 8-36: SGS External Lab GRE-04 Nb₂O_5, Sc, and TiO₂ Control Charts

A second riffled sample split of 558 Nb₂O₅, 559 TiO₂ and 403 Sc pulp duplicate samples, taken after pulverization, were sent to Actlabs as part of the routine sample submission from diamond drilling samples, which represent ~4.9% of total sample submissions from the 2011 and 2014 drilling program. A single Nb₂O₅ sample result was missing for the dataset and not all external

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samples received Sc analysis. The results are shown in Figure 8-37 and indicate a reasonable comparison between the original and duplicate assays.

External duplicate samples submitted to SGS, were accompanied by 50 blind pulp-split duplicates, inserted with in the sample sequence and assayed for Nb₂O₅ (50 assays), TiO₂ (50 assays), and Sc (30 assays). Results were charted and indicate a reasonable comparison between the original external pulp duplicate and its blind duplicate assay.

Source: Dahrouge, 2022

Figure 8-37: SGS (2014-2015) Labs Paired Relative Difference and an XY Scatter Comparison of Original Versus External Pulp Duplicate (Riffle Split) Analysis for Nb₂O₅ (%), Sc (ppm), and TiO₂ (%)

A review of the Paired Relative Difference and XY scatter plot (Figure 8-37) for Nb₂O₅ demonstrates that Actlabs reports consistently higher Nb₂O₅ grades greater than 0.5%. This bias high has been identified in the Actlabs results relative to the SGS results, for Nb₂O₅, and is defined by an averaged 7.5 % increase in relative difference (Figure 8-34). The difference increases slightly with samples that are greater than 1%, but the dataset for samples greater then 2% is too small to determine the significance. A review of the duplicate comparisons for Nb₂O₅, TiO₂, and Sc between Actlabs and SGS does not indicate a similar high bias (Figure 8-37).

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The following details are made to explain the continued high bias:

● Certified or Standard Reference Material (CRMs or SRMs) are designed using round robin analysis, with the best value being determined from results running both above and below the best value. Values that fall within the accepted ranges are classified as acceptable and within the risk ranges assigned to that standard.

● Duplicate samples are expected to a have a degree a variability between laboratories and analytical methods. For this reason, a 20% relative difference cut off is used when comparing duplicate samples.

○ Sample reviewed remain below an average 10% relative difference, identifying a degree of risk, but falling within accepted limits.

● The SX-Series certificates do not define the method or standard deviations, presenting only 95% confidence Intervals, resulting in potential analytical method mismatching and variations resulting from precision, equipment, or method used by Actlabs.

○ For example, SGS used a borate fusion with an XRF finish, compared to Actlabs’ use of a lithium metaborate/tetraborate fusion, which Actlabs state provides improved reliability.

The continuation of NioCorp’s rigorous analytical program is recommended for all future sampling programs, with the continuation of additional external laboratory check samples, and the incorporation of a third external check laboratory (or analytical method) to compare a sample subset in triplicate.

8.4 Qualified Person’s Opinion on the Adequacy of Sample Preparation, Security and Analytical Procedures

It is Dahrouge’s opinion that the sample preparation, security, and analytical procedures used by NioCorp are consistent with standard industry practices and that the data is suitable for the 2022 Mineral Resource Estimate. Additional recommendations have been identified (see Section 23) to ensure the continuation of a robust QA/QC program but there are no material concerns with the geological or analytical procedures used or the quality of the resulting data.

8.5 Specific Gravity

NioCorp collected specific gravity (SG) measurements in both 2011 and 2014 programs, covering the spatial and temporal aspect of all drill campaigns and considering the various lithologies present. Two methodologies were implemented: (1) water immersion specific gravity measurement and (2) volumetric dry density measurement. Initially only the water immersion measurements were taken but during the 2014 SRK site inspection it was recommended that a volumetric wet- and dry-density measurements should also be taken, due to zones of porous or vuggy core intervals causing possible errors in the water-immersion method. The two methods used are described below:

Water-immersion method determines the specific gravity by the following formula:

SG = (weight in air) / (weight in air – weight in water)

A 10 to 20 cm piece of whole, dry, HQ core is weighed dry on an Ohaus Scout Pro and the weight recorded. The weight in water is determined by attaching the core by a long nylon fishing line to the Ohaus balance, lowering the core piece into a large plastic tub located immediately below the scale and filled with purified water. The weight of the core while immersed is then recorded and applied to the formula for determining the SG. Porous core samples of altered carbonatite cannot

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be accurately measured using this method and are better represented by using the dry density measurement.

Dry Volumetric method determines the Density by the following formula:

SG= [(weight in air)]/ [(π) (core length) (core radius)²]

A 10 to 15 cm piece of whole, HQ or PQ core, is dried in a convection oven for 60 minutes at 200°F. If core still has moisture, it is left in the oven for a longer period. The exact length of the core is measured with a caliper and recorded. The sample is then weighed dry in air by suspending the core by a long nylon fishing line from an Ohaus Scout Pro balance and the weight of the core recorded. It is assumed the radius remains constant for each size of drill core: 31.75 mm for HQ and 41.50 mm for PQ. These measurements are applied to the formula for determining the SG. Calibration weights were occasionally used to verify the accuracy of the balance. The table used to complete the measurements is of wood construction and tested for level by the technician.

Although 2043 specific gravity measurements were recorded, those from the 2011 program and part of the 2014 program were by water immersion only. A total of 1225 samples have been analyzed using both methods and a comparison between the two methods (Figure 8-38) shows that the water-immersion method returns higher density values therefore only 1225 volumetric measurements were used in resource estimation. A statistical analysis of the mean grades of the two populations where both methods have been recorded shows a difference of approximately 1%. The correlation shown on the XY scatter indicates a strong correlation for most cases, but for some samples there are significant differences with the volumetric density returning higher grades. This may be a result of voids or porous material.

Dahrouge does not consider the difference to have a material impact but recommends the Company continue with the use of the volumetric measurements, as the water-immersion method could result in a high bias. To confirm the density of the samples, independent analysis was completed on the geotechnical samples and results fell within the expected ranges. Additional laboratory analysis should be completed on future programs to continue to increase the confidence and generate overlapping measurement.

Source: SRK: 2015

Figure 8-38 : Comparison of Density Measurements Using Volume

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9. DATA VERIFICATION

9.1 Understood and Optimize Group Data Validation

Understood and Optimize Group conducted several validation checks on the Elk Creek Deposit for the TRS. The verification process included a one-day site visit to the Property by Qualified Persons (QPs) from both Understood and Optimize Group to review drill core, core processing protocols, QA/QC methods, collar locations drill core chain of custody and infrastructure. Additionally, database validation of lithologies and analytical results was carried out.

9.1.1 Site Visit

A site visit to the Elk Creek Property was carried out April 27, 2022, by Matt Batty, P.Geo. (Understood), Bladen Allan (Optimize Group) and Trevor Mills, P.Geo. (Dahrouge), accompanied by Scott Honan, M.Sc, SME-RM from NioCorp. The one-day site visit included:

●       Review of drill core from holes NEC11-002, NEC14-014 and NEC14-021 (Figure 9-1)

●       Visual confirmation of some drill hole collar locations (Table 9-1)

●       Review and verification of the geological setting / environment of the Project.

●       Review of drilling, logging, sampling, analytical and QA/QC procedures

●       Review of overall site facilities

9.1.1.1 Drill Core Review

The QPs completed a detailed review of three selected drillholes (NEC11-002, NEC14-014 and NEC14-02) used in the resource estimate. The core was laid out on core racks and wooden sawhorses onsite (Figure 9-1). The QPs reviewed the lithologies, structure, alteration, and mineralization observable in the drillholes. The selected drillholes provided examples of low- and high-grade material (niobium, titanium, scandium, and rare earths) and an overall sense of the Elk Creek Deposit’s geology. A comparison of the drill logs with the drill core showed that the information recorded in the drill logs matched well with the drill core.

Figure 9-1: (left) 2022 core review at the Elk Creek Project. (right) Split Core from NEC11-002

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Figure 9-2: Example of visited physical drill collars: NEC14-009 (top left), NEC-14-013 (top right), NEC14-016 (bottom left), and NEC15-005 (bottom right).

9.1.1.2    Collar Verification

The QPs visually confirmed (no GPS used) the location of a small subset of the drillhole collars (Table 9-1; Figure 9-2).

Table 9-1: Confirmed collar locations

Drillhole ID	Easting (NAD83)	Northing (NAD83)	Elevation (meters)
NEC14-009/009a	739390.23	4461466.19	349.27
NEC14-013	739169.32	4461354.33	355.17
NEC14-016	739509.06	4461574.74	354.73
NEC14-MET-02	739171.12	4461372.42	355.81
NEC15-001	739245.31	4461336.57	354.63
NEC15-004	739472.23	4461506.97	354.6
NEC15-005	739514.76	4461418.5	351.16

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9.1.1.3  Core Processing Protocols

Core processing was completed by Dahrouge during the 2011 and later work programs. As such, the QPs have relied on Dahrouge’s database to review the core logging procedures, collection of samples, and chain of custody associated with those programs. Dahrouge provided the QPs with data exports from the project drillhole database (Datamine Fusion) and electronic copies of the original assay certificates and procedural documentation. The QA/QC protocols employed by Dahrouge included the routine insertion of field duplicates, laboratory pulp duplicates, blanks, and niobium, scandium, titanium and REE certified reference standards.

No significant issues were identified during the site visit. It is Dahrouge’s opinion that the geological data collection procedures and the chain of custody were found to be consistent with industry standards and in accordance with NioCorp’s internal procedural documentation.

9.1.2 Database Validation

SRK completed an exhaustive database validation between 2014 and 2017, with Nordmin completing additional validation in 2019. Both consultants concluded that no significant issues were identified in the database, that the methods for geological data collection and QA/QC meet industry best practice standards. Transcription errors in historical survey information were noted and corrected by resurvey using DGPS.  

Data validation for the current report is summarized below. A complete review of the control samples, completed during review, identified two mis-matched quarter core and pulp duplicate samples and one mis-matched coarse reject sample in the 2011 programs. A select sample set comparison of the original assay intervals to the identified no other sample errors in the review samples. 

Table 9-2: Drill hole Assay Intervals checked against Laboratory Certificates

Historical / Re-sampled Assays	2000 samples reviewed (~22%), including 1094 (100%) of the 2021 resample program
2011 Drilling Program	180 samples (~10%)
2014 Drilling Program	960 samples (~10%)

9.1.3 Review of NioCorp QA/QC

NioCorp has a robust QA/QC process in place, as described in Section 8. Assay results were actively monitored throughout the drill programs and QA/QC results were summarized. A number of failures for standard and blank reference materials were documented, resulting in the re-assaying of entire sample batches. Most of the reference materials performed as expected within tolerances of 2 to 3 standard deviations of the mean grade. Dahrouge and Understood are satisfied that the QA/QC process is performing as designed to ensure the quality of the assay data.

9.2 Limitations

Dahrouge and Understood were not limited in access to any of the supporting data use for the resource estimation or describing the geology and mineralization in this report. The database verification is limited to the procedures described above. All mineral resource data relies on industry professionalism and integrity of those who collected and handled the database.

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9.3 Qualified Person’s Opinion

It is the opinion of Dahrouge that the geological data collection and QA/QC procedures carried out by NioCorp, are of suitable quality to support the Mineral Resource and Reserve, and they meet industry best practice standards.

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10. MINERAL PROCESSING AND METALLURGICAL TESTING

Metallurgical test work was conducted at SGS Canada Inc. (SGS), Hazen Research (Hazen) and Kingston Process Metallurgy (KPM) throughout 2014, 2015, 2016 and into 2017 to properly design the required process units for the conversion of mined ore into niobium, titanium and scandium products. The preliminary test work was performed on flotation concentrate, which has since been abandoned due to the poor recovery it offered. Test work then focused on whole ore as a feed and consisted of the extensive exploratory bench and pilot scale hydrometallurgical test programs aimed at defining and proving out a final flowsheet using different reagents and technologies. The final process flowsheet was therefore established and proven by test work and piloting performed in all the process units.

10.1 Mineral Processing

The feasibility-level comminution test work was completed in two stages at SGS Canada Inc. (SGS) in Lakefield, Ontario. The primary stage test work (SGS 2016a) was conducted on six composite samples and 13 variability samples and included:

● Bond Rod Mill Work Index (Rwi) testing.

● Bond Ball Mill Work Index (Bwi) testing.

● Bond Abrasion Index (Ai) testing.

● Bond Low-energy Impact (Cwi) testing.

● JK Drop Weight (JKDW) testing.

● Semi-autogenous grinding (SAG) Mill Comminution (SMC) testing.

The second stage of comminution test work (SGS 2016b) was conducted on a single composite sample, using a LABWAL high-pressure grinding roll (HPGR) semi-pilot scale test work program. The test work results indicate that the Project ore is categorized as soft to moderately hard in terms of ore hardness, and amenable to standard grinding as well as an HPGR operation.

A bulk representative sample (approximately 3,000 kg) of ore was subjected to locked cycle pilot scale testing at NRRI-Coleraine in Minnesota. The ore tested indicates that it is amenable to processing via the HPGR. Autogenous layer buildup and flake generation were both acceptable, and there was, on average, 40% < 1 mm product generated from the HPGR when in steady state.

The most notable observations from the testing are:

a) Final product particle size is largely independent of press force and moisture

b) Specific energy increases as both moisture and press force increase

c) There is a decrease in specific throughput as the press force increases

d) There is a decrease in specific throughput as the feed moisture increases

Based on the results as indicated above, it would be recommended to run an installed HPGR at lower pressures, i.e., 3.0 N/mm² or less, and to remove as much free water from the circuit as possible. This will have the effect of reducing power requirements with limited to no impacts on size reduction.

It is Magemi Mining’s opinion that the data as collected to date is adequate and suitable for full HPGR scale up and process guarantees around envisioned plant operation conditions.

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10.2 Hydrometallurgy

10.2.1 Testing and Procedures

Sample Representativeness

Three different whole ore samples were received and tested at SGS minerals Lakefield site, for the development of the process described in this report:

● Master Whole Ore Composite Sample (originally on site from previous test work and further described in the previous reporting at the PEA level):

○ Assay reject sample from exploration activity, all passing 10 mesh. These samples were received at SGS Lakefield from the 2014 core drilling program and were used as feed material to test the pre-feasibility in preliminary test work programs. A total of 800 kg of feed samples were processed by SGS Lakefield. A total of ten representative samples representing different areas of the mine that could be reasonably expected during production were combined into a composite sample and used as feed to the hydrometallurgical program.

● 2016 Pilot Plant Ore (received April 2016):

○ Assay reject sample from exploration activity, all passing 10 mesh. Approximately 2,640 kg of coarse reject material was processed by SGS Lakefield.

● Quarter Core (received April 2016). Approximately 1,068 kg of quarter core material was processed by SGS Lakefield.

The received quarter core sample was stage crushed to ¾” and blended. Sub-samples of the quarter core sample were further stage crushed to 100% passing 6 mesh and 10 mesh and used in the test work. Head assays of the three ore samples are summarized in Table 10-1. The assay reject based samples were very similar in composition, while the quarter core sample was slightly lower in niobium, titanium, and scandium content.

Table 10-1: Whole Ore Sample Head Assays

Feed Assays (%)*	Master Hole Ore Composite Sample	2016 Pilot Plant Ore	Quarter Core 6 mesh
Si	4.86	4.72	4.86
Al	1.14	1.09	1.23
Fe	13.5	13.2	13
Mg	5.39	5.45	5.72
Ca	12.7	13.0	13.0
Na	0.30	0.24	0.11
K	1.08	1.3	1.5
Ti	1.98	1.78	1.57
P	0.33	0.34	0.30
Mn	0.51	0.52	0.50
Cr	0.01	0.02	0.02
V	0.03	0.03	0.03
Ba	-	4.73	4.30
Y (g/t)	174	166	151
Sc (g/t)	85	82	73
S	-	1.40	1.10
Nb	0.61	0.53	0.43
LOI	24.4	25.3	26.1

*Unless otherwise stated

Source: Tetra Tech, 2017

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Hydrochloric Acid Leach (605)

There were a number of preliminary hydrochloric acid leach tests performed at the bench-scale level using different hydrochloric acid concentrations, temperature and residence times in order to confirm the leachability of the Sc and material in the ore. The operating conditions were adjusted to maximize Nb recovery as well as the Nb to Ti selectivity in the processes downstream. A pilot test program, including two pilot campaigns (PP1-013 & PP2-013), was performed.

PP1-013 ran for 80 hours while PP2-013 ran for 88 hours for a total of 168 hours and processed a total of 1,680 kg of ore samples. The objective of the Hydrochloric Acid Leach pilot circuit was to leach out impurities and a large portion of the scandium from the ore while maintaining conditions to maximize Nb recovery as well as Ti selectivity. The pregnant leach solution (PLS) from the whole ore pre leach (WPL) circuit was collected for future test work aimed at the recovery of the leached scandium. The remaining solids were collected for future test work of downstream circuits aimed at recovering niobium, titanium, and unleached scandium.

A summary of the results from PP1-013 and PP2-013 can be found in Table 10-2 and Table 10-3, respectively. A summary of the design conditions and elemental extraction chosen for the Feasibility Study can be found in Table 10-4.

Table 10-2: PP1-013 Extraction Summary

	Weight Loss %	Fe %	Mg %	Ca %	Ti %	Sc %	Nb %	Th %	U %
Average	74	84	96	99	1	68	0	45	7
Min	70	78	95	99	1	62	0	40	6
Max	77	91	97	99	2	75	1	56	8
Revstd	4	0.06	0.01	0.00	0.61	0.07	0.76	0.14	0.10

Source: SGS 2016 report “Whole Ore Pre-Leaching as Part of the Flowsheet Development for the Elk Creek Deposit Project 14379-013.”

Table 10-3: PP2-013 Extraction Summary

	Weight Loss %	Fe %	Mg %	Ca %	Ti %	Sc %	Nb %	Th %	U %
Average	76	77	95	99	1	68	0	41	8
Min	66	69	95	99	1	61	0	30	6
Max	100	87	98	100	3	82	2	57	14
Revstd	0.15	0.06	0.01	0.00	0.61	0.09	1.40	0.20	0.32

Source: SGS 2016 report “Whole Ore Pre-Leaching as Part of the Flowsheet Development for the Elk Creek Deposit Project 14379-013”)

From the above results, a set of design conditions were obtained. Table 13-4 shows the hydrochloric acid leach design basis, in terms of extraction to the PLS.

Table 10-4: HCI Leach - Summary Design & Extraction Extent

Si	0	%
Al	19.1	%
Fe	81.9	%
Mg	95.3	%
Ca	99.1	%

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Na	20.6	%
K	18	%
Ti	0.8	%
P	77.9	%
Mn	99.4	%
Ba	0.0	%
Sc	62.5	%
Sr	99.1	%
Nb	0.0	%
U	4.3	%
Th	40.5	%
Zr	36.0	%
Cr	6.2	%
V	58.3	%

Source: Tetra Tech, 2017

Acid Bake (610) and Water Leach (615)

The residues from the Hydrochloric Acid Leach testing were used in a series of Acid Bake tests, directed to extracting the niobium after sulphation using sulphuric acid at elevated temperature. Following the preliminary bench-scale test work, two pilot campaigns were performed on the Acid Bake. A first campaign (PP1-015) was performed using a pug mill followed by a rotary kiln feeding into the Water Leach. Excessive abrasion and corrosion wear on the pug mill due to inappropriate material of construction forced the second campaign (PP2-015) to be operated with batch acid mixing followed by a continuous run of rotary kiln feeding into the Water Leach. Metal extraction is different between a batch and continuous operation. For this reason, PP2-015 results were excluded from the design data of the Acid Bake and Water Leach. PP1-015 ran over the course of 103 hours and produced a total of 159 kg of solids fed to Water Leach. Table 10-5 shows a summary of the results from PP1-015 Acid Bake and Water Leach.

The results from PP1-015 were used to derive the design conditions. A summary of the design conditions and elemental extraction chosen for the Feasibility Study can be found in
Table 10-6.

Table 10-5: PP1-015 Acid Bake and Water Leach Extractions

	Weight Loss %	Si %	Fe %	Mg %	Ca %	Ti %	P %	Sc %	Nb %	Th %	U %
Average	55%	0.12	84	97	89	85	90	90	92	90	97
Min	50%	0.08	77	95	84	80	84	85	88	59	96
Max	61%	0.15	89	99	92	89	92	93	95	96	98
Revstd	7%	0.21	0.04	0.02	0.03	0.03	0.03	0.03	0.02	0.13	0.01

Source: SGS 2017 report “An Investigation into Flowsheet Development for the Elk Creek Flowsheet — Acid Bake Through to Titanium Precipitation for the Elk Creek Deposit SGS Project No.: 14379-015

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Table 10-6: Acid Bake and Water Leach - Extraction Results

Hydrochloric Acid Leach Residue
AB Temperature	300	°C
AB Acid Ratio	0.775	t/t
WL Temperature	95	°C
Fe	79.8	%
Mg	97.4	%
Ca	91.9	%
Ti	87.3	%
Nb	93.8	%
U	97.1	%
Th	95.5	%

Source: Tetra Tech, 2017

Integrated Operation Iron Reduction (620), Niobium Precipitation (625) and Titanium Precipitation (635)

The water leach liquors were processed in a series of preliminary tests in the Iron Reduction, Niobium Precipitation and Titanium Precipitation aimed at producing and confirming the operating conditions to be used in the pilot campaign. An integrated pilot plant (PP3-015) was operated continuously from August 28 to September 2, 2016, providing approximately 111 hours of operation. The campaign processed water leach liquor that was produced in the PP1-015 and PP2-015 campaigns.

Iron Reduction (620)

A summary of the design conditions is provided in Table 10-7.

Table 10-7: Iron Reduction

Temperature	Ambient	°C
Residence Time	1.25	H
Fe addition	1.125
Ratio of Briquettes and Powder	90:10

Source: Tetra Tech, 2017

Niobium Precipitation (625)

Additional pilot-scale tests on Iron Reduction (IR) and NbP were conducted in November 2016 to investigate different physical aspects of the process, and follow-up pilot plant campaigns were conducted in December 2016 and March 2017. All test work was carried out under continuous operating conditions.

The feed to the continuous NbP test campaigns was produced by processing Hydrochloric Acid Leach residue from an earlier pilot campaign (PP1-013) which was processed through Acid Bake and Water Leach steps. The Water Leach filtrate was then advanced to continuous testing of the Iron Reduction (IR) and NbP processes. The conditions for IR were left unchanged from the latest bench and pilot work. Based on the established conditions, a full pilot campaign (PP1-018) was conducted over five days in December 2016. The overall dilution ratio was decreased to 2 (effective ratio of ~2.3 based on flow differential) resulting in a drop from 96% to 94% niobium

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recovery. Following the March 2017 pilot campaign (PP2-018) calculations and investigations confirmed that a 91.51% extraction of Nb could be achieved at a dilution ratio of 0.6:1. To conservatively design the plant, two additional tanks were included to increase the residence time. While being lower than the 96% at 5:1 or the 94% at 2:1, a reduced dilution ratio of 0.6:1 is preferable by greatly reducing the equipment size and the reagent consumption.

Figure 10-1 shows the trends in Nb and Ti recoveries in the NbP as a function of the dilution ratio. The trend shows that the 0.5-0.8 dilution ratio is in the greatest inflection portion of the curve. A regression based on the results provides the required verification that 91.51% recovery can be achieved with a 0.6:1 dilution ratio.

Source: SGS 2017 report “An Investigation into Flowsheet Development for the Elk Creek Flowsheet — Acid

Bake Through to Titanium Precipitation for the Elk Creek Deposit SGS Project No.: 14379-015

Figure 10-1: Average Estimated Precipitation Versus Dilution Ratio

The results from the March campaign were used to derive the design conditions. A summary of the design conditions and elemental extractions is provided in Table 10-8.

Table 10-8: Niobium Precipitation - Elemental Extractions

Temperature	100	°C
Residence Time	4	H
Si	0	%
Al	0.3	%
Fe	0.1	%
Mg	0.3	%
Ca	0.3	%
Na	0.0	%
K	0.7	%
Ti	53.6	%
Mn	5.4	%

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Cr	16.1	%
V	12.4	%
Ba	67.6	%
Sc	4.2	%
S	1.2	%
U	37.0
Th	8.5
Nb	91.5	%

Source: Tetra Tech, 2017

Caustic Leach — Phosphate Removal

The Niobium Precipitates were used in a series of caustic leach tests, aimed at confirming the process for reducing the phosphate concentration in the final Niobium Precipitate. Ten caustic leach tests investigated a selection of NaOH solutions at various concentrations, temperatures and contact times. A summary of the retained design conditions is presented in Table 10-9.

Table 10-9: Phosphate Removal - Summary Extractions

Si	37.2	%
Al	14.4	%
Mg	0.09	%
Ti	0.04	%
P	95.1	%
Mn	0.2	%
Cr	0.5	%
V	28.2	%
Nb	0.2	%
Sr	100	%
K	57.4	%
S	93.7	%

Source: Tetra Tech, 2017

Titanium Precipitation (635)

A continuous neutralization circuit was operated from August 29 to September 2, 2016, providing approximately 86 hours of operation. The circuit processed filtrate from the NbP circuit. Feed was pumped into the circuit, and as the purpose of the pilot campaign was to test the titanium precipitation, sodium carbonate was added to lower the acidity from approximately 100 g/L sulphuric acid in the feed to between 15 g/L and 20 g/L in the discharge of the circuit with a target of 15 g/L.

The use of sodium carbonate was initially considered for the full-size plant along with magnesium carbonate but was later discarded due to the high cost of the reagent and the difficulties in regenerating. Further testing was performed using lime and limestone that confirmed a significant amount of scandium is being trapped in the gypsum formed by the partial neutralization of the NbP filtrate. To counter the significant scandium losses and the high cost of reagent, a “Calcium Loop” was designed that uses a small amount of fresh lime with recycled lime to partially neutralize the NbP filtrate. A purge back to HCI Acid Leach controls the scandium concentration inside the loop and recovers any elements that may have been trapped. This loop uses a calciner

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in the same fashion as with the sulphates coming from the Acid Regeneration and the Tailings Neutralization to regenerate and recycle lime out of the gypsum. As these were tested in continuous pilot scale operation, no further piloting was performed on the “Calcium Loop” calciner.

The neutralization was followed by a continuous Titanium Precipitation circuit that was also operated from August 29 to September 2, 2016, providing approximately 81 hours of operation. Table 10-10 shows the design basis of the TiP in terms of elemental extraction.

Table 10-10: Titanium Precipitation - Elemental Extractions

Si	18.9	%
Al	0.9	%
Fe	1.2	%
Mg	0.1	%
Ca	0.6	%
Na	0.0	%
K	0.6	%
Ti	93.5	%
Cr	3.6	%
V	5.2	%
Sc	0.8	%
U	6.4	%
Th	4.3	%
Nb	76.1	%

Source: Tetra Tech, 2017

Scandium Precipitation (628)

The Titanium Precipitation filtrate solution was used in a series of scandium precipitation tests. The limited quantity of scandium contained in the filtrate restricted the number and size of test programs. In all nine bench-scale tests and two bulk campaigns were performed. A total of 659 L of the filtrate were treated, producing a combined 970 g of the precipitate.

A summary of the results can be found in Table 10-11. From these tests and bulk campaigns, design conditions were determined. A summary of the retained design conditions is presented in Table 10-12.

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Table 10-11: Scandium Precipitation Summary

Test ID	ScTiP1	ScTiP2	ScTiP3	ScTiP4	ScTiP5	ScTiP6	ScTiP7	ScTiP8	ScTiP9	ScTiP10	ScTiP11
Iron Powder Added	Yes	Yes	No	No	Yes	No	Yes	Yes	Yes	Yes	Yes
Phosphoric Acid Added	Yes	No	Yes	No	Yes	Yes	Yes	Yes	Yes	Yes	Yes
Reagent	MgCO3	MgCO3	MgCO3	MgCO3	Ca(OH)2	Ca(OH)2	MgCO3	MgCO3	MgCO3	MgCO3	MgCO3
Reagent Addition, kg/m3	17.2	17.8	20.0	18.1	16.6	57.4	14.9	14.9	13.6	13.2	15.6
Iron Powder Addition, kg/m3	2.2	4.0	-	-	5.1	-	9.1	10.8	9.7	9.4	3.3
Phosphoric Acid Addition, kg/m3	1.4	-	1.4	-	1.4	3.5	1.4	1.0	0.7	0.3	1.4
Test Temperature, °C	75	75	75	75	75	75	75	75	75	75	75
Final Target Pulp pH	3.35	4.00	4.00	4.00	3.25	3.25	3.25	3.25	3.25	3.25	3.25
Final Filtrate pH	2.88	3.95	3.71	4.00	2.92	3.13	2.83	2.78	2.87	2.71	2.70
Precipitation (%)	ScTiP1	ScTiP2	ScTiP3	ScTiP4	ScTiP5	ScTiP6	ScTiP7	ScTiP8	ScTiP9	ScTiP10	ScTiP11
Sc	93.22	12.09	95.85	-	96.02	97.06	95.66	96.59	96.25	84.92	96.45
Al	70	-	-	59	86	59	77	67	57	19	66
Fe	3	3	18	13	4	4	1	1	2	1	2
Mg	-	-	-	-	1	2	-	-	-	-	-
Ca	1	1	3	2	88	97	1	1	1	1	1
Na	-	-	-	-	-	4	-	-	-	-	-
K	2	-	4	3	5	26	3	2	1	1	3
Ti	100	100	100	100	100	100	100	100	100	100	100
P	53	34	99	31	80	99	61	57	76	65	51
Mn	-	-	1	3	22	43	2	1	1	1	1
Cr	83	88	99	98	95	98	95	76	74	35	92
V	15	97	93	96	34	82	55	17	13	4	19
Th	98	84	100	73	100	100	100	100	100	100	100
Zr	-	-	-	-	-	-	65	59	56	49	95
S	-	-	-	-	-	-	100	-	-	-	-
Nb	-	-	-	-	-	-	100	94	93	91	-

Source: Tetra Tech, 2017

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Table 10-12: Scandium Precipitation - Elemental Extractions

Criteria	Value	Unit
Iron Addition	2.9	kg/m3
Phosphoric Acid Addition	4.5	kg/m3
MgCO3 Addition	15.3	kg/m3
Final pH target	3.25	-
Sc	96.4	%
Ti	100.0	%
Zr	100.0	%
Nb	95.0	%
Th	99.7	%

Source: Tetra Tech, 2017

Sulphate Calcining and Mixed Oxides Handling (630)

Sulphate calcining was tested at three facilities all using the Hydrochloric Acid Regeneration solids. Solids were initially calcined at SGS using equipment available. The wet filter cake from the Hydrochloric Acid Regeneration pilot plant was processed through a rotary kiln at a temperature of 1100°C at SGS. The temperature limit was set by the available equipment and not by the process requirement. At such a temperature, it was expected that the Calcium sulphate in the feed material would not be converted, while the free sulphuric acid and any iron sulphate or magnesium sulphate would be converted. In bench tests under similar conditions, typically 70-80% total sulphur removal was noted. The continuous pilot plant operated for 91.5 hours. Overall, the pilot was successful, resulting in 53 kg of calcine produced (from 183 kg of wet filter cake fed). Total sulphur removal was calculated to be 80%, as expected. The calcined product from the pilot at SGS was primarily gypsum with associated iron and magnesium oxides. It was, however, desired to convert the remaining sulphur associated with the gypsum. This led to test work being performed at Hazen Research in Golden, Colorado as well as at Kingston Process Metallurgy Inc. in Kingston Ontario Canada.

Test campaigns were initiated at Hazen using calcined material from the pilot campaign at SGS Lakefield. Conditions were set to provide a reducing atmosphere. This allowed for the better conversion of the gypsum into an oxide mix at a lower temperature than is typically required for gypsum under an oxidizing atmosphere.

Bench-scale tests were initiated followed by a bulk pilot campaign processing a total of 45 kg of calcined material producing 14.5 kg of mixed oxide material and 6 kg of kiln cleanout material containing mixed oxides and a small amount of etched refractory.

Table 10-13 provides a summary of the results of both the bench scale tests and the pilot campaign. The pilot campaign showed that gypsum was essentially wholly converted to oxides with only a trace amount of sulphur left in the mixed oxide material.

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Table 10-13: Sulphate Calcining Results Summary

Mineral Identified	Selected Sample Composition, wt%
	ARC Solids Bulk Shipment 54764	SGS Arc-1 (bucket sample) (reed w/bulk)	Bulk Process Sectional Kiln Average Calcine	Bulk Process Sectional Kiln 3 Larger Aggloms.	Batch Kiln Calcine
					BK-1	BK-2	BK-3	BK-4 c/
Anhydrite (CaSO4)	67	65	< 1	nd	< 1	25	12	< 1
Magnetite (Fe3O4)	25	28	8	13	nd	nd	nd	nd
Hematite (Fe2O3)	< 1	nd	nd	nd	nd	nd	nd	nd
Periclase (MgO)	7	7	16	6	29	16	18	2
Srebrodolskite (Ca2Fe2O5)	nd	nd	42	14	44	44	51	52
Lime (CaO)	nd	nd	3	nd	18	9	11	7
Oldhamite (CaS)	nd	nd	< 1	nd	5	< 1	2	15
Portlandite (Ca(OH)2)	nd	nd	nd	< 1	3	< 1	1	8
Corundum (Al2O3)	nd	nd	3	6	nd	nd	nd	nd
Quartz (SiO2)	nd	nd	< 1	nd	nd	nd	nd	nd
Larnite (Ca2SiO4)	nd	nd	22	18	nd	nd	nd	nd
Merwinite (Ca3Mg(SiO4)2)	nd	nd	4	17	nd	nd	nd	nd
Jasmundite (Ca10.5Mg0.5Si4SO16)	nd	nd	nd	nd	nd	nd	nd	15
Mulite (3Al2O32SiO2)	-	-	-	7	-	-	-	-
Nepheline ((Na,K)AlSiO4)	-	-	-	4	-	-	-	-
Sodium Aluminum Silicate (NaAlSiO4)	-	-	-	6	-	-	-	-
Gehlenite (Ca2Al(AISi)07)	-	-	-	< 1	-	-	-	-
Ilvaite (CaFeSilicate)	-	-	-	6	-	-	-	-
Other Results
Total Sulphur	16.2	14.8	0.8	0.5	2.2	6.3	4.2	7.7
Neutralization Potential, kg HCl/ton sample b/	na	110	na	na	809	502	624	na

Source: Tetra Tech, 2017

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Scandium Extraction (640)

The pre-leach liquors were treated in a series of scandium extraction (ScSx) tests and a pilot campaign, aimed at confirming the process for extracting scandium from leach liquors. Within the overall flowsheet, 62.5% of the available scandium is dissolved in the Hydrochloric Acid Leach circuit at design conditions. Scandium that is not dissolved in the HCI Leach circuit is further recovered and eventually is combined in the Pregnant Leach Liquor and fed to the solvent extraction circuit. Two successful and separate solvent extraction pilot plant campaigns (PP2 and PP3) were performed and the circuit operated for a total 215 hours, and approximately 3800 L of PLS was processed, producing 118 g of scandium solids containing on average 42.9% Sc (77.7 g of Sc₂O₃ equivalent content).

The extractant used was Di-(2-ethylhexyl) phosphoric acid (D2EHPA) prepared with tridecanol (as modifier) in Orfom SX80.

Scandium extraction averaged more than 99% throughout the two campaigns with scandium levels in the raffinate consistently below the analytical detection limit (<0.07 mg/L). Thorium and iron extractions averaged 0.1%. Titanium extraction extent averaged 93 and 95%, which showed that no selectivity against titanium took place in the extraction circuit. However, titanium was efficiently removed in the scrub circuit. A single wash stage was included in which more than 55% of the iron and more than 20% of the thorium was removed from the loaded organic. Some of the titanium, 7%, was also removed; less than 0.1% scandium was removed in this circuit. The wash liquor was mixed with the PLS and put back into the extraction circuit.

The scrubbing circuit was designed to remove the remaining iron, thorium and titanium from the washed organic using a solution of H₂SO₄ and H₂O₂. More than 99% of the iron was removed from the washed organic. Thorium in the scrubbed organic was below the detection limit (<2.5 mg/L) throughout the campaigns. Titanium and scandium scrubbing reached 98% and 7%, respectively. The scrubbed scandium is combined with the Ti Precipitation feed liquor, where titanium is recovered as TiO₂ and scandium is subsequently recovered via precipitation from the Ti Precipitation filtrate and brought back into the Scandium SX circuit as impure re-leach liquor.

The strip circuit used a NaOH solution to strip the scandium from the scrubbed organic and at the same time precipitate scandium hydroxide. The aqueous phase was sent to a filter to remove the scandium solids and recycle the strip liquor as strip feed (after adjusting its NaOH concentration).

The overall recovery of scandium to the solids, ranged from 93% to 95% at the beginning of the PP2 campaign and 91% during PP3. Approximately 8% of the total scandium reported to the scrub liquor. This scandium fraction is recovered in the Scandium Precipitation circuit (628). Therefore, overall scandium recovery in the Solvent Extraction (to solids and to the scrub liquor) is greater than 99%. Table 10-14 and Table 10-15 provide a summary of the overall deportment of metals in the Scandium Solvent Extraction unit pilot campaigns PP2-014 and PP3-014.

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Table 10-14: PP2 Overall Metal Distribution

Stream	Sc %	Th %	Fe %	Ti %	Ca %	Mg %	Mn %	Al %	Ba %	Be %	Co %	Cr %	K %	Na %	Ni %	P %	Sr %	V %	Zn %
Raffinate	1	100	100	8	100	100	100	100	100	99	99	100	100	77	100	100	100	100	100
Scrub Sol’n	8	0	0	88	0	0	0	0	0	1	1	0	0	0	0	0	0	0	0
Conditioning	0	0	0	0	0	0	0	0	0	0	0	0	0	22	0	0	0	0	0
Strip Solution	91	0	0	4	0	0	0	0	0	0	0	0	0	0	0	0	0	0	0
Total	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100
Stream	Sc	Th	Fe	Ti	Ca	Mg	Mn	Al	Ba	Be	Co	Cr	K	Na	Ni	P	Sr	V	Zn
Raffinate	1	100	100	8	100	100	100	100	100	99	99	100	100	77	100	100	100	100	100
Scrub Sol’n	8	0	0	88	0	0	0	0	0	1	1	0	0	0	0	0	0	0	0
Conditioning	0	0	0	0	0	0	0	0	0	0	0	0	0	22	0	0	0	0	0
Strip Solution	91	0	0	4	0	0	0	0	0	0	0	0	0	0	0	0	0	0	0
Total	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100

Source: Tetra Tech, 2017

Table 10-15: PP3 Overall Metal Distribution

Stream	Sc %	Th %	Fe %	Ti %	Ca %	Mg %	Mn %	Al %	Ba %	Be %	Co %	Cr %	K %	Na %	Ni %	P %	Sr %	V %	Y %	Zn %
Raffinate	0	100	100	5	100	100	100	100	100	99	99	100	100	75	100	100	100	100	100	100
Scrub Sol’n	8	0	0	91	0	0	0	0	0	0	1	0	0	0	0	0	0	0	0	0
Conditioning	0	0	0	0	0	0	0	0	0	0	0	0	0	25	0	0	0	0	0	0
Strip Solid	92	0	0	3	0	0	0	0	0	0	0	0	0	0	0	0	0	0	0	0
Total	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100	100

Source: Adapted from SGS 2017 report “An Investigation into An Integrated Scandium Solvent Extraction Pilot Plant for The Elk Creek Project 14379-014

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Scandium Refining (645)

The scandium hydroxide [Sc(OH)₃] produced in the scandium solvent extraction test work was used in the steps to test the scandium refining portion of the process. A number of leaches both in HCl and H₂SO₄ were performed with the intent of testing the removal of impurities such as titanium and niobium from the scandium produced. HCl was initially tested but was rejected due to poor results. H₂SO₄ was retained in combination with a solvent extraction step. A series of Sc(OH)₃ re-leaches tests using H2_SO₄ were performed in order to provide approximately 5 L of Sc rich solution. Then the solvent extraction was tested in a series of 6 campaigns treating a total of 5 L. Alamine 336 was first used and provided best results for Zr but only partially removed Ti and Nb. Aliquat 336 was then used and provided a very good result for Ti and Nb but only partially removed Zr. A mix of Alamine 336 and Aliquat 336 is finally used to get the best removal of Zr, Ti and Nb.

A summary of the retained design conditions is presented in Table 10-16.

Table 10-16: Scandium Refining - Impurity Extractions

Criteria	Value	Unit	Comment
Zr extraction	98.8	%
Ti extraction	91.3	%
Nb extraction	94.9	%
Sc extraction	99.3	%	Only traces loading

Source: Tetra Tech, 2017

The Sc rich raffinate was then mixed with an oxalic acid solution in a batch wise fashion to form scandium oxalate crystals. A series of 15 tests were performed to crystallize scandium oxalate. The initial solution used came from the initial re-leach with HCI. The solution used was later changed to a sulphate-based scandium re-leach following the development in the impurity’s removal steps. The crystallization proved to be straight forward and provided very good recovery of scandium oxalate well above 98-99%. Scandium oxalate was further filtered, washed and calcined to produce scandium trioxide with a purity of 99.9%. The limited quantity of scandium available prevented large scale testing of filtration or calcining.

Hydrochloric Acid Regeneration (660)

Using scandium solvent extraction raffinate, a five-day continuous pilot plant seeking to regenerate HCI from the raffinate using sulphuric acid (H₂SO₄) was conducted between October 18, 2016, and October 22, 2016. Two separate runs (PR1, PR2) were operated. The process was successful and showed that 99.94% of the chlorides in the feed stream were converted to HCI into the vapour phase. All the Calcium and most of the iron and magnesium were precipitated as insoluble metal sulphates, forming a filter cake of 50-60% moisture under pilot conditions.

When collecting the HCI without any water addition, the HCI content was calculated to exceed the azeotropic point of HCI, reaching 28% HCI by weight prior to the shutdown of the pilot plant. Reaching even higher concentration would have required increased cooling capacity at a lower temperature. This represents a significant opportunity to reduce the upgrading requirement prior to recycling the stream to pre-leach.

The feed solution averaged 40 g/L Fe, 24 g/L Mg, 56 g/L Ca, and 300 g/L CI, while the discharge solution from AR6 averaged 0.9 g/L Fe, 3.5 g/L Mg, 0.2 g/L Ca, and 39 mg/L Cl. This represents a

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major decrease in the amount of dissolved metal and near complete removal of chloride from the feed. Table 10-17 shows the composite solid discharge assay.

Table 10-17: PP1 Composite Discharge Solids Assays

AR6 D/C Slurry - Solids	Solids Assays (%, unless otherwise noted)
	Fe	Mg	Ca	Ti	Mn	S	CI (g/t)	SO4*
Average	8.41	3.84	12.7	0.01	0.38	22.3	190	67
Min	7.97	3.74	11.7	0.01	0.36	22.0	60	66
Max	8.74	3.93	13.2	0.01	0.39	22.7	439	68
Revstd	3%	2%	4%	0%	3%	1%	70%	1%

Source: SGS 2017 report “An Investigation into Acid Regeneration Pilot Plant Elk Creek Deposit Project 14379-016

In total, 290 kg (or 223 L) of Scandium Solvent Extraction raffinate was processed, and the pilot generated -230 kg of wet solids ranging from 40-56% solids, resulting in approximately 106 kg of dry equivalent.

Tailings Neutralization (665)

Two bulk neutralization campaigns were performed on tailings solution. In all 511 L of tailings, the solution was neutralized with limestone locally obtained from a mine in Weeping Water, NE. Neutralization proved to follow theoretical models. Tailings Neutralization solids were also calcined in bulk to provide samples that were used in the sulphate calcining campaigns performed at Hazen Research.

10.2.2 Relevant Results

A number of individual extractions were compiled to define the total recovery of each of the pay metals. A summary of the results and design conditions is shown in Table 10-18.

Table 10-18: Recovery Summary

	Recovery
	Nb From Test Work	Sc From Test Work	Ti From Test Work
HCI Leach	100.0%	100.0%	100.0%
Acid Bake - Water Leach	93.8%	97.0%	87.3%
Nb Precipitation	91.5%	98.3%	49.4%
Partial Neutralization	100.0%	100.0%	99.9%
Ti Precipitation	-	99.7%	93.5%
Sc Precipitation - Re-Leach	-	98.6%	-
Sc Solvent Extraction	-	100.0%	-
Sc Purification	-	99.3%	-
Overall	85.8%	93.14%	40.3%

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10.2.3 Significant Factors

It is the L3’s opinion that adequate test work was conducted to support a feasibility-level design for the Hydromet Plant; however, optimization was not achieved in all areas. Certain areas will certainly benefit from further “post-Feasibility Study” test work, preferably before detailed engineering activities begin. For instance, the test work performed for the Feasibility Study shows indications that several factors influence the precipitation of niobium as well as the selectivity with respect to titanium. Test work indicates that an increase in Fe/Nb ratio positively affects Nb precipitation while also promoting selectivity against Ti. Precipitant (dilution water) acidity inversely affects Nb precipitation but increases the selectivity against Ti precipitation. Final free acid titration (FAT) has very little effect on the Nb precipitation, but it greatly increases the selectivity against Ti precipitation. The above factors have not been optimized in this study, and further testing will be required to achieve optimal results. Such optimization could also be achieved with performing process simulation of the yearly or monthly elemental feed composition using the METSIM model and the compositions from the mine plan.

10.3 Pyrometallurgy

Pyrometallurgical test work was carried out at Kingston Process Metallurgy (KPM) in Kingston, Ontario, Canada, which is a non-certified but independent lab and testing company being well-known in the metallurgy field. Testing has been done using a high temperature MoSi₂ tube furnace having the capability to perform the trials at 1650^oC (3002^oF) and to operate at the atmospheric conditions. Stochiometric amount of Fe₂O₃ powder and aluminum shot have been added into the crucible containing the NbP precipitate to conduct each test performed into the furnace. A small amount of CaO has been part of the ingredients too in order to give some fluidity properties to the slag. Several tests were performed in order to ensure repetition of the aluminothermic reaction and consistency in the results for FeNb alloy chemistry.

It is the opinion of MCS that adequate test work was conducted to support a feasibility-level design for the Pyromet Plant since the Hydromet Plant analysis demonstrated higher levels of TiO₂ at the exit of the NbP section. With a higher TiO₂ content in the NbP precipitates, the purpose of the Pyromet testing was changed slightly from its original scope. The objective of the Pyromet was to produce a saleable FeNb metal, but in addition, it would play a purification role by eliminating the excess TiO₂ left by the Hydromet. The testing performed at KPM facilities was required in order to demonstrate the capability of the Pyromet to produce an acceptable quality of FeNb alloy that meets the product specifications of the Nb-steel producers. The testing performed at the KPM facilities demonstrated several points listed below:

●      The aluminothermic reduction of Nb₂O₅ precipitate to produce FeNb alloy was demonstrated regardless of the high level of TiO₂ in the precipitate.

●      Niobium recovery from the Hydromet precipitate reached 95.9%, found in the Fe-Nb alloy.

●      Only 4.1% of the Nb has been found in the slag associated with oxygen.

●      Hematite powder (Fe₂O₃) was successfully added as the iron source for the aluminothermic reduction.

As far as for a Fe-Nb pyrometallurgy plant running on daily and yearly basis, the testing has respected the thermodynamic equations, which shows the positive expectations for the future plant. However, there was no agitation in the MoSi₂ tube furnace during the aluminothermic reaction testing and this is the reason for some heterogeneity in the Fe-Nb alloy observed in the lab results. Observations using a scanning electron coupled with a spectroscopy showed the Nb-Fe alloy had a range of Nb content varying from 60% to 70%, but the average was close to the

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nominal target of 65.5% expected. The high level of TiO₂ and Al₂O₃ in the amalgam solid formed during the aluminothermic reaction prevented getting a more precise result in the Nb-Fe alloy chemical analysis, but for this stage of study, the results are aligned with the expected thermodynamic. The small amount of Nb precipitate available has limited the number of trials too, however the quantity of trials would have probably not changed the conclusion of the testing performed since they would have been performed in the same equipment. It is important to mention that the future plant will be designed with an electric arc furnace, bringing the required agitation to end it up with a more homogenous Fe-Nb alloy.

The Nb precipitates has a high TiO₂ content and the weight ratio Nb₂O₅/TiO₂ is a way lower than 1 (in fact, it is likely almost at 0.6, but with the effort putting on the hydromet processes, this ratio is fortunately going up). On a high-level production, slag drossing frequency might present some challenges, which would likely be solved by a robust dross system.

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11. MINERAL RESOURCE ESTIMATES

11.1 Introduction

The 2022 Mineral Resource Estimate for the Elk Creek Deposit was completed by Understood Mineral Resources Ltd. The effective date of the enclosed mineral resource is June 30, 2022, representing the date of when the last assay was received.

The drill hole data was provided to Understood as individual spreadsheets, including collar surveys, downhole deviation surveys, lithology logs, assay data, and specific gravity data. The database contains all 138 drill holes drilled on the property, including the 45 drill holes that define the Elk Creek Deposit, which were used to inform the 2022 Mineral Resource Estimate. A further 5 holes drilled in 2015 for hydrogeological and geotechnical testing were not assayed and did not form part of the resource estimate.

The deposit hosts niobium, titanium, scandium, and REE mineralization. Elevated concentrations of niobium, titanium, and scandium are observed within the logged magnetite dolomite carbonatite unit. The magnetic mineralization is observed to be continuous along a northwest to southeast trend with an average thickness of 200 m. The rare earth concentrations are observed to increase from southwest to the northeast, across the trend of magnetite carbonatite domain. Therefore, three wireframes were constructed for the deposit to reflect the geologic and grade observations using the available drilling data: the magnetic carbonatite domain (referred to as “Bound 1” or “MCarb”), the domain southwest of the magnetic carbonatite domain (referred to as “Bound 2” or “SW”), and the domain northeast of the magnetic carbonatite (referred to as “Bound 3” or “NE”). The NE domain has been primarily logged as a carbonatite with localization of lamprophyre dikes and the SW domain is chiefly carbonatite. The domains were clipped to the extents of the overlying sedimentary units, which was also modelled. Leapfrog Geo (version 2021.2.4) was used to create the MCarb domain and the overlying sedimentary unit wireframes. The two wireframes were exported from Leapfrog and imported into Maptek’s Vulcan (version 2021.5) for the creation of the NE and SW domains, compositing, and block estimation.

The following fields were composited at 1 m lengths with a merge tolerance of 0.5 m in the domains: Density, Nb₂O₅ (%), TiO₂(%), Sc (ppm), La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), Nd₂O₃ (%), Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), and Y₂O₃ (%).

LREO (%), MREO (%), HREO (%), and TREO (%) were also composited as a check on the individual REO (Rare Earth Oxide) composites and estimates, where:

● Light Rare Earth Oxides (LREO) variable (%) is the summation of La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), and Nd₂O₃ (%) concentrations;

● Medium Rare Earth Oxides (MREO) variable (%) is the summation of Sm₂O₃ (%), Eu₂O₃ (%), and Gd₂O₃ (%) concentrations;

● Heavy Rare Earth Oxides (HREO) grade (%) is the summation of Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), and Y₂O₃ (%) concentrations; and

● Total Rare Earth Oxide (TREO) grade (%) is the summation of LREO (%), MREO (%) and HREO (%).

The assays were capped prior to compositing, resulting in a stationary dataset for each field for each domain.

Downhole, omni-directional, and directional experimental variograms were created in Vulcan for the each of the variables in the three domains. The directional variograms were determined to be

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unstable, therefore omni-directional variograms were used. A variogram model was fitted to the experimental variograms with the nugget contribution being applied to each model as observed in the downhole variogram.

A block model was constructed to encompass the three domains using 5 m by 5 m by 5 m blocks. The blocks were populated through the use of OK using an omnidirectional variogram within the three domains. The block model was checked for data replication, including mean comparison, volumetric comparison, visual inspection, swath plots, histogram comparison, bivariate plot comparisons, and correlation checks.

The classification of the 2022 Mineral Resource Estimate has been prepared in accordance with the S-K 1300 classification system and was reported at a US$180 diluted NSR breakeven cut-off grade. The REOs were evaluated as a potential by-product to the mining of niobium, titanium, and scandium; thus, the estimated values of the REOs are reported using the diluted NSR as derived from the Nb₂O₅, TiO₂, and Sc Mineral Resources.

Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the Mineral Resource will be converted into a Mineral Reserve. The 2022 Mineral Resource Estimate is comprised of Indicated and Inferred Mineral Resources.

11.2 Source Database

The drill hole data was provided to Understood Mineral Resources Ltd. as individual spreadsheets, including collar surveys, downhole surveys, lithology logs, assay data, and specific gravity data. The database was constructed by Dahrouge from the 1980 Molycorp data and raw data captured by Dahrouge during the 2011 and 2014 drill campaigns, and the 2021 re-analysis campaign.

The collar spreadsheet comprised of 138 entries that detailed the drill hole name, the collar locations in NAD 1982 UTM Zone 14 N grid coordinates, method used to collect coordinate information (extracted from a map or from differential GPS), company that drilled the hole, and year the hole was drilled. All but one drillhole used in the 2022 Mineral Resource Estimate, EC-018 (2011), has the collar coordinates extracted from differential GPS. Notably, the location of 24 of the 29 Molycorp drill collars at the Elk Creek deposit were re-excavated to confirm the collar coordinates in 2011 by CES Group P.A. Engineers & Surveyors, based in Kansas City, Missouri.

The survey spreadsheet contains 4,164 records from 133 drill holes, which averages to a survey data point every 15.85 m of borehole. The following five drill holes from the 1980 drill campaign were not accounted for in the survey spreadsheet: CA-001, CA-002, CA-003, CA-005, and NN-1. The data discrepancy is not material to the Mineral Resource Estimate as the drill holes with no survey data are distal to the Elk Creek Deposit.

The lithology spreadsheet contains 4,065 logged intervals from 129 drill holes. The entries define the drill hole, the interval, the major logging unit, the rock code, the lithology log, and a description. The following holes contain survey data but do not have lithology data: EC-012, EC-013, EC-106, NEC11-004, and NEC11-005. Again, the data discrepancy is not material to the Mineral Resource Estimate due to distance from the Elk Creek Deposit.

The rock codes used at Elk Creek was first defined by Molycorp then simplified by Dahrouge in 2011 and 2014 for interpretation purposes. The simplified lithology codes in Table 11-1 were used in Leapfrog to create geological interpretations for estimation.

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Table 11-1: Dahrouge lithology codes

Name (Dahrouge)	Code
Glacial Till	Till
Pennsylvanian Sediments	Sed
Mudstone	Mdst
Phosphatic Shale	PhosShale
Limestone	Lmst
Carbonatite/Lamprophyre	CarbLamp
Carbonatite/Lamprophyre Breccia	CarbLampBc
Carbothermal Vein	CarbtherVn
Barite Dolomite Carbonatite	dolCarb
Dolomite Carbonatite Breccia	dolCarbBc
Lamprophyre	Lamph
Lamprophyre Breccia	LampBc
Lamprophyre/Carbonatite	LampCarb
Lamprophyre/Carbonatite Breccia	LampCarbBc
Mafic dyke, vein or fragment	maf
Mafic dyke, vein or fragment Breccia	mafBc
Magnetite Carbonatite\Lamprophyre	mCarbLamp
Magnetite Dolomite Carbonatite	mdolCarb
Magnetite Dolomite Carbonatite Breccia	mdolCarbBc
Syenite	sy
Granite	Gr

The specific gravity spreadsheet contains 2,043 measurements from 19 drill holes of the 2014 drill campaign: values range between 2.02 and 4.19. The assay spreadsheet contains 20,462 entries with 48 variable fields; the variables were heterotopically sampled, thus the dataset is unequal (Table 11-2). A value of -9 was assigned to all missing intervals. Details of procedures for specific gravity measurements are discussed in Section 8.5.

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Table 11-2: Assay variables

Variable	Unit	No. of Samples	Variable	Unit	No. of Samples	Variable	Unit	No. of Samples
Nb2O5	%	19,496	Er	ppm	14,930	Eu2O3	%	14,930
TiO2	%	13,177	Eu	ppm	14,930	Gd2O3	%	14,930
Sc	ppm	13,186	Gd	ppm	14,930	Ho2O3	%	14,930
Ag	ppm	14,930	Ho	ppm	14,930	La2O3	%	14,930
As	ppm	14,930	La	ppm	14,930	Lu2O3	%	14,930
Ba	ppm	14,930	Lu	ppm	14,930	Nd2O3	%	14,930
CaO	%	14,930	Nd	ppm	14,930	Pr2O3	%	14,930
Cr	ppm	14,930	Pr	ppm	14,930	Sm2O3	%	14,930
Fe2O3	%	13,186	Sm	ppm	14,930	Tb2O3	%	14,930
MgO	%	14,517	Tb	ppm	14,930	Tm2O3	%	14,930
Pb	ppm	14,930	Tm	ppm	14,930	Y2O3	%	14,930
Th	ppm	14,930	Y	ppm	14,930	Yb2O3	%	14,930
U	ppm	14,930	Yb	ppm	14,930	LREO	%	14,930
Zn	ppm	13,185	Ce2O3	%	14,930	MREO	%	14,930
Ce	ppm	14,930	Dy2O3	%	14,930	HREO	%	14,930
Dy	ppm	14,930	Er2O3	%	14,930	TREO	%	14,930

The REO concentrations were calculated from the individual elemental rare earth concentration prior to Understood receiving the data. As a check, the values were recalculated and compared using the conversion coefficients in Table 11-3; the assay spreadsheet values match the re-calculated values.

Table 11-3: Elemental percentage conversion ratios to oxide percentage

Oxide	Conversion
Ce2O3	1.1713
Dy2O3	1.1477
Er2O3	1.1435
Eu2O3	1.1579
Gd2O3	1.1526
Ho2O3	1.1455
La2O3	1.1728
Lu2O3	1.1371
Nd2O3	1.1664
Pr2O3	1.1703
Sm2O3	1.1596
Tb2O3	1.1510
Tm2O3	1.1421
Y2O3	1.2699
Yb2O3	1.1387

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11.2.1 Drill Holes

The drill hole data spreadsheets were imported into a Vulcan database and checked for the following:

●         Unique collar locations;

●         Overlapping assays;

●         Empty table check for assays, collars, lithology, and surveys;

●         Increasing depth field in surveys, assays, lithology, and specific gravity field;

●         Consecutive variation tolerance (max of 30 degrees) for dip and azimuth;

●         Unique sample ID for assay and specific gravity measurements;

●         Ensure azimuth survey measurements are between 0 and 360;

●         Ensure dip survey measurements are between -90 and 0;

●         Ensure Nb₂O₅ (%) grades are between -9 and 100;

●         Ensure TiO₂ (%) grade are between -9 and 100;

●         Ensure Sc (%) grade are between -9 and 100; and

●         Ensure TREO (%) grades are between -9 and 100.

There are no overlapping assays, depth errors, or gross numerical errors in recorded assay grades. There are two holes with the same collar coordinate locations (NEC14-009 and NEC14-009a), but this is not in error. NEC14-009 was abandoned during drilling, then restarted with a wedge at 485.51 m as NEC14-009a and drilled to final depth of 897.0 m. There are no duplicate assay sample IDs, but there is a duplicate sample ID for specific gravity. The specific gravity duplicate sample ID was determined to be a typo and the data was deemed acceptable.

The following twenty-two drill holes contain empty lithology, survey, and/or assay table(s): CA-001, CA-002, CA-003, CA-005, EC-012, EC-013, EC-023, EC-042, EC-046, EC-047, EC-078, EC-081, EC-085, EC-094, EC-103, EC-104, EC-106, NEC11-004, NEC11-005, NEC14-MET-01, NEC14-MET-02, and NN-1. All the listed drill holes except NEC14-MET-01 and NEC14-MET-02 are distal from the Elk Creek Deposit and therefore are not material to the Mineral Resource Estimate. NEC14-MET-01 and NEC14-MET-02 are holes dedicated to metallurgical testing and therefore were not sampled for assays, hence an empty assay table.

Not all of the drillholes within the Project were used in the Mineral Resource Estimation or included in the model database that accompanies this report, as many are located at a significant distance away from the deposit. Of the 1 total drillholes within the carbonatite complex, 48 drillholes are within the Elk Creek deposit area. An additional five of the 48 Drillholes within the Resource Area, including EC-025, EC-033, EC-035, EC-036, and EC-051, could not be included in the 2019 Nordmin or the current Resource Estimate, because they lack Sc, TiO2, and REE analytical results, preventing their incorporation into the multi-element Resource. It has been recorded that original sample material for these holes could not be located for reanalysis and because they fell at the boundaries of the deposit it was not considered priority. It is recommended that a follow-up sample search is completed given the potential for future Resource expansion and a recently noted improved organization of the historical material at the Mead storage facility, which is operated by the Conservation and Survey Division of the University of Nebraska Lincoln. Drillholes NEC14-MET-01 and NEC14-MET-02, drilled for metallurgical test work, were not assayed on an interval basis and were used for geological control, but excluded from the Resource.

There are 14 holes that exceed the azimuth survey tolerance of 30 degrees between neighbouring data locations on the same drill string, but all 14 holes are vertical holes. Vertical holes can show large apparent changes in azimuth with little true deviation; therefore, the surveys are deemed to be

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valid. There were no records that exceed the dip survey tolerance of 30 degrees between neighbouring data locations.

11.3 Geological Domaining

Three-dimensional geological interpretations were created to represent the observed geology and mineralization in the drill core and create stationary datasets to facilitate estimation. The interpretations were generated in Seequent’s Leapfrog Geo (version 2021.2.4) and Maptek’s Vulcan (version 2021.5).

Niobium, titanium, and scandium are observed in elevated concentrations within the logged magnetite dolomite carbonatite unit (Figure 11-1, Figure 11-2 and Figure 11-3). The magnetic mineralization is observed to be continuous along a northwest to southeast trend with an average thickness of 200 m. Understood is of the opinion that modeling the unit results in a geologically reasonable and statistically stationary dataset suitable for estimation. The magnetic dolomite carbonatite unit was modelled in Leapfrog by selecting the appropriate logged units to create hangingwall and footwall point clouds that defined the outer contacts. Extension distance for the wireframe was approximately 70 m laterally past the last drill intercept. The sedimentary units unconformably overlying the carbonatite was also modelled in Leapfrog by creating a point cloud from the logged contacts between the sediments and carbonatite. The magnetic dolomite carbonatite domain was subsequently clipped to the sedimentary wireframe and is referred to as the “Bound 1“ wireframe or “MCarb” (Figure 11-4). The two wireframes were exported into Vulcan.

Figure 11-1: Niobium concentration box plot by logged lithology.

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Figure 11-2: Titanium concentration box plot by logged lithology.

Figure 11-3: Scandium concentration box plot by logged lithology.

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Figure 11-4: Plan view of the MCarb domain (upper image) and cross section looking northwest of the MCarb domain and the modelled overlying sediments (lower image). Both diagrams contain the informing drill holes displaying lithology logs.

The rare earth concentrations increase from southwest to the northeast across the trend of MCarb domain (Figure 11-5). Understood decided to domain the data into the subsets to reflect the gradual increase in rare earth concentrations: the already established MCarb domain, a domain to the southwest of the MCarb domain (referred to as “Bound 2” or “SW”), and a domain to the northeast of the MCarb domain (referred to as “Bound 3” or “NE”). The SW and NE wireframes were assembled in Vulcan by tying two-dimensional cross-section polygons together

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using the drill hole data for a reference, capturing all of the drill hole data on either side of the MCarb domain and clipping the wireframes to the MCarb and overlying sedimentary unit wireframe (Figure 11-6). The NE domain is primarily carbonatite with localization of lamprophyre dikes and the SW domain is chiefly carbonatite.

Figure 11- 5: Oblique view looking northwest drill holes displaying TREO (%) assay results and outline of the MCarb domain.

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Figure 11-6: Plan view of the MCarb, SW, and NE domains (upper image) and cross section looking northwest of the MCarb, SW, and NE domains (lower image). Both diagrams contain drill holes traces displaying TREO assay results.

11.4 Exploratory Data Analysis

11.4.1 Compositing

Over 50 % of assays within the MCarb, SW, and NE domains were sampled in 1 m intervals, and 77% of the samples were sampled at lengths between 0.5 m and 1.5 m (Figure 11-7). Therefore, the

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dataset was composited at 1 m lengths with a merge tolerance of 0.5 m to maintain the variability contained within the dataset. Assays were composited in Vulcan starting at the first mineralized wireframe boundary from the collar and resetting at each new wireframe boundary. Composites less than 0.5 m, which were located at the bottom of the mineralized intercept, were added to the previous composite. Drillhole locations with no or missing values were ignored during compositing (less than 1% of data location have missing values). The following fields were composited in the MCarb, SW, and NE domains: Nb₂O₅ (%), TiO₂(%), Sc (ppm), La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), Nd₂O₃ (%), Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), and Y₂O₃ (%). LREO (%), MREO (%), HREO (%), and TREO (%). Note that the density samples were measured independently and prior to the collection of assay samples, and, consequently, the density sample intervals overlap and are at different lengths than the assay sample intervals. The density samples were included in the composite database by assigning the density sample to an assay composite if the density sample covered more than 50 % of the assay interval. If the coverage was less than 50%, the density measurement was not used. Combining the density values with the assay composites was completed to review relationships between all the assay variables and the density variable.

Figure 11-7: Histogram of assay lengths within the MCarb, SW, and NE domains.

11.4.2 Declustering

A global representative distribution of every variable is essential for unbiased resources calculation, and one step in determining a representative distribution is the consideration of the spatial arrangement of the data. Cell declustering is a widely used technique that assigns each datum a weight based on its closeness to surrounding data, where closely spaced data will receive a smaller weight than sparsely spaced data. The cell size selected for declustering of each domain was selected based on the relative spacing of the sparsely sampled areas. The MCarb domain used a 90 m x 90 m x 90 m cell, and the SW and NE domains used a 180 m x 180 m x 180 m cell.

11.4.3 Outlier Capping

The uncapped composited data with the declustered weights for each variable in each domain was reviewed in probability plots, histograms, and cartesian space for stationarity and outliers. Upon

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review, capping levels were selected for the variables in the domains, as per Table 11-4. The assays were re-composited with capping levels applied on the raw assays, and the re-composited capped dataset was reviewed again for stationarity. Figure 11-8 is an example of the review process of scandium composites in the MCarb domain. In general, the uncapped populations are relatively stable with no extreme outliers, thus the global effect of capping was minimal and only small changes in the mean and variance of the distribution were observed.

Table 11-4: Capping levels

Variable	Capping Grade
	B1 - MCARB	B2 - SW	B3 - NE
Nb2O5 %	3.6	1.8	1
TiO2 %	8.25	6.3	N/A
Sc ppm	200	160	N/A
La2O3 %	0.7	0.6	0.6
Ce2O3 %	0.9	0.8	N/A
Pr2O3 %	0.075	0.09	N/A
Nd2O3 %	0.25	0.25	N/A
Sm2O3 %	0.06	0.04	N/A
Eu2O3 %	N/A	0.0125	N/A
Gd2O3 %	0.053	0.037	N/A
Tb2O3 %	0.007	0.005	0.0035
Dy2O3 %	0.028	0.021	0.02
Ho2O3 %	0.004	0.004	0.0035
Er2O3 %	N/A	0.01	0.008
Tm2O3 %	0.001	0.0015	0.001
Yb2O3 %	0.005	0.006	0.005
Lu2O3 %	N/A	0.001	N/A
Y2O3 %	0.107	0.125	0.107
LREO %	1.8	1.5	2.5
MREO %	0.13	0.08	0.08
HREO %	N/A	0.17	0.15
TREO %	2	1.6	2.5
S.G.	4.1	N/A	N/A

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Figure 11-8: Probability plot and histogram of uncapped and capped Sc (ppm) composite distributions in the MCarb domain.

11.4.4 Representative Distributions Statement

In Understood’s opinion, the selected capping values are reasonable and, in conjunction with the declustered weights, produce representative distributions for the Elk Creek Mineral Resource Estimate. All distributions have a relatively low degree of variance around the representative mean (i.e. all co-efficients of variance (CV) are below 1.0) with a stable shape (i.e. no bimodal distributions are noted). The Nb₂O₅ data is exhaustively sampled at all data locations and less than 1% of the locations missing data for scandium, titanium, or REO.

The capped distributions with no weightings applied are used for the estimate, and the estimate is compared to the capped distribution with the declustered weightings. The summary statistics of the representative distributions, such as count, average, standard deviation, and CV, for each variable per domain are listed in Table 11-5.

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Table 11-5: Summary statistics of the variables per domain

Variable	B1 - MCARB				B2 - SW				B3 - NE
	Count	Mean	Std dev.	CV	Count	Mean	Std dev.	CV	Count	Mean	Std dev.	CV
Nb2O5%	13492	0.58	0.39	0.68	5923	0.21	0.14	0.66	1193	0.23	0.14	0.6
TiO2 %	13474	2.48	1.2	0.48	4994	1.15	1.1	0.96	1193	1.55	1.15	0.74
Sc ppm	13435	64.7	25.38	0.39	4962	33.7	17.51	0.52	1094	34.1	17.6	0.52
La2O3 %	13474	0.0908	0.057	0.63	4994	0.0701	0.068	0.97	1193	0.1317	0.099	0.75
Ce2O3 %	13474	0.1538	0.0847	0.55	4994	0.1238	0.11	0.89	1193	0.2287	0.167	0.73
Pr2O3 %	13474	0.0161	0.0081	0.5	4994	0.0133	0.011	0.85	1193	0.0252	0.019	0.77
Nd2O3%	13474	0.0581	0.0269	0.46	4994	0.0485	0.038	0.78	1193	0.0867	0.064	0.74
Sm2O3%	13474	0.0141	0.0066	0.47	4994	0.0086	0.005	0.59	1193	0.0131	0.008	0.59
Eu2O3 %	13474	0.005	0.0028	0.56	4994	0.0025	0.001	0.55	1193	0.0036	0.002	0.51
Gd2O3%	13474	0.012	0.007	0.58	4994	0.0064	0.003	0.54	1193	0.0081	0.004	0.46
Tb2O3 %	13474	0.0013	0.0008	0.63	4994	0.0008	0	0.58	1193	0.001	0.001	0.52
Dy2O3 %	13474	0.0049	0.0032	0.65	4994	0.0038	0.002	0.6	1193	0.0044	0.003	0.61
Ho2O3%	13474	0.0007	0.0004	0.64	4994	0.0006	0	0.63	1193	0.0007	0	0.67
Er2O3 %	13474	0.0015	0.0009	0.57	4994	0.0014	0.001	0.65	1193	0.0015	0.001	0.69
Tm2O3%	13474	0.0002	0.0001	0.54	4994	0.0002	0	0.66	1193	0.0002	0	0.7
Yb2O3 %	13474	0.001	0.0004	0.43	4994	0.0009	0.001	0.63	1193	0.0009	0.001	0.63
Lu2O3 %	13474	0.0001	0.0001	0.44	4994	0.0001	0	0.61	1193	0.0001	0	0.64
Y2O3 %	13474	0.0198	0.0125	0.63	4994	0.0178	0.012	0.67	1193	0.0191	0.013	0.69
LREO %	13474	0.3188	0.1737	0.54	4994	0.2556	0.224	0.88	1193	0.4725	0.347	0.73
MREO%	13474	0.0311	0.0158	0.51	4994	0.0175	0.009	0.54	1193	0.0248	0.013	0.51
HREO %	13474	0.0295	0.0182	0.62	4994	0.0257	0.017	0.65	1193	0.0278	0.018	0.66
TREO %	13474	0.3795	0.1824	0.48	4994	0.299	0.234	0.78	1193	0.5252	0.364	0.69
S.G.	1225	2.99	0.2315	0.08	335	2.8	0.238	0.08	69	2.9	0.152	0.05

11.5  Exploratory Data Analysis

Exploratory data analysis was conducted on the capped distributions from each domain. Correlation matrices were generated using homotopic observations to view relationships, excluding density, for each domain (Figure 11-9). Density was excluded from the dataset due to extreme inequality of density sampling relative to the oxide/scandium dataset. The correlations observed in Figure 11-9 are to be reasonably replicated in the block model estimate.

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Figure 11-9: Correlation matrices of the homotopic observations of capped oxides and scandium distributions within the MCarb (Bound 1), SW (Bound 2), and NE (Bound 3) domains.

The MCarb domain is of particular interest, as it hosts 71% of the reported Indicated Mineral Resources and 97% of the reported Mineral Reserves. Therefore, bivariate plots were generated to better understand the relationship between the variables defining the NSR calculation (niobium, titanium, scandium, and density) in the MCarb domain (Figure 11-10).

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Figure 11-10: Bivariate plots of the homotopic composite observations of niobium, titanium, scandium, and density distributions within the MCarb (Bound 1) domain.

Niobium and titanium are strongly, positively, and nearly linearly correlated with one another, and scandium is weakly and non-linearly correlated with niobium – these relationships are important to replicate upon estimation for accurate reporting. The density variable is weakly and positively correlated with niobium and titanium within the MCarb domain. The controls on density were further investigated by reviewing the raw measured density measurements relative to logged lithological units through box plots and histogram analysis (Figure 11-11). The mean of the logged MCarb density distribution is slightly higher than the other logged lithologies and the distribution’s upper tail extends appreciably beyond the other distributions. Understood concluded that logged lithology is the primary predictor in density measurements’ value with niobium and titanium concentrations being a lesser predictor. The modelled MCarb domain preserves the observation of elevated density measurements of the MCarb unit relative to the neighbouring lithology domains (Figure 11-12).

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Figure 11-11: Box plot and histogram of the specific gravity grouped by logged lithology

Figure 11-12: Box plot and histogram of the specific gravity grouped by estimation domains

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11.6  Variography

Independent variography was performed in Vulcan on each variable of the capped composite datasets for each domain (MCarb, SW, and NE domains). Directional experimental variograms were created, but were determined to be unstable, therefore omni-directional variograms were used. The omni-directional variograms reflect the nature of the mineralization of the deposit, where mineralization appears to be isotopic in all directions within each domain. The lack of directional variability is evident in the niobium fan variogram in Figure 11-13. A variogram model was fitted to the experimental variograms with the nugget contribution being applied to each model as observed in the downhole variogram. In total, 69 experimental and model variograms were created. Figure 11-14 is an example of experimental and model variograms for scandium within the MCarb domain Table 11-6 summarizes the variogram models for niobium, titanium, scandium, TREO, and density within the MCarb domain.

Figure 11-13: Fan variogram of niobium

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Figure 11-14: Omni-direction experimental and model variogram for scandium in the MCarb domain

Table 11-6: Variogram models for select variables within the MCarb domain

Variable	Type	Sill	Major	Semi	Minor
Nb2O5%	Nugget	0.20	-	-	-
	Spherical	0.51	20.16	20.16	20.16
	Spherical	0.29	74.85	74.85	74.85
TiO2 %	Nugget	0.20	-	-	-
	Spherical	0.39	18.289	18.289	18.289
	Spherical	0.41	40.64	40.64	40.64
Sc ppm	Nugget	0.15	-	-	-
	Spherical	0.33	19.64	19.64	19.64
	Spherical	0.09	71.13	71.13	71.13
	Spherical	0.43	173.41	173.41	173.41
TREO %	Nugget	0.30	-	-	-
	Spherical	0.37	9.235	9.235	9.235
	Spherical	0.07	88.54	88.54	88.54
	Spherical	0.26	600	600	600
S.G.	Nugget	0.40	-	-	-
	Spherical	0.21	28.13	28.13	28.13
	Spherical	0.39	101.88	101.88	101.88

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11.7  Block Model Resource Estimation

11.7.1  Estimation Overview

A three-dimensional block model was constructed in Vulcan to encompass the MCarb, SW, and NE domains. The following fields were estimated within the block model: Density, Nb₂O₅ (%), TiO₂(%), Sc (ppm), La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), Nd₂O₃ (%), Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), Y₂O₃ (%), LREO (%), MREO (%), HREO (%), and TREO (%). Understood believes there is sufficient specific gravity data for independent estimation, and that estimation of density will be more accurate and precise than assignment of density based on global observations. LREO, MREO, HREO, and TREO were estimated as a check on the individual REO estimates.

The blocks variables were independently interpolated through the use of OK as informed by omnidirectional variogram models. The estimate was completed in a single run for each domain using an isotropic search of 200 metres. A minimum of 4 to a maximum of 50 composites were used per estimate with no direct restrictions on the number of holes per estimate. A high yield limit was placed on TiO₂ estimates in the NE domain, effectively reducing the search ellipse dimensions to 50 m by 50 m by 50 m for samples over 3.0 % TiO₂.

11.7.2  Block Model Definition

The minimum extents of the block model are 738775.0 in the X direction, 4461000.0 in the Y direction, and -650.0 in the Z direction. A block size of 5 m by 5 m by 5 m was selected for the block model. The size of the block approximates the size of an underground drift round (the smallest mining unit for consideration), and adequately captures the geologic features of the modelled domains. The block model is not rotated or tilted and is made up of 185 columns (X direction), 130 rows (Y direction), and 170 levels (Z direction) for a total of 4,088,500 blocks. A whole block approach was used whereby the block was assigned a numerical code based on the domain where its centroid is located. Blocks within the MCarb were coded as 1, SW as 2, and NE as 3. The models fully enclose the modelled resource wireframes and only the blocks within the domain are estimated. The variables of the block model are listed in Table 11-7. The final block model is named ELK_5x5x5_ok_2022Q1_rev2-5.bmf.

Table 11-7: Block model variables

Variable	Description	Variable	Description
nsamp	Number of samples per estimate	ore	Domains (MCarb =1, SW =2, NE =3)
nholes	Number of holes per estimate	class	Classification (3 = Inf, 2 =Ind)
est_avg_dist	Average distance to samples per est.	nb2o5_pct	Est Variable Nb2O5 %
est_samp_dist	Distance to nearest sample per est.	tio2_pct	Est Variable TiO2 %
nn	Nearest neighbour grade	sc_ppm	Est Variable Sc ppm
nn_distance	Distance to nearest neighbour	la2o3_calc_pct	Est Variable La2O3 %
est_flag_nb	Estimation flag for Nb2O5 ID	ce2o3_calc_pct	Est Variable Ce2O3 %
est_flag_ti	Estimation flag for TiO2 ID	pr2o3_calc_pct	Est Variable Pr2O3 %
est_flag_sc	Estimation flag for Sc ID	nd2o3_calc_pct	Est Variable Nd2O3 %
est_flag_den	Estimation flag for den ID	sm2o3_calc_pct	Est Variable Sm2O3 %

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est_flag_reo	Estimation flag for REO ID	eu2o3_calc_pct	Est Variable Eu2O3 %
krig_var	Kriging variance variable	gd2o3_calc_pct	Est Variable Gd2O3 %
blk_var	Block variance variable	tb2o3_calc_pct	Est Variable Tb2O3 %
krig_eff	Kriging efficiency variable	dy2o3_calc_pct	Est Variable Dy2O3 %
nb2o5_dil	Nb Oxide % - Diluted	ho2o3_calc_pct	Est Variable Ho2O3 %
tio2_dil	Ti Oxide % - Diluted	er2o3_calc_pct	Est Variable Er2O3 %
sc_ppm_dil	Elemental Scandium ppm - Diluted	tm2o3_calc_pct	Est Variable Tm2O3 %
rev_nb	Revenue Nb2O5	yb2o3_calc_pct	Est Variable Yb2O3 %
rev_ti	Revenue TiO2	lu2o3_calc_pct	Est Variable Lu2O3 %
rev_sc	Revenue Sc	y2o3_calc_pct	Est Variable Y2O3 %
rev_nb_dil	Revenue Nb2O5 - Diluted	lreo_calc_pct	Est Variable LREO %
rev_ti_dil	Revenue TiO2 - Diluted	mreo_calc_pct	Est Variable MREO %
rev_sc_dil	Revenue Sc - Diluted	hreo_calc_pct	Est Variable HREO %
nsr	Net Smelter Return per tonne	treo_calc_pct	Est Variable TREO %
nsr_dil	NSR_diluted	den	Est Variable S.G.
tonnes_dil	Tonnes_diluted

11.7.3 Estimation Strategy and Testing

It is Understood’s opinion that a Mineral Resource Estimate should honour the data, replicate the relationships therein, be globally unbiased, be geologically reasonable, and meet the applicable reporting standards. An estimation strategy was constructed, and multiple models were tested to meet the stated criteria.

11.7.3.1  Estimation Strategy

As previously mentioned, the mineralization appears to be isotropic within each domain, hence the fitting of the omni-directional variograms and the necessity for isotropic search ellipses.

Minor localizations of high-grade and low-grade concentrations of variables are observed within the domains. OK assumes an unknown mean (and consequently a local varying mean during estimation) and was therefore selected as the kriging method to manage the local trends within the domains. Co-kriging was also considered but given the number of variables and that the data is nearly homotopic observed (less than 1% of the locations missing data), it was decided that OK would be simpler to implement and still honour the relationships in the domains.

The estimate is to closely replicate the declustered mean of each variable within the domains to ensure the estimate is globally unbiased. Target variance of key variables (niobium, titanium, scandium, and TREO) within the MCarb domain were derived from the composite data using a Discrete Gaussian Model (DGM). The DGM accounts for change of support using a variogram model, a normal score transformation, and Hermite polynomials (Harding & Deutsch, 2019). Change of support means that as the support of the core sample increases to the size of a mining unit (or block size) the observed variability will decrease and the distribution will become more symmetric (Harding & Deutsch, 2019; Figure 11-15). The details of the DGM will be further described in Section 11.8 Model Validation.

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Figure 11-15: Diagram demonstrating the change of support principle (Harding & Deutsch, 2019)

11.7.3.2  Testing and Strategy Refinement

The first series of OK models implemented search ellipses according to the ranges of the variograms and tested different composites per estimate, targeting the declustered mean and variance from the DGM. An example of the targeting and composite testing is graphically displayed in Figure 11-16.

Figure 11-16: OK model of TiO₂ (MCarb domain) using different number of composites per estimate compared against the target mean and variance.

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The models replicate the data and were globally unbiased, but regardless of the number of composites included per estimate, were visibly disjunctive (similar to a Nearest Neighbour model) and geologically unlikely. The model requires smoothing to be geologically reasonable, which was completed by expanding the search ellipses.

Honouring the scandium variance is important for accurate and precise reporting of resources, as it contributes, on average, 74% of the revenue towards the NSR calculation (Figure 11-17). The scandium variogram model is stable with a relatively large range. The range of the scandium variogram informed the search ellipse size for all variables, effectively smoothing the model for variables with shorter ranges yet maintaining the variance observed in scandium. Figure 11-18 is an example and comparison of the smoothed and unsmoothed models.

Figure 11-17: Contribution percentage of scandium revenue to NSR (Diluted) from block model.

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Figure 11-18: Cross section looking northwest of the unsmoothed and smoothed OK TiO₂ models within the MCarb Domain.

Another control on smoothing is constraining the number of drill holes used per estimate. Numerous tests were completed with varying number of samples and drill holes used per estimate. Constraining the estimate by number of drill holes resulted in a “striped” model (Figure 11-19) that over-smooths (i.e., averages out the high and low values). Understood considers the striped, overly smoothed model to be geologically unreasonable and statistically imprecise. In the final model, no drill hole constraints were used during estimation.

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Figure 11-19: Cross section looking northwest of the DDH constrained and unconstrained OK TiO₂ models within the MCarb Domain.

The expanded, uniform search ellipse strategy increased the likelihood that estimated blocks will be populated with all the variables. The result is a model that meets the previously stated criteria for a Mineral Resource Estimate with focus on the principal driver to the economics of the deposit – further demonstrations are available in Section 11.8 Model Validation.

Nearest Neighbour (NN) estimates and Inverse Distance Squared (ID2) estimates were also created to compared against the OK models for global bias assessment.

11.7.4  Estimation/Interpolation Methods

The blocks variables were independently estimated in Vulcan using OK and the omnidirectional variogram models. The interpolation was completed in a single pass for each variable in each domain using an isotropic search of 200 m by 200 m by 200 m. The search ellipses are isotropic; therefore, no orientation was required. Hard boundaries were used to limit the use of composites between domains. A minimum of 4 to a maximum of 50 composites were used per estimate with no restrictions on the number of holes per estimate. The estimate only selected composites with values between -1 and 999 to avoid selecting “missing data” during estimation.

The following fields were estimated within the block model: Density, Nb₂O₅ (%), TiO₂(%), Sc (ppm), La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), Nd₂O₃ (%), Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%),

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Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), Y₂O₃ (%), LREO (%), MREO (%), HREO (%), and TREO (%). If a block was not estimated for density, it was assigned a default of 2.90. If other variables were not estimated for a block, the variable was set to a default of 0.

Initially, TiO2 estimates in the NE domain were overestimating relative to the declustered representative composite distribution, representing a global bias. A high yield limit was placed on TiO2 estimates in the NE domain, effectively reducing the search ellipse dimensions to 50 m by 50 m by 50 m for samples over 3.0 % TiO2. The TiO2 estimate in the NE domain is the only estimate that required a high-yield limit.

11.8  Model Validation

Understood validated the block model by mean comparison, volumetric comparison, visual inspection, swath plots, histogram comparison, bivariate plot comparisons, and correlation checks. Overall, there is a good correlation between the block estimates and the supporting composite grades. The MCarb domain was investigated more than the other domains, as it hosts 62% of the reported Indicated Mineral Resources and 97% of the reported Mineral Reserves.

11.8.1  Rare Earth Considerations

LREO, MREO, HREO, and TREO were estimated as a check on the individual REOs. The estimates of the individual REOs that constitute the LREO, MREO, HREO, and TREO variables were summed and validated against the estimated LREO, MREO, HREO, and TREO values (Figure 11-20). The comparison demonstrates a high degree of correlation and minimal variance between the summed and estimated values. Therefore, it is permissible to use the LREO, MREO, HREO, and TREO as global checks in place of checking each individual REO.

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Figure 11-20: The summed estimates of the individual REOs that constitute the LREO, MREO, HREO, and TREO variables versus the estimated LREO, MREO, HREO, and TREO values.

11.8.2  Global Checks

The average block grades were compared to the means of the representative distributions for an assessment of global bias (Table 11-8). The means of the variables within the MCarb domain were reproduced very well, with a max difference of 3%. The means of the variables within the SW and NE domains, excluding the variables associated with LREO in the SW domain, are reasonably replicated as well. The underestimation of LREO in the SW domain is attributed to the constraining of a high-grade blocks along the boundary of the domain (Figure 11-21). The declustering weights of the representative distribution do not consider the boundary constraints.

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Table 11-8: Comparison of block grades to representative distribution

Variable	Composite Avg (Declustered)			Block avg (Reported Material)			% Difference
	MCarb	SW	NE	MCarb	SW	NE	MCarb	SW	NE
Nb2O5 %	0.58	0.21	0.23	0.58	0.21	0.24	0%	-1%	6%
TiO2 %	2.48	1.15	1.55	2.47	1.10	1.48	-1%	-5%	-4%
Sc ppm	64.7	33.7	34.1	64.8	32.5	35.9	0%	-3%	5%
La2O3 %	0.0908	0.0701	0.1317	0.0919	0.0594	0.1280	1%	-15%	-3%
Ce2O3 %	0.1538	0.1238	0.2287	0.1556	0.1050	0.2238	1%	-15%	-2%
Pr2O3 %	0.0161	0.0133	0.0252	0.0163	0.0114	0.0244	1%	-15%	-3%
Nd2O3 %	0.0581	0.0485	0.0867	0.0588	0.0419	0.0845	1%	-14%	-2%
Sm2O3 %	0.0141	0.0086	0.0131	0.0143	0.0078	0.0130	1%	-10%	-1%
Eu2O3 %	0.0050	0.0025	0.0036	0.0051	0.0023	0.0036	1%	-8%	1%
Gd2O3 %	0.0120	0.0064	0.0081	0.0120	0.0060	0.0083	0%	-6%	2%
Tb2O3 %	0.0013	0.0008	0.0010	0.0012	0.0008	0.0010	-2%	-4%	0%
Dy2O3 %	0.0049	0.0038	0.0044	0.0048	0.0037	0.0043	-3%	-3%	-3%
Ho2O3 %	0.0007	0.0006	0.0007	0.0007	0.0006	0.0006	-2%	-2%	-5%
Er2O3 %	0.0015	0.0014	0.0015	0.0015	0.0014	0.0014	-2%	-2%	-5%
Tm2O3 %	0.0002	0.0002	0.0002	0.0002	0.0002	0.0002	-2%	-2%	-5%
Yb2O3 %	0.0010	0.0009	0.0009	0.0010	0.0009	0.0009	-1%	-2%	-3%
Lu2O3 %	0.0001	0.0001	0.0001	0.0001	0.0001	0.0001	0%	-3%	-3%
Y2O3 %	0.0198	0.0178	0.0191	0.0194	0.0174	0.0183	-2%	-2%	-4%
LREO %	0.3188	0.2556	0.4725	0.3228	0.2172	0.4608	1%	-15%	-2%
MREO %	0.0311	0.0175	0.0248	0.0314	0.0161	0.0248	1%	-8%	0%
HREO %	0.0295	0.0257	0.0278	0.0288	0.0251	0.0268	-2%	-2%	-4%
TREO %	0.3795	0.2990	0.5252	0.3830	0.2585	0.5123	1%	-14%	-2%
S.G.	2.99	2.80	2.90	3.00	2.83	2.89	0%	1%	0%

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Figure 11-21: Cross section looking northwest of LREO composites and blocks within the SW domain showing the constrained high-grade blocks

Wireframe volumes were compared to block volumes for each domain, as summarized in Table 11-9. Results show that there is good agreement between the wireframe and block model volumes, with a maximum difference of |0.17|%.

Table 11-9: Block to wireframe volume comparison

Domain	Block Volume	Wireframe Volume	% Difference
MCarb	73,282,375.00	73,286,491.37	0.01%
SW	62,055,250.00	62,031,175.26	-0.04%
NE	29,315,000.00	29,364,084.70	0.17%

The DGM created for the niobium, titanium, scandium, and TREO in the MCarb domain produced a target distribution that accounts for the change of support (Figure 11-22). The blocks were compared to the DGM target distributions reasonably well, but some minor over-smoothing is observed. The over-smoothing was intentional to produce a geologically reasonable model. Notably, scandium, the economic driver to the reporting of NSR, replicates the target distribution reasonably well.

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Figure 11-22: Histogram comparison of blocks relative to DGM target distributions

11.8.3  Visual Inspection

Block grades were visually compared with drill hole composites on cross-sections, longitudinal sections, and plan views. The block grades and composite grades correlate very well visually within the Elk Creek Deposit. Figure 11-23 contains long sections of niobium, titanium, scandium, and TREO comparing the informing composites to the estimated blocks. Blocks shown are restricted to Indicated and Inferred Mineral Resources.

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Figure 11-23: Plan view of the Elk Creek domains and long sections looking southwest of the niobium, titanium, scandium, and TREO block model grades with informing composite grades. Blocks shown are restricted to Indicated and Inferred Mineral Resource material.

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11.8.4  Swath Plots

A series of swath plots were generated for niobium, titanium, scandium, TREO, and density from slices throughout the MCarb domain (Figure 11-24). The swath plots compare the block model grades against the composite grades to evaluate any potential local grade bias; no bias was identified in the model. As expected, the composite database is more variable than the block model, but the block model captures general trends observed in the data.

Figure 11-24: Niobium, titanium, scandium, TREO, and density swath plots in the MCarb domain

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11.8.5  Correlation Review

The bivariate plots generated of composited primary variables in the MCarb domain (Figure 11-25) were compared to the bivariate plots of the block model (Figure 11-10). The correlation coefficients and bivariate distributions of the reviewed variables closely reproduce the composite data.

Figure 11-25: Bivariate plots of block model estimate of niobium, titanium, scandium, and density distributions within the MCarb domain

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11.9  Mineral Resource Classification

The 2022 Elk Creek Mineral Resource Estimate contains Indicated and Inferred Mineral Resources (Figure 11-26). The classification was assigned to regions of the block model based on the Qualified Person’s confidence and professional judgement related to the geological understanding and continuity of mineralization in conjunction with data quality, spatial continuity, block model representativeness, and data density. On average, the Indicated Mineral Resources are informed with a drill hole spacing between 50-75 m and extend approximately 35-50 m laterally beyond the last drill intercept. Additionally, portions of the block model that had noticeable “striping” were avoided in the declaration of Indicated Mineral Resources. The Inferred Mineral Resources capture the sparser drilled areas with an average drill hole spacing of 75 to 125 m and extend approximately 50-75 m laterally beyond the last drill intercept (Figure 11-27). Exceptions to the stated classification guidelines occurs but are rare and typically err conservatively.

Figure 11-26: Long-section looking southwest of the 2022 Inferred and Indicated Domains underlain with niobium composites

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Figure 11-27: Plan section (elevation of 25) of the 2022 Inferred and Indicated domains with niobium composites displayed as spheres

11.10  Reasonable Prospects of Eventual Economic Extraction

To fulfill the requirement to meet “reasonable prospects for eventual economic extraction”, Understood estimated a potential underground mining cut-off grade using assumptions from the previous technical studies, which were based on known operating costs for UG mines operating in the region. The project is amendable to the processing method for FeNb, TiO₂, and Sc2O₃. Details regarding the estimation of the cut-off grade are presented in Section 11.11, below. It is Understood’s opinion that all relevant technical and economic factors likely to influence the prospect of economic extraction can be resolved with further work.

11.11  Cut-Off Grade

The 2022 Mineral Resource Estimate is reported at a diluted NSR of US$ 180 per tonne based on NioCorp’s estimated break-even OPEX mining cost of US$ 180 per tonne, as per Table 11-10.

Table 11-10: Mining cost assumptions

Cost Item	Value	Unit
Mining Cost	$    50.00	US$/t mined
Processing	$  125.00	US$/t mined
General and Administrative	$      5.00	US$/t mined
Total Cost	$  180.00	US$/t mined

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The calculation of diluted NSR from Nb₂O₅, TiO2, and Sc is as follows:

The diluted tonnes are a 6% increase in the total tonnes of the block. The diluted revenue from Nb₂O₅, TiO₂, and Sc per block used the following factors:

● Nb₂O₅ Revenue: a 94% grade recovery, a 0.696 factor to convert Nb₂O₅ to Nb, 82.36% assumption for plant recovery, and a US$ 39.60 selling price per kg of ferroniobium.

● TiO₂ Revenue: a 94% grade recovery, a 40.31% assumption for plant recovery, and a US$ 0.88 selling price per kg of titanium oxide.

● Sc Revenue: a 94% grade recovery, a 1.534 factor to convert Sc to Sc₂O₃, 93.14% assumption for plant recovery, and a US$ 3,675 selling price per kg of scandium oxide.

The pricing for Niobium, Scandium and Titanium used to support the Mineral Resource estimate was provided by NioCorp. NioCorp has relied on third party market information, relying on Roskill (2019) for niobium, OnG Commodities (2019) for Scandium and USGS (2019) for titanium. It is Dahrouge’s opinion that the pricing used is adequate and appropriate for use in estimating mineral resources. The pricing for the prospective products from the Project is effective as of June 30, 2022. Further details on the market assumptions and timeframes analyzed are disclosed in Chapter 16 of this report.

11.12  Mineral Resource Tabulation

The 2022 Mineral Resource Estimate for the Elk Creek Deposit adheres to the S-K 1300 classification system and was reported at a US$ 180 diluted NSR breakeven cut-off grade. The REOs were evaluated as a potential by-product to the mining of niobium, titanium, and scandium; thus, the reported REOs are coincident with above-cut-off diluted NSR values as derived from the Nb₂O₅, TiO₂, and Sc estimates. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the Mineral Resource will be converted into a Mineral Reserve.

Table 11-11: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) excluding reserves

Class	NSR Cutoff	Tonnage (Mt)
Indicated	180	151.7	Nb2O5 (%)	Nb2O5 (kt)
			0.43	649.8
			TiO2 (%)	TiO2 (kt)
			2.02	3,067
			Sc (ppm)	Sc (t)
			56.42	8,558
Inferred	180	108.3	Nb2O5 (%)	Nb2O5 (kt)

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0.39	426.6
TiO2 (%)	TiO2 (kt)
1.92	2,082

Table 11-12: Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) excluding reserves

Class	NSR Cut-off	Tonnage (Mt)
Indicated	180	151.7	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0766	116.2	0.1320	200.2	0.0140	21.3
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0511	77.5	0.0116	17.6	0.0040	6.0
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.0096	14.6	0.0011	1.6	0.0044	6.7
			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0006	1.0	0.0015	2.2	0.0002	0.3
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.0010	1.5	0.0001	0.2	0.0187	28.4
			LREO (%)	LREO (kt)	HREO(%)	HREO (kt)	TREO (%)	TREO (kt)
			0.2737	415.2	0.0528	80.0	0.3265	495.2
Inferred	180	108.3	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0943	102.1	0.1576	170.6	0.0163	17.7
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0575	62.2	0.0116	12.6	0.0038	4.1
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.0090	9.8	0.0010	1.1	0.0042	4.6
			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0006	0.7	0.0014	1.5	0.0002	0.2
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.0010	1.1	0.0001	0.1	0.0182	19.7
			LREO (%)	LREO (kt)	HREO(%)	HREO (kt)	TREO (%)	TREO (kt)
			0.3257	352.6	0.0512	55.5	0.3769	408.1

Notes:

a. Classification of Mineral Resources in the above tables is in accordance with the S-K 1300 classification system. Mineral Resources in this table are reported exclusive of Mineral Reserves

b. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

c. The Mineral Resources are reported at a Diluted Net Smelter Return (NSR) Cut-off of US $180/tonne.

d. The diluted NSR is defined as:

●

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● The diluted revenue from Nb₂O₅, TiO₂, and Sc per block used the following factors:

● Nb₂O₅ Revenue: a 94% grade recovery, a 0.696 factor to convert Nb₂O₅ to Nb, 82.36% assumption for plant recovery, and a US$ 39.60 selling price per kg of ferroniobium as of June 30, 2022.

● TiO₂ Revenue: a 94% grade recovery, a 40.31% assumption for plant recovery, and a US$ 0.88 selling price per kg of titanium oxide as of June 30, 2022.

● Sc Revenue: a 94% grade recovery, a 1.534 factor to convert Sc to Sc₂O₃, 93.14% assumption for plant recovery, and a US$ 3,675 selling price per kg of scandium oxide as of June 30, 2022.

● The diluted tonnes are a 6% increase in the total tonnes of the block.

e. Price assumptions for FeNb, Sc₂O₃, and TiO₂ are based upon independent market analyses for each product.

f. Numbers may not sum due to rounding. The rounding is not considered to be material.

g. Rare Earth Oxides (REO) were evaluated as a potential by-product to the mining of niobium, titanium, and scandium; thus the estimated values of the REOs are reported using the previously determined diluted NSR as derived from the Nb₂O₅, TiO₂, and Sc Mineral Resources and are assigned a price of $0.

h. The stated Light Rare Earth Oxides (LREO) grade (%) is the summation of La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), and Nd₂O₃ (%) estimates.

i. The stated Heavy Rare Earth Oxides (HREO) grade (%) is the summation of Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), and Y₂O₃ (%) estimates.

j. The stated Total Rare Earth Oxide (TREO) grade (%) is the summation of LREO (%) and HREO (%).

k. The effective date of the Mineral Resource, including by-products, is June 30, 2022.

Table 11-13: Elk Creek 2022 In Situ Mineral Resource Estimate (niobium, titanium, and scandium) including reserve material

Class	NSR Cutoff	Tonnage (Mt)
Indicated	180	188.8	Nb2O5 (%)	Nb2O5 (kt)
			0.51	970.3
			TiO2 (%)	TiO2 (kt)
			2.24	4,221
			Sc (ppm)	Sc (t)
			60.06	11,337
Inferred	180	108.3	Nb2O5 (%)	Nb2O5 (kt)
			0.39	426.6
			TiO2 (%)	TiO2 (kt)
			1.92	2,082
			Sc (ppm)	Sc (t)
			52.28	5,660.20

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Table 11-14: Elk Creek 2022 In Situ Mineral Resource Estimate (rare earth oxides) including reserve material

Class	NSR Cut- off	Tonnage (Mt)
Indicated	180	188.8	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0773	145.8	0.1335	251.9	0.0143	26.9
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0524	98.9	0.0129	24.3	0.0046	8.6
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.0110	20.8	0.0012	2.3	0.0048	9.1
			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0007	1.3	0.0015	2.9	0.0002	0.3
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.0010	1.9	0.0001	0.3	0.0199	37.6
			LREO (%)	LREO (kt)	HREO(%)	HREO (kt)	TREO (%)	TREO (kt)
			0.2774	523.6	0.0579	109.3	0.3353	632.9
Inferred	180	108.3	La2O3 (%)	La2O3 (kt)	Ce2O3 (%)	Ce2O3 (kt)	Pr2O3 (%)	Pr2O3 (kt)
			0.0943	102.1	0.1576	170.6	0.0163	17.7
			Nd2O3 (%)	Nd2O3 (kt)	Sm2O3 (%)	Sm2O3 (kt)	Eu2O3 (%)	Eu2O3 (kt)
			0.0575	62.2	0.0116	12.6	0.0038	4.1
			Gd2O3 (%)	Gd2O3 (kt)	Tb2O3 (%)	Tb2O3 (kt)	Dy2O3 (%)	Dy2O3 (kt)
			0.009	9.8	0.0010	1.1	0.0042	4.6
			Ho2O3 (%)	Ho2O3 (kt)	Er2O3 (%)	Er2O3 (kt)	Tm2O3(%)	Tm2O3 (kt)
			0.0006	0.7	0.0014	1.5	0.0002	0.2
			Yb2O3 (%)	Yb2O3 (kt)	Lu2O3 (%)	Lu2O3 (kt)	Y2O3 (%)	Y2O3 (kt)
			0.001	1.1	0.0001	0.1	0.0182	19.7
			LREO (%)	LREO (kt)	HREO(%)	HREO (kt)	TREO (%)	TREO (kt)
			0.3257	352.6	0.0512	55.5	0.3769	408.1

Notes:

a. Classification of Mineral Resources in the above tables is in accordance with the S-K 1300 classification system. Mineral Resources in this table are reported inclusive of Mineral Reserves

b. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

c. The Mineral Resources are reported at a Diluted Net Smelter Return (NSR) Cut-off of US $180/tonne.

d. The diluted NSR is defined as:

●

● The diluted revenue from Nb₂O₅, TiO₂, and Sc per block used the following factors:

● Nb₂O₅ Revenue: a 94% grade recovery, a 0.696 factor to convert Nb₂O₅ to Nb, 82.36% assumption for plant recovery, and a US$ 39.60 selling price per kg of niobium as of June 30, 2022.

● TiO₂ Revenue: a 94% grade recovery, a 40.31% assumption for plant recovery, and a US$ 0.88 selling price per kg of titanium oxide as of June 30, 2022.

● Sc Revenue: a 94% grade recovery, a 1.534 factor to convert Sc to Sc₂O₃, 93.14% assumption for plant recovery, and a US$ 3,675 selling price per kg of scandium oxide as of June 30, 2022.

● The diluted tonnes are a 6% increase in the total tonnes of the block.

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e. Price assumptions for FeNb, Sc₂O₃, and TiO₂ are based upon independent market analyses for each product.

f. Numbers may not sum due to rounding. The rounding is not considered to be material.

g. Rare Earth Oxides (REO) were evaluated as a potential by-product to the mining of niobium, titanium, and scandium; thus the estimated values of the REOs are reported using the previously determined diluted NSR as derived from the Nb₂O₅, TiO₂, and Sc Mineral Resources and are assigned a price of $0.

h. The stated Light Rare Earth Oxides (LREO) grade (%) is the summation of La₂O₃ (%), Ce₂O₃ (%), Pr₂O₃ (%), and Nd₂O₃ (%) estimates.

i. The stated Heavy Rare Earth Oxides (HREO) grade (%) is the summation of Sm₂O₃ (%), Eu₂O₃ (%), Gd₂O₃ (%), Tb₂O₃ (%), Dy₂O₃ (%), Ho₂O₃ (%), Er₂O₃ (%), Tm₂O₃ (%), Yb₂O₃ (%), Lu₂O₃ (%), and Y₂O₃ (%) estimates.

j. The stated Total Rare Earth Oxide (TREO) grade (%) is the summation of LREO (%) and HREO (%).

k. The effective date of the Mineral Resource, including by-products, is June 30, 2022.

11.13 Mineral Resource Uncertainty

Mineral deposits, including the Elk Creek Deposit, are inherently uncertain because of variability at all scales and sparse sampling. In addition to uncertainty associated with estimation, there are specific risks and sources of uncertainty associated with the Elk Creek Deposit. These risks should be evaluated by potential and current investors.

S-K 1300 and other similarly purposed International Codes (JORC, 2012; NI 43-101, 2014) are to disclose risks to the public as identified and evaluated by the Qualified Person. The Qualified Person addresses the technical risks in various sections and considers that no material technical risks are identified.

The risks listed below are not considered exhaustive and there may be additional risks and uncertainties not presently known, such as market or technology changes, that are currently deemed immaterial but may also affect the business.

11.13.1  Specific Identified Risks

Scandium on average contributes 74% of the revenue towards the NSR Calculation, which is partly attributed to the price assumptions used for scandium. According to the US Geological Survey, the global scandium market is small relative to most other metals and the metal is typically traded between private parties, thus prices can fluctuate greatly, which would affect the NSR calculation.

Niobium data is exhaustively sampled at all data locations and less than 1% of the locations missing data for scandium, titanium, or REO. However, there are significantly less density measurements, with only ~10% of the data locations having density measurements. The lack of data increases uncertainty in the estimation of density across the deposit, equating to increased uncertainty in estimated tonnage.

Further advances in geostatistical estimation may be expected including more use of directional anisotropy (through directional variograms rather than the use of omni-directional variograms), the use of co-kriging for consideration of secondary data for estimation (rather than independent estimation of each variable), and conditional simulation to quantify estimation risk and optimize drill sampling grids.

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The drill sampling methods used at the Elk Creek Deposit meet or exceed industry standards and the assay results have been comprehensively reviewed and validated. The geostatistical estimates of in situ tonnages and grades are reasonable and validated by comprehensive model checking. The Understood Qualified Person considers that these methods are appropriate to produce the declared Mineral Resources.

11.13.2  Generic Mineral Resource Uncertainty

Mineral resources are uncertain because of variability at all scales and sparse sampling. The variables constituting the mineral resource, the volume of the geological interpretation and the grade estimates within that volume, are the sources of uncertainty. These uncertainties are typically a function of drill spacing, with denser spacing equating to less uncertainty and sparser spaced areas having more uncertainty. Understood classified the Mineral Resource into Indicated and Inferred Resources categories based on geological and grade continuity as well as drill hole spacing; therefore, adhering to the well-studied concept that drilling reduces uncertainty.

Changes to the geologic interpretation would greatly alter the estimation. If new interpretations of geological complexities are presented, the Mineral Resource would need to be updated to reflect the new interpretations.

NioCorp cannot be certain that any part or parts of a deposit or Mineral Resource estimate will ever be confirmed or converted into Mineral Reserves or that mineralization can in the future be economically or legally extracted.

11.14 Mineral Resource Sensitivity

The resources were calculated at various diluted NSR cut-off thresholds as a review of the deposit’s sensitivity to change in mining costs. Niobium, titanium, scandium, and TREO are summarized in Table 11-15 and the individual REOs are visualized in Figure 11-28 and Figure 11-29.

Table 11-15: Grade/tonnage by diluted NSR cut-off

	NSR Cutoff	Tonnage (Mt)	Nb2O5 (%)	Nb2O5 (kt)	TiO2 (%)	Ti O2 (kt)	Sc (ppm)	Sc (t)	TREO (%)	TREO (kt)
Indicated	0	213.3	0.48	1,019.4	2.08	4,435	55.7	11,884.7	0.3356	715.9
	100	213.3	0.48	1,019.4	2.08	4,435	55.7	11,884.5	0.3356	715.8
	180	188.8	0.51	970.3	2.24	4,221	60.1	11,336.9	0.3353	632.9
	200	176.3	0.53	942.2	2.33	4,101	62.3	10,986.8	0.3424	603.5
	300	137.3	0.62	847.3	2.55	3,506	69.0	9,470.8	0.3585	492.2
	400	102.5	0.69	703.4	2.65	2,719	74.5	7,634.0	0.3589	367.8
	500	55.2	0.79	437.1	2.86	1,578	82.6	4,560.9	0.3703	204.6
	600	19.8	0.92	183.0	3.16	627	92.4	1,833.2	0.4030	80.0
	700	3.9	1.03	40.1	3.37	132	106.4	415.4	0.4472	17.5
	800	0.6	1.16	7.2	3.24	20	126.6	78.9	0.5694	3.5
	900	0.2	1.30	2.8	3.41	7	134.9	28.5	0.6353	1.3

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Inferred	0	156.6	0.33	510.3	1.61	2,518	42.9	6,715.0	0.3829	599.6
	100	156.6	0.33	510.3	1.61	2,518	42.9	6,714.9	0.3829	599.6
	180	108.3	0.39	426.6	1.92	2,082	52.3	5,660.2	0.3769	408.1
	200	97.5	0.41	404.2	2.02	1,968	54.9	5,351.2	0.3820	372.5
	300	63.3	0.51	324.3	2.26	1,434	63.2	4,004.3	0.3954	250.4
	400	33.5	0.63	210.5	2.47	829	71.6	2,400.2	0.3905	130.9
	500	13.7	0.72	98.4	2.53	347	84.4	1,158.8	0.3650	50.1
	600	4.7	0.81	38.6	2.72	129	93.5	443.6	0.3787	18.0
	700	0.6	0.59	3.5	2.03	12	119.6	72.2	0.3669	2.2
	800	0.004	0.93	0.04	3.10	0.1	125.4	0.5	0.4576	0.02
	900	-	-	-	-	-	-	-	-	-

Figure 11-28: Grade-tonnage curve of individual REOs for Indicated Mineral Resources

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Figure 11-29: Grade-tonnage curve of individual REOs for Inferred Mineral Resources.

11.15 Relevant Factors

Understood is not aware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the Mineral Resource Estimate that is not discussed in this TRS.

A variety of factors may affect the 2022 Elk Creek Mineral Resource Estimate, including but not limited to: changes to product pricing assumptions, re-interpretation of geology, geometry and continuity of mineralization zones, mining and metallurgical recovery assumptions, and additional infill or step out drilling.

In Understood’s opinion, the estimation methodology is consistent with standard industry practice and the Indicated and Inferred Mineral Resource Estimates for Elk Creek are considered to be reasonable and acceptable.

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12. MINERAL RESERVE ESTIMATES

12.1 Conversion Assumptions, Parameters and Methods

Indicated Mineral Resources were converted to Probable Mineral Reserves by applying the appropriate modifying factors, as described within this sub-section, to potential mining block shapes created during the mine design process. No Measured Resources are estimated and, as a result, no Proven Reserves are stated.

The undiluted tonnes and grade of each potential mining block are based on the resource block model estimated by Understood Mineral Resources Ltd. as described in Section 14 of this report. All Mineral Reserve tonnages are expressed as “dry” tonnes (i.e., no moisture) and are based on the density values stored in the block model.

12.1.1  Dilution

Mining dilution of approximately 6% was applied to all stopes and development, based on 3% for the primary stopes, 9% for the secondary stopes, and 5% for ore development. The mining dilution was added to the designed tonnage to account for unplanned sources of dilution such as backfill and host rock around the periphery of the ore mass. Mining dilution of host rock from around the periphery of the ore mass has been applied with zero grade as a conservative assumption even though some sources of this type of dilution carry grade. The primary stopes will have ore, host rock and backfill as material that will slough into them while being extracted. The ore portion of the sloughed material is not included in the 3% dilution factor, as this ore is accounted for in the adjacent stopes. Secondary stopes have more sources of material with no grade and less ore from adjacent stopes; therefore, a higher dilution factor of 9% has been applied to them.

The thickness of external dilution is estimated as equivalent linear overbreak/slough (ELOS), for moderately weathered carbonatite, and for fresh to slightly weathered carbonatite. Sidewall and back dilution are not expected to be a problem because in the primary stopes dilution (from adjacent secondary stopes) will be at grade, and dilution from secondary stopes is managed by controlling backfill strength.

As shown in Figure 12-1, sources of mining dilution for primary stopes include:

● Backfill material on the floor/sill with no grade.

● Backfill material from the hangingwall end with no grade if the stope is adjacent to a previously mined stope.

● Low grade periphery rock dilution in the hangingwall or footwall if the stope is not adjacent to other stopes.

As shown in Figure 12-1, sources of mining dilution for secondary stopes include:

● Backfill material on the floor/sill with no grade.

● Backfill material from the hangingwall end with no grade if the stope is adjacent to a previously mined stope.

● Low grade periphery rock dilution in the hangingwall or footwall if the stope is not adjacent to other stopes.

● For most situations, backfill material on both sidewalls with no grade.

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Source: Nordmin, 2019

Figure 12-1: Sources of Mining Dilution for Typical Stope Layout (Not to Scale)

Table 12-1 shows the sources of mining dilution by stope type (primary and secondary) for typical stope geometry.

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Table 12-1: Sources of Mining Dilution for Typical Geometry by Stope Type

	Primary Stopes			Secondary Stopes
	P0	P1	P2	S0	S1	S2
Hanging Wall Dilution - Rock	Yes	No	No	Yes	No	No
Footwall Dilution - Rock	No	No	Yes	No	No	Yes
Hanging Wall Dilution - Backfill	No	Yes	Yes	No	Yes	Yes
Footwall Dilution - Backfill	No	No	No	No	No	No
Sidewalls - Rock (Ore)	No	No	No	No	No	No
Sidewalls - Backfill	No	No	No	Yes	Yes	Yes
Floor/Sill Dilution - Backfill	Yes	Yes	Yes	Yes	Yes	Yes

Source: Nordmin, 2019

12.1.2  Recovery

A stope recovery factor of 95% was used. The following items were used to calculate this factor:

● Material loss into backfill (floor) or 0.4 m.

● Material loss to side and end walls (under blast) of 0.2 m.

● Material loss to mucking along sides and in blind corners.

● Additional loss factor due to rockfalls, unanticipated regional stress loads, and other geotechnical reasons.

A development recovery factor of 95% was used for all horizontal development.

A recovery of 62.5% in sill pillar stopes was used.

12.1.3  Cut-Off Grade Calculation

Net Smelter Return (NSR) is a commonly accepted method of evaluating a mineral deposit where revenue is generated from multiple elements. NSR is defined as the proceeds from the sale of mineral products after deducting off-site processing and distribution costs. NSR is typically expressed on a dollar per tonne basis. For this Project, the NSR calculation considers revenue for three products, FeNb, TiO₂, and Sc₂O₃. A factor of 0.696 was used to convert Nb₂O₅ in the block model to Nb contained in the FeNb product. Similarly, a factor of 1.534 (1/0.652) was used to convert Sc to Sc₂O₃.

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Recoveries used are based on metallurgical test work discussed in Section 10. The NSR was evaluated for each block in the 3D geologic resource block model. Table 12-2 shows NSR parameters and an example NSR calculation for an individual block.

Table 12-2: Example of an NSR Block Calculation

Input Parameters	Total	Nb2O5	TiO2	Sc (1)
Example Block Model Mass	100 t
Example Block Model Grades		0.70%	2.50%	60 ppm
Metallurgical Recoveries (2)		82.36%	40.31%	93.14%
Amount Payable		100.0%	100%	100.0%
Conversions from input grade to product		69.6%	100.0%	153.4%
Refining Charges		0	0	0
Price		US$ 39.60/kg	US$ 0.88/kg	US$ 3,675/kg
Calculate Contained Metal
Nb2O5		700 kg
TiO2			2,500 kg
Sc				6 kg
Calculate Saleable Metal (conversion to product, discounted by recovery)
Nb		401 kg
FeNb		617 kg
TiO2			1,008 kg
Sc (as Sc2O3)				8.57 kg
Calculate Block Dollar Value for Each Metal
FeNb		US$ 15,890
TiO2			US$ 887
Sc				US$ 31,504
Total Block Value	US$ 48,281
Block Value per tonne	US$ 482.81/t

Source: Nordmin, 2019

(1) Stored as PPM in the block model. Sc % = Sc ppm/10,000.

(2) Overall metallurgical recovery, including all losses

Figure 12-2 through Figure 12-4 provide a grade-tonne curve for the deposit using various NSR cut-off grades, (CoG). It includes only Measured and Indicated material and shows average grades for each grade variable. All Inferred material is treated as having a zero-grade value in this mineral reserve estimation.

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Source: Optimize Group, 2022

Figure 12-2: NioCorp Grade (Nb₂O₅)-Tonne Curves Based on NSR Cut-Off

Source: Optimize Group, 2022

Figure 12-3: NioCorp Grade/Tonne Curves Based on NSR Cut-Off (TiO₂)

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Source: Optimize Group, 2022

Figure 12-4: NioCorp Grade (Sc ppm) - Tonne Curves Based on NSR Cut-Off

To establish the initial boundary of the mine design and to assure inclusion of all potential Mineral Reserves, a minimum CoG of US$ 180/t was used based on the estimated costs shown in Table 12-3.

Table 12-3: Operating Costs Used for Mine Design NSR Cut-off

Item	Estimated Costs (US$/t)
Mining (1)	50.00
Processing	125.00
G&A	5.00
Total (2)	US$ 180.00

Source: Nordmin, 2019

(1) Includes backfill

(2) Values used here differ from the economic model generated from the final overall site design. Optimize Group is satisfied that the values used were applicable to establishing the correct and optimum mining design.

12.1.4  Mine Design

Potential mining areas were identified using stope optimization within Deswik.SO StopeOptimizer software. The stope optimizer output was reviewed on a level-by-level basis, and a 3D mine design was generated. The estimated cut-off NSR value (CoNSR) of US$ 180/t from the SRK study was used as a starting point for this analysis. Generally, stopes would be selected based on the minimum CoG or CoNSR. As the CoNSR value is much lower than the resulting average stope NSR value, the Co NSR was not the decisive factor in the stope optimization process. Rather than using only a minimum CoNSR, the mine design also targeted an average cut-off Nb₂O₅ grade of 0.679% and targeted higher annual ferroniobium production during the first five years of production. With a milling constraint of 2,764 tpd, the steady-state life of mine average annual ferroniobium

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production during full production years was 7,450 tonnes annually. This strategy results in a LOM NSR average value of US$ 563.06/t. The identified mining blocks provide an approximate 38-year LOM. The design includes stopes, development accesses, and necessary infrastructure. Figure 12-5 shows the completed mine design.

Source: Optimize Group, 2022

Figure 12-5: Completed Mine Design

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12.2 Reserves

The 2019 Mineral Reserves were classified using SK 1300 Standards. Indicated Mineral Resources were converted to Probable Mineral Reserves by applying the appropriate modifying factors, as described earlier in this section, to potential mining block shapes created during the mine design process.

The underground mine design process resulted in a mine plan with a Mineral Reserve Estimate of 36.7 Mt (diluted) with an average grade of 0.81% Nb₂O₅, 2.92% TiO₂, and 70.2 ppm Sc. This estimate is based on a mine design using elevated CoGs and applying the US$ 180/t NSR CoG to capture all potential Mineral Reserves within the design and an average cut-off grade of 0.68% Nb₂O₅. These numbers include a 95% mining ore recovery to the designed wireframes (sill pillar recovery is 62.5%) in addition to applying approximately 6% unplanned dilution as described in Section 12.1.1.

Table 12-4 summarizes the underground reserves as of June 30, 2022.

Table 12-4: Underground In Situ Mineral Reserves Estimate for Elk Creek, Effective Date June 30, 2022

Classifi- cation	Tonnage (kt)	Nb2O5 Grade (%)	Contained Nb2O5 (t)	Payable Nb (t)	TiO2 Grade (%)	Contained TiO2 (t)	Payable TiO2 (t)	Sc Grade (ppm)	Contained Sc (t)	Payable Sc2O3 (t)
Proven	-	-	-	-	-	-	-	-	-	-
Probable	36,656	0.81	297,278	170,409	2.92	1,071,182	431,793	70.2	2,573	3,677
Total	36,656	0.81	297,278	170,409	2,92	1,071,182	431,793	70,2	2,573	3,677

Source: Optimize Group, 2022. All figures are rounded to reflect the relative accuracy of the estimates. Totals may not sum due to rounding.

● The Qualified Person for the Mineral Reserve estimate is Optimize Group Inc. The estimate has an effective date of June 30, 2022

● The Mineral Reserve is based on the mine design and mine plan, utilizing an average cut-off grade of 0.679% Nb₂O₅ with an NSR of US$ 180/mt.

● The estimate of Mineral Reserves may be materially affected by metal prices, environmental, permitting, legal, title, taxation, socio-political, marketing, infrastructure development, or other relevant issues.

● The economic assumptions used to define Mineral Reserve cut-off grade are as follows:

○ Annual life of mine (LOM) production rate of ~7,450 tonnes of FeNb/annum during the years of full production.

○ Initial elevated five-year production rate ~ 7,500 tonnes of FeNb/annum when full production is reached.

○ Mining dilution of ~6% was applied to all stopes and development, based on 3% for the primary stopes, 9% for the secondary stopes, and 5% for ore development.

○ Mining recoveries of 95% were applied in longhole stopes and 62.5% in sill pillar stopes.

Parameter	Value	Unit
Mining Cost	42.38	US$/t mined
Processing	106.70	US$/t mined
Water Management and Infrastructure	16.62	US$/t mined
Tailings Management	2.01	US$/t mined
Other Infrastructure	5.47	US$/t mined
General and Administrative	8.91	US$/t mined
Royalties/Annual Bond Premium	8.34	US$/t mined
Other Costs	6.29	US$/t mined
Total Cost	196.72	US$/t mined

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Nb2O5 to Niobium conversion	69.60	%
Niobium Process Recovery	82.36	%
Niobium Price	39.60	US$/kg
TiO2 Process Recovery	40.31	%
TiO2 Price	0.88	US$/kg
Sc Process Recovery	93.14	%
Sc to Sc2O3 conversion	153.40	%
Sc Price	3,675.00	US$/kg

● Price assumptions are as follows: FeNb US$ 39.60/kg Nb, Sc2O3 US $3,675/kg, and TiO2 US $0.88/kg. Price assumptions are based upon independent market analyses for each product as of June 30, 2022

● Price and cost assumptions are based on the pricing of products at the “mine-gate,” with no additional down-stream costs required. The assumed products are ferroniobium (metallic alloy shots consisting of 65%Nb and 35% Fe), a titanium dioxide product in powder form, and scandium trioxide in powder form.

● The Mineral Reserve has an average LOM NSR of US$ 563.06/tonne.

● Optimize Group has provided detailed estimates of the expected costs based on the knowledge of the style of mining (underground) and potential processing methods (by 3rd party Qualified Persons).

● Mineral reserve effective date is June 30, 2022. The financial model was run after the estimate of the NSR above, which reflects a total cost per tonne of US$ 196.72 versus US$ 189.91. This is not considered a material change.

● Price variances for commodities are based on independent market studies versus earlier projected pricing. The independent market studies do not have a negative effect on the reserve.

12.3  QP Opinion and Relevant Factors

It is Optimize Group’s opinion that there are no known environmental, permitting, legal, socio-economic, marketing, political, or other factors which could materially affect the underground Mineral Reserve Estimate. In addition, realistic and justifiable mining factors in determining the mine plan and schedule for reporting mineral reserves. These factors include geotechnical considerations, ore loss, dilution, mine extraction rates and metallurgical recovery.

The pricing for Niobium, Scandium and Titanium used to support the Mineral Reserve estimate was provided by NioCorp. NioCorp has relied on third party market information, relying on Roskill (2019) for niobium, OnG Commodities (2019) for Scandium and USGS (2019) for titanium. It is Optimize Group’s opinion that the pricing used is adequate and appropriate for use in estimating Mineral Reserves. Further details on the market assumptions and timeframes analyzed are disclosed in Chapter 16 of this report.

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13. MINING METHODS

13.1 Geotechnical Design Parameters

Refer to Section 7.3.

13.2 Hydrogeology Design Parameters

Refer to Section 7.4.

13.3 Mine Design

13.3.1 Selection of Mining Method

The mining method selected for this ore body was based on economic parameters and geotechnical information, ensuring it was suitable for the mineralization geometry. Due to its depth and requirement for selectivity in mill feed grades, the underground longhole stoping method (LHS) was selected. Given the bulky geometry of the deposit, a block caving or sub-level caving method could have possibly been economically viable. However, the limited selectivity of such methods would not allow for optimizing the higher value of this deposit given their production constraints. To maximize the recovery of the high-grade zones, a longhole stoping method utilizing paste backfill was used.

The stopes dimensions are 15 m wide, and stope length varies based on Nb₂O₅ mineralization grade to a maximum of 25 m per panel with a level spacing of 40 m. The variation on stope length allowed for optimization of the Nb₂O₅ grade with a minimal increase to operating costs. The level spacing of 40 m was beneficial to operating and sustaining capital costs. Each block is mined with a bottom-up sequence. A partial sill pillar level is designed to be left between these two mining fronts/blocks. The extraction of ore from the partial sill pillar level is expected to be 62.5% using production up-holes through 25 m of the 40 m thick sill pillar and is accounted for within the reserves. This methodology will allow partial mining of ore on the sill pillar level, while at the same time allowing the development of the lower mining block and establishing an early start to the mining of the upper mining block. Using this approach minimizes the impact on initial capital investment. The backfill was designed to have an adequate strength to allow for mining adjacent to filled stopes, thus eliminating the need for rib pillars. The backfill will have an adequate strength to allow for mining adjacent to filled stopes, thus eliminating the need for rib pillars.

The mine design process involved using a minable shape optimization software to determine potentially mineable areas based on an estimated minimum cut-off NSR (CoNSR) value, Nb₂O₅ grades and mining dimensions parameters. As the CoNSR value is much lower than the resulting average stope NSR value, the CoNSR was not the decisive factor in the stope optimization process. Mining dilution of approximately 6% was applied to all stopes and development, based on 3% for the primary stopes, 9% for the secondary stopes, and 5% for ore development. The mining dilution was added to the designed tonnage to account for unplanned sources of dilution such as backfill and host rock around the periphery of the ore mass. An ore recovery factor of 95% was applied to account for unrecoverable ore left within the stopes.

The mine design and schedule were based on a milling constraint (2,764 tpd), provided by NioCorp, to produce approximately 7,450 t/y of ferroniobium during the years of full production and a LOM of over 38 years. Optimization work indicated that the grade of Nb₂O₅, (0.81%) at a unit NSR over US$ 500/t could sustain and produce a consistent ferroniobium production over the

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LOM. The mill production rate was established at 2,764 t/d from which an annualized production averaging approximately 7,450 t of ferroniobium per year during the years of full production is derived. Scandium trioxide and titanium dioxide accompany the ferroniobium production in the mine plan. Nordmin favoured a higher NSR value in its approach to maximize the LOM NPV for production scheduling while at the same time maintaining the annual ferroniobium requirement.

13.3.2  Stope Optimization

As mentioned in Section 13.3.1, the minable shape optimization software provided by Deswik was used to determine potentially mineable areas based on NSR, Nb₂O₅ grades and mining dimensions parameters. The estimated cut-off NSR value (CoNSR) of US$ 180/t from the SRK study was used as a starting point for this analysis. Generally, stopes would be selected based on the minimum CoG or CoNSR. As the CoNSR value is much lower than the resulting average stope NSR value, the CoNSR was not the decisive factor in the stope optimization process. Rather than using a minimum CoNSR, the mine design also targeted an average cut-off Nb₂O₅ grade of 0.679% with a milling constraint of 2,764 tpd which resulted in a steady-state average annual ferroniobium production of 7,450 tonnes annually during the years of full production. This strategy results in a LOM NSR average value of US$ 563/t. Figure 13-1 and Table 13-1 show the stopes optimized for varying CoNSR scenarios. An average dilution of approximately 6% was added to the designed tonnage which accounts for unplanned sources of dilution such as backfill and the host rock around the periphery of the ore mass while a recovery factor was applied to account for unrecoverable material which left within the stopes.

Source: Optimize Group, 2022

Figure 13-1: Undiluted Stope Optimization Results for Varying NSR Cut-Offs

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Table 13-1: Undiluted Stope Optimization Results for Varying NSR Cut-offs

NSR Cut-off	Tonnes	Nb2O5	Sc	TiO2	NSR	LREO	MREO	HREO	TREO
(US$/t)	(t)	(%)	(ppm)	(%)	(US$/t)	(%)	(%)	(%)	(%)
100	213,332,432	0.478	55.7	2.08	408	0.28	0.0273	0.0287	0.335
180	195,071,823	0.504	58.9	2.2	432	0.276	0.0281	0.0292	0.333
200	183,223,208	0.523	61.1	2.28	447	0.28	0.0289	0.0296	0.338
300	144,533,304	0.6	67.8	2.51	501	0.294	0.0319	0.0306	0.356
400	114,081,216	0.664	72.7	2.62	542	0.293	0.0341	0.0314	0.359
500	69,666,027	0.753	80	2.76	601	0.293	0.0379	0.0337	0.364
600	29,008,380	0.878	89.1	3.06	678	0.309	0.0445	0.039	0.392
650	17,334,253	0.935	93.5	3.19	715	0.32	0.047	0.0409	0.408
700	8,391,296	0.988	99.4	3.31	758	0.331	0.0502	0.044	0.424

Source: Optimize Group, 2022

13.3.3  Stope Design

Figure 13-2 shows typical stopes cross section. The stope width is a constant 15 m, and vertical height is 40 m from floor to floor. The length of the stopes is on average 19 m ranging from 10 m to a maximum panel length of 25 m. Figure 13-3 shows a typical level arrangement of the stopes, x-cut, footwall drive, ramp and other infrastructures servicing a level. The mine plan stope orientation is perpendicular to the general strike of the deposit, which is 20° off the measured principal stress. This offset will not have a significant impact on stope stability. The actual mine plan stope lengths have a maximum length of 25 m in both fresh and moderately weathered rock, which is a conservative design in relation to the stability assessment described in Section 13.1, Geotechnical Design Parameters.

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Source: Nordmin, 2019

Figure 13-2: Stopes and Cross-Cut Accesses (Cross Section View)

Source: Optimize Group, 2022

Figure 13-3: Level Layout with Stopes and Footwall Accesses (Rotated View Looking North)

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13.3.4  Development Design

The stopes are accessed through a footwall drive with about 25 m offset from the stopes. The cross-cuts (x-cuts) are driven in the center of the stopes from the footwall drives, as shown in Figure 13-4. These drives are connected by the ramp system, ventilation raises and on some levels they are connected to the production shaft. Most of the mine infrastructure is located in waste, but some areas can be in lower grade material as it gets closer to the ore body.

The designed vertical extent of the mine is 600 m with a bottom elevation of -535 m. The ventilation shaft is designed to a 530 m depth. The production shaft is designed to a 755 m depth. The production shaft and related crushing and conveying system is complemented with an ore pass and waste pass system that results in an overall material handling system that has suitable ore storage above and below the crusher station.

Source: Optimize Group, 2022

Figure 13-4: Completed Mine Design (Plan View)

Figure 13-5 shows the completed mine design main infrastructure area. The shafts, internal raises, and underground infrastructure included in the design are discussed in other subsections. The two mining horizons are generally mined simultaneously. Altogether, they provide an estimated LOM of 38 years.

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Source: Optimize Group, 2022

Figure 13-5: Completed Mine Design - Main Infrastructure (Looking South)

Figure 13-6 and Figure 13-7 show the mine design coloured by Nb₂O₅ grade and NSR, respectively.

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Source: Optimize Group, 2022

Figure 13-6: Mine Design Coloured by Nb₂O₅ Grade.

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Source: Optimize Group, 2022

Figure 13-7: Mine Design Coloured by NSR

Table 13-2 summarizes the mine design by activity type.

Table 13-2: Mine Design Summary - by Activity Type

General Summary	LOM Statistics
Ore Tonnes (t)	36,655,676
FeNb (t)	262,168
Nb2O5 Grade (%)	0.811
Sc Grade (ppm)	70.2
TiO2 Grade (%)	2.92
Development Ore Tonnes (t)	908,725

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Stope Production Tonnes (t)	35,746,951
Waste Tonnes (t)	3,071,390
Total Tonnes Moved (t)	39,727,066
Lateral Development:
Main Ramp - 5.0 x 5.3 (m)	5,361
Ramp Access to Level - 4.5 x 5.3 (m)	1,105
Shaft Access to Level (m)	775
Footwall Access - 4.5 x 5.3 (m)	8,259
Fresh Air Raise Access - 4.5 x 5.3 (m)	698
Return Air Raise Access - 4.5 x 5.3 (m)	1,047
Ore Pass Access - 4.5 x 5.3 (m)	1,148
Waste Pass Access - 4.5 x 5.3 (m)	258
Stope Access Drift - 4.3 x 4.0 (m)	46,439
Other Lateral Development - Shop, Crusher, Sumps, etc. (m)	1,622
Total Lateral Development (m)	66,712
Vertical Development:
Production Shaft - 6.0 m Finished Diameter (m)	755
Ventilation Shaft - 6.0 m Finished Diameter (m)	525
Fresh Air Raise - 3.3 m Diameter (m)	643
Return Air Raise - 3.3 m Diameter (m)	660
Ore Pass - 3.0 x 3.0 (m) and Ore Pass Fingers - 2.0 x 2.0 (m)	939
Waste Pass - 3.0 x 3.0 (m)	334
Waste Pass Fingers - 2.0 x 2.0 (m)	143
Other Vertical Development - Bins, Conical Sump (m)	95
Total Vertical Development (m)	4,094

Source: Optimize Group, 2022

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13.3.5  Mine Access

The underground mine will be accessed from the surface at a collar elevation of 354.6 m (1,163.4 ft) ASL, via twin 6.0 m (19’-8”) diameter concrete lined shafts, named the “production shaft” and the “ventilation shaft” (see Figure 13-8). Coordinates for the shafts are N4461430.850 m, E739499.590 m for the production shaft and N4461310.000, E739663.000 m for the ventilation shaft. The shafts are excavated by means of conventional shaft sinking and will be combined with freezing down to the potential water-bearing contact between the Pennsylvanian sediments and carbonatite unit. Below that contact, inflow control in the shafts will be controlled by grouting in advance of shaft sinking. This method, unlike a raisebore method of excavation, allows control of water inflows.

The production shaft will facilitate the movement of larger mining equipment, workforce, services, material hoisting, and act as the supply route for the mine ventilation system. The production shaft is excavated to a lower elevation than in the previous Feasibility Studies. This allows earlier access to higher grade ore in the central portion of the mine and to also access higher grade ore in the lower mining block with a more efficient material handling system.

The ventilation shaft will be dedicated to moving workforce and smaller material, hoisting for initial lateral development, as well as act as an exhaust route for the mine ventilation system. A second temporary hoist, hoist room, and headframe is installed for the ventilation shaft sinking process and will be utilized to hoist waste from initial lateral mine development prior to the completion and installation of the permanent hoisting arrangement in the production shaft.

Main access to the lower working levels will be gained by means of the production shaft which records a shaft bottom elevation of -400.4 m (-1313.65 ft), with stations at the -15.4 m (-50.5 ft), -175.4 m (-575.5 ft), -295.4 m (-969 ft), and -335.4 m (-1,100.5 ft), and access to the spill pocket at -370.4 m (-1,215 ft).

Stations and underground development on the -15.4 m (-50.5 ft), -175.4 m (-575.5 ft) levels, allow for easy access between the production shaft and the ventilation shaft.

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Source: Nordmin, 2019

Figure 13-8: Underground Mine Access Via Twin Concrete Lined Shafts

13.3.5.1 Shaft Layouts

Access to the underground mine is via either the 6.0 m diameter concrete lined production shaft or the 6.0 m diameter concrete lined ventilation shaft. Atop the production shaft lies a 71.5 m (235 ft) tall headframe, with three sheave decks for five rope sheaves. The production shaft will host the two production skips, the main service cage and counterweight, and auxiliary cage as well as house all services to the underground including:

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● 8.0” diameter dewatering lines, a quantity of three;

● 8.0” compressed air line;

● 8.0” and 6.0” slick lines;

● 4.0” process water line;

● 2.0” fuel line;

● 4.0”, 2.0” and 1.0” spare cables;

● 13.8 kV power lines; and

● communication, fibre, leaky feeder and ground cables (see Figure 13-9).

Source: Nordmin, 2019

Figure 13-9: Production Shaft Layout

Services within the production shaft are located for ease of access for required inspections from the skips, service cage, and the auxiliary cage. The production shaft will be partitioned into two

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main areas with the use of buntons and brattice. The twin skips running on steel guides will be separated from the rest of the shaft by brattice panels throughout the length of the shaft. The other section of the shaft will house the auxiliary cage, the service cage and the service cage counterweight. A 5 m shaft set spacing is used in both the ventilation shaft and the production shaft.

Access to the shaft will be from within the headframe. Additional safety gates and hydraulically actuated collar doors will reside atop the shaft at collar elevation.

The 6.0 m diameter concrete lined ventilation shaft will be used to host a secondary auxiliary cage, identical to the production shaft auxiliary cage and minimal services, including:

● 13.8 kV power cables;

● communication, fibre, leaky feeder and ground cables; and

● 8.0” and 6.0” slick lines.

Similar to the production shaft headframe, the ventilation shaft headframe will house a single sheave deck, a set of collar doors and safety gates.

The ventilation shaft is partitioned into two main areas, one dedicated for the auxiliary cage, and one dedicated for the ventilation system. Both sections will be segregated by the use of brattice panels from the lower depths of the shaft to the top of the ventilation sweep for the exhaust system. This will guarantee an unobstructed ventilation pathway for the mine air exhaust system (see Figure 13-10).

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Source: Nordmin, 2019

Figure 13-10: Ventilation Shaft Layout

13.4 Production Schedule

The production schedule is based on the mine design and reserves discussed in previous sections.

13.4.1  Productivity

Productivities were developed from first principles. Input from mining contractors, blasting suppliers and equipment vendors, were used for the key parameters. The rates developed from first principles were adjusted based on benchmarking and the experience and judgment of the mine design team.

The productivity rates used for mine scheduling are shown in Table 13-3, followed by a description of the general and activity-specific parameters upon which the productivity rates are based.

Typical dimensions by heading types are presented in Table 13-4. These will all be developed by contractors in accordance with the productivity rates and levelled in the schedule by crew assignments.

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Table 13-3: Productivity Rates

Activity	Type	Dimensions	Rate
Lateral	Priority Face	See Table 16-9	5.0 m/d
Development	Non-Priority Face		3.0 m/d
	Shaft Station	Varies	2.0 m/d
Vertical	Production Shaft	6.0 m diameter	2.3 m/d
Development	Ventilation Shaft	6.0 m diameter	2.3 m/d
	Fresh Air Raise	3.3 m diameter	3.6 m/d
	Return Air Raise	3.3 m diameter	3.6 m/d
	Ore Pass	3 m x 3 m	3.6 m/d
	Ore Pass Fingers	2 m x 2 m	3.6 m/d
	Waste Pass	3 m x 3 m	3.6 m/d
	Waste Pass Fingers	2 m x 2 m	3.6 m/d
	Bins and Conical Sump	Varies	1.3 m/d
Individual	Slot Development	-	10.5 d
Stoping	Drilling	-	250 m/d
	Stope Production	-	930 t/d
	Backfill Preparation	-	10.0 d
	Backfilling	-	1200 m3/d
	Backfill Curing	-	28.0 d

Source: Nordmin, 2019

Table 13-4: Dimensions by Heading Types

Heading Types	Width (m)	Height (m)	Area (m2)
Ramp	5.0	5.3	27
Elect Sub	4.5	5.3	24
FAR Access	4.5	5.3	24
FW Drift	4.5	5.3	24
Ore/Waste Access	4.5	5.3	24
X-Cuts	4.3	4.0	17
Remucks	4.5	5.0	23
Sump	4.5	5.3	24
Refuge Station	4.5	5.3	24

Source: Nordmin, 2019

General Parameters

Table 13-5 provides the general schedule parameters applicable to all underground mining

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activities.

Table 13-5: Workforce Schedule Parameters for Underground Mining

Schedule Parameters	Units	Value
Annual Mining Days	days/year	365
Mining Days per Week	days/week	7
Shifts per Day	shifts/day	2
Scheduled Shift Length	hrs/shift	12
Scheduled Deductions:
- Travel Time Between Underground and Surface	hrs/shift	1.00
- Workplace Examinations and Equipment Pre-shift Inspections	hrs/shift	0.25
- Lunch	hrs/shift	0.50
- Breaks	hrs/shift	0.50
Total Scheduled Deductions	hrs/shift	2.25
Operating Time (Scheduled Shift Length Less Scheduled Deductions)	hrs/shift	9.75
Effective Time (Operating Time Reduced to a 50 Minute Hour, i.e., Multiplied by 83.3%)	hrs/shift	8.125

Source: Nordmin, 2019

Table 13-6 provides the ground support requirements.

Table 13-6: Ground Support Requirements

Geotechnical Zone	Q*	Excavation	Support Categories	Bolt Length	Bolt Spacing	Other Support
Footwall, highly weathered (6%)	0.4 to 6.2 (Very Poor)	Main Ramp	3-Bolts, mesh and shotcrete	2.5 m	1.2 m	Fully grouted rebar, mesh, 5 cm shotcrete
		FW Access	2-Systematic bolting	2.5 m	1.2 m	Split sets and mesh
		Stope Access	2-Systematic bolting	2.5 m	1.6 m	Split sets and mesh
Footwall, moderately weathered (24%)	3.2 to 13.8 (Poor-Fair)	Main Ramp	2-Systematic bolting	2.5 m	1.2 m	Fully grouted rebar, mesh
		FW Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets
		Stope Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets
Footwall,	5.9 to	Main Ramp	1-Spot bolting	2.5 m	1.6 m	Grouted

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slightly weathered (70%)	28.1 (Fair-Good)		(15/10 m)			rebar
		FW   Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets
		Stope Access	1-Spot bolting (15/10 m)	2.5 m	1.6 m	Split sets

*(%) Amount of Expected Ground

Source: SRK, 2017

The mine plan has conservatively allowed for grouted rebar in the back (roof) of all excavations. Split sets were acceptable in walls of excavations, but not in the back.

13.4.2  Shaft Sinking – Production Shaft and Ventilation Shaft

Shaft sinking operations at both shafts will be carried out simultaneously. This allows the initial lateral development to begin from the bottom of the ventilation shaft while the production shaft continues to be excavated to a lower elevation to facilitate extraction of higher niobium grade stopes located at the lower levels.

The 6.0 m (inside diameter) Production Shaft is excavated to a depth of 755 m. The shaft is excavated using conventional shaft sinking methods in conjunction with a freezing process through the first 200 m from the surface to ensure ground and water control. Upon completion of the first 200 m section, the shaft sinking continues, but freezing is no longer required to reach the bottom elevation. The rate of excavation averages 2.30 m/d; this rate was developed in collaboration with contractors for the material expected to be encountered. The average rate includes sinking, lining, furnishing and adjustments in rate due to rock types and shaft depth. The material is removed by a temporary shaft sinking hoist, hoist room, and headframe system and placed in the lined temporary stockpile location adjacent to the shaft.

The Ventilation Shaft is excavated with the same diameter and method as the production shaft, but only to a depth of 530 m. Conventional shaft sinking is combined with freezing down to the potential water-bearing contact between the Pennsylvanian sediments and carbonatite unit. This method, unlike a raisebore method of excavation, allows control of potential water inflows. A second temporary hoist, hoist room, and headframe is installed for the sinking process and will be utilized to hoist waste from lateral mine development prior to the completion and installation of the permanent hoisting arrangement in the production shaft.

13.4.3  Development and Production Schedule

The production and development schedules were completed using the Deswik scheduling software. The production schedule is based on the rate assumptions shown in Table 13-7.

A delay of 28 days was used before driving on paste backfill or mining adjacent to a paste backfilled stope. These delays account for curing time as well as multiple pours.

The mining operation schedule is based on 365 days/year, 7 days/week, with two 12 hour shifts each day. A production rate of 2,764 t/d was targeted with a ramp-up to full production as quickly as possible. The schedule timeframe is monthly for the pre-production period, two years for production, then quarterly for three years, and annually for the remainder of the LOM.

Production shaft and ventilation shaft sinking preparation begins eight months after the commencement of detailed engineering with the actual sinking beginning five months later and subsequent lateral mine development beginning nine months later. Production stoping begins

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sixteen months after the start of lateral development, with a production ramp-up period through the next six months, after which the mine and plant are operating at full capacity.

Table 13-7 shows the annual mine production schedule, and Figure 13-11 shows the mine production schedule coloured by year.

Table 13-7: Mine Production Schedule

Year	Ore Tonnes	Nb2O5	TiO2	Sc	Waste Tonnes	Backfill Volume
	(t)	(%)	(%)	(ppm)	(t)	(m3)
Year 0	0	0.00	0.00	0.00	147,139	0
Year 1	5,594	0.74	2.85	72.23	591,609	0
Year 2	743,439	0.84	3.21	77.55	505,301	162,671
Year 3	1,044,000	0.82	3.06	78.06	486,568	317,214
Year 4	1,044,000	0.80	2.88	74.27	192,683	316,849
Year 5	1,043,999	0.82	3.03	76.68	176,912	332,281
Year 6	1,044,001	0.82	3.02	73.20	140,449	349,031
Year 7	1,044,000	0.81	2.93	75.96	127,374	321,897
Year 8	1,044,000	0.78	2.87	69.86	79,924	340,996
Year 9	1,044,000	0.79	2.82	70.91	88,099	332,332
Year 10	1,044,000	0.79	2.90	67.35	92,819	342,684
Year 11	1,044,002	0.80	2.79	68.03	42,639	352,356
Year 12	1,044,033	0.78	2.73	67.80	8,465	343,903
Year 13	1,044,001	0.80	2.87	67.23	21,727	340,628
Year 14	1,044,001	0.80	2.92	68.92	36,905	344,280
Year 15	1,044,002	0.80	2.90	69.89	79,784	341,113
Year 16	1,043,998	0.79	2.90	69.75	45,485	329,332
Year 17	1,043,999	0.80	2.87	72.33	8,876	351,974
Year 18	1,044,000	0.80	2.93	65.92	54,616	332,620
Year 19	1,044,000	0.80	2.93	72.03	23,922	349,750
Year 20	1,044,001	0.79	2.96	68.83	0	355,535
Year 21	1,044,000	0.79	2.80	68.91	0	355,205
Year 22	1,044,000	0.80	2.76	69.47	4,309	346,011
Year 23	1,044,041	0.80	2.94	69.31	0	375,852
Year 24	1,044,001	0.78	2.82	67.25	0	355,948
Year 25	1,043,999	0.82	3.05	61.02	0	349,544
Year 26	1,043,999	0.83	2.88	59.24	0	340,005

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Year 27	1,043,999	0.83	3.01	69.70	103,439	342,093
Year 28	1,044,000	0.79	2.75	71.41	12,346	342,081
Year 29	1,044,001	0.82	2.96	70.04	0	385,824
Year 30	1,043,999	0.85	3.12	62.96	0	359,392
Year 31	1,043,998	0.83	2.93	71.69	0	371,231
Year 32	1,044,004	0.81	2.96	69.03	0	350,524
Year 33	1,044,001	0.79	2.90	74.92	0	367,303
Year 34	1,044,000	0.82	3.00	68.77	0	390,992
Year 35	1,034,060	0.91	3.08	72.22	0	377,351
Year 36	748,810	0.88	2.89	72.31	0	323,803
Year 37	511,665	0.82	2.88	71.02	0	276,378
Year 38	204,030	1.03	2.92	103.26	0	106,592
Year 39	0	0.00	0.00	0.00	147,139	0
Totals	36,655,676	0.81	2.92	70.2	3,071,390	12,373,574

Source: Optimize Group, 2022

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Source: Optimize Group, 2022

Figure 13-11: Mine Production Schedule - Coloured By Year

13.5 Mining Operations

13.5.1  Production

The ore feed to the plant primarily comes from the stope production as the development contributes to less than 3% of the total ore. Stopes are mined using the longhole open stoping method. Individual stope blocks are designed to be 15 m wide, up to 25 m long oriented roughly parallel to the main stress. Levels are spaced 40 m apart, and each stope block has top and bottom access called the crosscut (x-cut: 4.3 m x 4 m flat back drifts).

Stopes are drilled downward from the top access using 114 mm (4.5 in) diameter holes. Initial opening is done using stope slots drilled with a slot reamer machine and blast holes. A level-by-level bottom up, primary/secondary extraction sequence is followed. Primary stopes are backfilled with high strength cemented paste backfill. Secondary stopes are backfilled with high strength cemented paste backfill when more than one panel needs to be mined adjacent to one another. When development waste rock is not available for backfilling secondary stopes, low strength paste backfill is utilized as needed.

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All blasting is performed with bulk emulsion. The slot is expanded first followed by one or two mass blasts to complete the stope.

Ore is mucked from the lower x-cut access using a 6.2 m³ (14 t) LHD with remote control capability. The ore is transported by the LHD to either an ore pass directly or to a remuck bay to maximize the efficiency of the stope mucking operations as a function of the haulage distance. When needed, a second LHD and a fleet of 40-tonne haul trucks are used to transport ore from the remuck bays to the grizzly feeding the underground material handling system. Multiple remuck bays are used on each level to avoid interference between the LHD and the haul trucks.

13.5.2  Development

Lateral development includes interlevel ramps, level accesses, stope accesses, and short connecting drifts for ventilation. The interlevel ramp system is 5 m wide by 5.3 m high at a maximum 15% gradient. Level accesses is 4.5 m wide by 5.3 m high and is mined higher at the remuck bays to allow the haul trucks to be loaded by the LHD. Stope access drifts 4.3 m wide by 4 m high. Stope accesses are oriented perpendicular to the strike of the orebody.

The lateral development is sized for the operation of the mining equipment fleet selected for the operation. The development profiles include allowances for ventilation ducting and services.

Raiseboring is used to establish ventilation connections between level access drifts.

13.5.3  Truck and LHD Haulage

The mine plan assumes that 6.2 m3 (14 t) LHDs load the 40-tonne haul trucks from remuck bays that are strategically located throughout the development workings. Ore haulage distance and cycle times were calculated using a haulage model built in Microsoft Excel^® and are based on estimated underground speeds, loading travel time, efficiency, productivity, distance, bucket capacity, filling factor, turnaround time, time to dump and utilization rate, for LHD and truck as shown in Table 13-8 to Table 13-10. At levels 210 to 610, the cycle time corresponds to the sum of the time spent moving the LHDs from the Stopes to the ore pass plus the time the LHDs in the end of the ore pass at level 610 take to complete a cycle to the grizzly. At levels 650 to 890, the cycle time is calculated by summing the LHD cycle haulage from the stope to the Mucking bay, where the truck will be completely filled, requiring 3 trips plus the time for the fully loaded truck to dump the material into the grizzly in the level 610. The outputs from the haulage profile module are a one-way haulage distance and an average truck and LHD cycle time (round trip).

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Table 13-8: LHD Hauling parameters upper mining block

LEVEL UPPER MINING BLOCK	DTM GRIZZLY WEST	DTM GRIZZLY EAST	250		290		330		370		410		450		490		530		570		610
			East	West	East	West	East	West	East	West	East	West	East	West	East	West	East	West	East	West	East	West
AVG Total Distance (km)	0.3	0.1	0.30	0.00	0.17	0.13	0.19	0.09	0.16	0.17	0.15	0.15	0.18	0.20	0.07	0.18	0.10	0.12	0.11	0.12	0.19	0.15
Horizontal Distance (Km)	0.3	0.1	0.30	0.00	0.17	0.13	0.19	0.09	0.16	0.17	0.15	0.15	0.18	0.20	0.07	0.18	0.10	0.12	0.11	0.12	0.19	0.15
Horizontal Velocity Empty and load (Km/h)	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8	8
Bucket capacity (m3)	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2	6.2
Bucket capacity (t)	14.0	14.0	14	14	14	14	14	14	14	14	14	14	14	14	14	14	14	14	14	14	14	14
filling factor	0.9	0.9	0.85	0.00	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85
Bucket load (min)	0.2	0.2	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20	0.20
Turnaround time (min)	0.2	0.2	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22	0.22
Load travel time (min)	2.0	0.8	2.2	0.0	1.3	1.0	1.4	0.7	1.2	1.3	1.1	1.1	1.4	1.5	0.5	1.4	0.8	0.9	0.8	0.9	1.4	1.1
Empty travel time (min)	2.0	0.8	2.2	0.0	1.3	1.0	1.4	0.7	1.2	1.3	1.1	1.1	1.4	1.5	0.5	1.4	0.8	0.9	0.8	0.9	1.4	1.1
Dump (min)	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08	0.08
Utilization Rate (%)	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85	0.85
Cycle time (min)	0.09	0.04	0.10	0.00	0.06	0.05	0.07	0.04	0.06	0.06	0.05	0.05	0.06	0.07	0.03	0.06	0.04	0.05	0.04	0.04	0.06	0.05

Source: Optimize, 2022

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Table 13-9: LHD Hauling parameters lower mining block 

LEVEL LOWER MINING BLOCK	650	690	730	770	810	850	890
LHD
AVG Total Distance (km)	0.43	0.38	0.40	0.44	0.39	0.36	0.37
Horizontal Distance (Km)	0.29	0.23	0.25	0.29	0.24	0.21	0.23
Vertical Distance (Km)	0.15	0.15	0.15	0.15	0.15	0.15	0.14
Horizontal Velocity Empty and load (Km/h)	8	8	8	8	8	8	8
Vertical Velocity Empty (Km/h)	10	10	10	10	10	10	10
Vertical Velocity Loaded (Km/h)	6	6	6	6	6	6	6
Bucket capacity (m3)	6.2	6.2	6.2	6.2	6.2	6.2	6.2
Bucket capacity (t)	14	14	14	14	14	14	14
Filling factor	0.85	0.85	0.85	0.85	0.85	0.85	0.85
CYCLE
Bucket load (min)	0.20	0.20	0.20	0.20	0.20	0.20	0.20
Turnaround time (min)	0.22	0.22	0.22	0.22	0.22	0.22	0.22
Load travel time (min)	3.6	3.2	3.4	3.6	3.3	3.1	3.1
Empty travel time (min)	3.0	2.6	2.8	3.1	2.7	2.5	2.6
Dump (min)	0.08	0.08	0.08	0.08	0.08	0.08	0.08
Utilization Rate (%)	0.85	0.85	0.85	0.85	0.85	0.85	0.85
Cycle time (min)	8.3	7.4	7.8	8.5	7.6	7.1	7.2

Source: Optimize, 2022

Table 13-10:Truck Hauling parameters lower mining block 

–TRUCK PRODUCTIVITY
LEVEL LOWER MINING BLOCK	650	690	730	770	810	850	890
Average Distance (km)	0.3	0.6	0.9	1.2	1.5	1.8	2.1
Horizontal Distance (Km)	0.2	0.2	0.2	0.2	0.2	0.2	0.2
Vertical Distance (Km)	0.1	0.4	0.7	1.0	1.3	1.6	1.9
Horizontal Velocity Empty (km/h)	8	8	8	8	8	8	8
Vertical Velocity Empty (km/h)	10	10	10	10	10	10	10
Horizontal Velocity Loaded (km/h)	8	8	8	8	8	8	8
Vertical Velocity Loaded (km/h)	6	6	6	6	6	6	6
Capacity (t)	40	40	40	40	40	40	40
Fill Factor	0.95	0.95	0.95	0.95	0.95	0.95	0.95
Useful volume of cargo equipment (m3)	14.0	14.0	14.0	14.0	14.0	14.0	14.0
Buckets quantities	3.0	3.0	3.0	3.0	3.0	3.0	3.0
CYCLE

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Time of loading (min)	25.0	22.3	23.4	25.4	22.8	21.2	21.7
Dump Time (min)	1.8	1.8	1.8	1.8	1.8	1.8	1.8
Travel time Load (min)	3.0	6.0	8.4	11.4	14.4	17.4	20.4
Travel time empty (min)	2.4	4.2	5.4	7.2	9.0	10.8	12.6
Turnaround time (min)	0.8	0.8	0.8	0.8	0.8	0.8	0.8
Utilization Rate (%)	0.80	0.80	0.80	0.80	0.80	0.80	0.80
Cycle time (hours)	0.69	0.73	0.83	0.97	1.02	1.08	1.19

Source: Optimize, 2022

	Road Grade (%)	Speed (km/h)
Loaded	0%	8.0
	-15%	6.0
Empty	0%	8.0
	15%	10.0

Source: Nordmin, 2019

The ore haulage distances were evaluated from the mine design. Based on this evaluation, ore haulage routes were measured according to the distance from the LHD and truck to the loading area to dump, per level. Microsoft Excel^® was then used to generate a one-way ore haulage distance and an average cycle time (round trip) using the parameters shown in Table 13-8 to Table 13-10.

The average one-way ore haulage distances are approximately 345 m in the first eight years of the LOM; and increases to approximately 1,378 m from years 9-16; and then, 1,015 m for the remainder of the LOM; the LOM average is 950 m. At the peak, five haul trucks are required to transport the ore and waste. Figure 13-12 and Figure 13-13 show the haulage distance and cycle time by monthly and yearly time periods. The cycle times reflected in this summary are indicative as there is a fixed component including the loading time, dumping time, positioning time, and additional delays that are included in the productivity determinations was included in the information summarized in these figures.

 

Source: Optimize, 2022

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Figure 13-12: Haulage Distance — One-Way Length

 

Source: Optimize, 2022

Figure 13-13: Haulage Cycle Time – Roundtrip

During the pre-production period, before mining of stopes and the commissioning of the plant, waste and mineralized material is hoisted to the surface and stored separately in a designated lined storage facility. During the pre-production period, 33 first months of LOM, the mine produces approximately 1,131,026 tonnes of waste and 488,034 tonnes of ore that is stockpiled until the processing plant is commissioned and ramps up to full production.

13.5.4  Backfilling

The mine production sequence includes the use of cemented paste backfill to fill the voids left by the stopes to maintain the mine structural integrity. The mine utilizes a high strength backfill paste that has a 2% cement only content in the primary stopes. For secondary stopes, lower strength paste with a lower cement content and a possible fly ash blend could be used to supplement development waste rock, whenever development waste rock is not available to backfill stopes.

Section 15.7 discusses the paste backfill surface plant and system to move the paste backfill underground to the stopes. A backfill operations crew installs barricades in the lower access drift to the stopes, extends the pipe delivery system from the production shaft via the upper access drift into the stopes, and monitors the backfill as the stope fills. Once the stope is filled the backfill is allowed to cure (28-days) to design strength of over 1 MPa before blasting on the adjoining stope.

During the LOM, 2% cemented paste is used in the primary stopes, while a combination of lower cement content and a possible fly ash blend paste backfill and rockfill (up to a maximum of 50%) is used in the secondary stopes. If the secondary stope is the last one to be mined, then it may be filled with up to 100% rockfill if sufficient material is available.

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In the mine development phase, waste rock is hoisted to surface and store until the stopes become available for backfilling. During the LOM, only 730,474 m³ of rockfill is generated from waste development headings and is available for backfilling in secondary stopes. Therefore 11,643,100 m³ of paste backfill is required.

Table 13-11 provides the LOM backfill breakdown by volume and type. Figure 13-14 shows the annual backfill production requirements over the LOM.

Table 13-11: Backfill Volume Summary - By Type

Backfill Type	Volume (m3)
High Strength Backfill (2% Cement only Paste backfill)	6,059,071
Low Strength Backfill (Lower cement content with possible fly ash blend paste backfill)	5,584,029
Rockfill	730,474
Total Backfill	12,373,574

Source: Optimize, 2022

 

Source: Optimize, 2022

Figure 13-14: Backfill Production during LOM.

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13.5.5  Ground Support

The lateral development assumptions include systematic bolting with screen for all areas. Some specific excavations include shotcrete either because they have been identified as weaker ground conditions or for long term stability requirements. As there are faults and discrete structures that may impact stability locally, the unit cost of the various type of development has been increased to consider that a portion of the development will require shotcrete locally. Table 13-12 shows the estimated amount of the total development that is projected to require extra support, whether it is shotcrete or cable bolting. The extra support accounts for intersections with wider span or geological structures that would require special attention. It is assumed that all the stope brows will need extra support, which attributes for 25% of the x-cuts requiring extra support.

Table 13-12: Extra Support Assumptions by Heading Type 

Heading Type	% Extra Support
Ramp	15%
Electric Substation	100%
FAR Access	20%
FW Drift	15%
Ore/Waste Pass Access	100%
X-Cuts	25%
Remucks	10%
Sump	100%
Refuge Station	100%

Source: Nordmin, 2019

13.5.6  Grade Control and Reconciliation

The objective of an underground grade control program, as part of a routine mining sequence, is to maximize the value of ore mined and fed to the surface plant. The grade control (or ore control) process involves the predictive delineation of the tonnes and grade of ore that will be recovered by mining. The predictions have several common characteristics across all mineralization and mining types, for instance, from small, low production rate, metalliferous underground mines to large world-class open pits.

Accordingly, a dedicated grade control sampling practice must ensure the following:

● The program aims to deliver the most economic tonnes to the mill via an accurate definition of “ore” and waste;

● The program aims to identify variations in the dip, strike and width, impact on a local scale from faulting effects, and grade continuity/type. Variations in geometry at the edge of the mineralization require a geological understanding to ensure optimum grade, minimal dilution and maximum mining recovery;

● Safe practices are followed during the sampling process;

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● Sampling remains as unbiased as possible;

● Representative (i.e. correct in terms of Gy’s sampling theory (“Wikipedia Contributors,” 2019); and

● Timely (so that the results can usefully define the ore blocks).

A successful program in an underground environment is completed through detailed geological mapping and grade sampling ahead of the mining. The mine geologist is to perform daily mapping and define the ore/waste contact for the mining team to progress. The mapping is incorporated into a digital format to improve the geological model further and enable the development of short-term estimation. The grade control strategy is related to the mining method and orebody type. For underground operations sampling methods include chip, channel and panel samples, grab/muck pile samples, and drill-based samples.

NioCorp will establish a daily grade control strategy as outlined above before commencing mining. The strategy will involve an infill drilling and underground chip sampling program. The goal of the infill drilling and underground chip sampling program is to monitor and provide close spaced sampling that is required to define the boundaries of mineable ore blocks. The amount of sampling is constrained by practical limitations and cost considerations and will be adjusted over the LOM based upon the needs of the mining operation.

The infill drilling will occur from established drilling stations on each level, with underground diamond drilling using an NQ core diameter drilled across the width of the known mineralization. Drill logging is required to collect geological and structural measures in conjunction with assaying. The current protocols and procedures developed by NioCorp for exploration work require further development to support a daily production environment. NioCorp will drill multiple holes in a fan pattern from each station to gain information for levels above and below as required. The current geometry of the orebody supports completion of the infill drilling in advance of the mining to enable the design of the ore blocks to be based on true grade control sampling. Accordingly, Figure 13-15 outlines many of the critical areas that require further infill definition drilling.

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Source: Nordmin, 2019

Figure 13-15: The Critical Areas Requiring Further Infill Definition Drilling

Additionally, NioCorp will rely on underground chip sampling to provide infill sampling for grade control purposes. Underground chip sampling continues to depend primarily on manual methods of extraction, i.e., collecting rock chips using a hammer and chisel. The proposed mining method will involve the development of cross-cuts at regular intervals across the width of the mineralization at the top and bottom of a stope before mining. Samples will be taken across the full width of the exposed mineralization with sufficient volume to ensure accurate assay. The sample weight will be the equivalent of at minimum half NQ core for the sampling interval. The samples will be logged geologically marking the width of the mineralization and any hanging wall or footwall mineralization.

Additionally, blasted material is available for grab sampling to test grades, which will be input into a production database and be used to confirm head grades and used for reconciliation purposes.

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The close spaced sampling collected from drilling, chip sampling and grab sampling are utilized within the mine reconciliation process, where mine reconciliation is completed progressively by accumulating predictions for treated ore and comparing them to the production and mill results.

NioCorp will establish a consistent reconciliation framework, monitored on a daily/weekly/monthly and yearly process. The essential steps of the framework include the following:

● Establish an audit trail for all data.

● Agree to report results routinely in a consistent format and ensure that there are cross-functional reconciliation meetings in place to discuss results and develop action plans.

● Tabulate the data.

● Report variations based on consistent volumes (bench by bench, stope by stope) or periods (monthly, quarterly, annually).

● Graph the variations (or factors) for each parameter to determine trends.

● Analyze the differences and annotate the graphs to explain the differences.

● Alter the input parameters systematically to reduce future reconciliation differences.

A consistent framework of establishing reconciliation has been established by Harry Parker (2012) that used the following definitions:

F1=	short range model depletions	i.e. F1=	GRADE CONTROL (PREDICTION)
	long range model depletions		ORE RESERVE (PREDICTION)

and

F2=	received at mill	i.e. F2=	MILL (PRODUCTION)
	delivered to mill		GRADE CONTROL (PREDICTION)

and

F3=	received at mill	i.e. F3=	MILL (PRODUCTION)
	long range model depletions		ORE RESERVE (PREDICTION)

then it is now evident that F3 = F1 * F2

Source: Shaw, W.J, et al., 2013

By ensuring that reconciliation calculations are all done as factors (for tonnes, grade and metal), and each stage of the chain is used as a numerator when compared to the previous component in the chain, all of the various components of a mine reconciliation scheme can be rationalized and compared (see Figure 13-16).

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Source: Shaw, W.J, et al., 2013

Figure 13-16: Elaboration of the Reconciliation Process Defining Additional Steps

A series of protocols covering all grade control tasks, reconciliation from mapping to sampling, and integration with the database shall be implemented at the mine operation. This includes an ongoing review of the quality assurance/quality control monitoring program to ensure protocols and staff are updated as required.

13.5.7  Workforce

Workforce levels are estimated based on the production schedule and equipment needs. The productivities used reflect a mix of local and skilled labour with an experienced management team.

The estimate is based on the utilization of a contractor for mining development and operations with an ownership senior management team to oversee mining activities. The rotating contractor crews will be using an operating schedule consisting of 12 hours per shift, two shifts per day, and seven days per week. A four-crew arrangement supports the 12-hour shift with two crews onsite at any given time (per rotation). The ownership, senior management and technical team are planned to work five 8-hour days per week.

Table 13-13 shows the maximum required workforce. There are 96 people on a two-week rotation and 24 ownership senior management and technical team on a weekly basis. The workforce increases over time to a maximum of 216 in year five. There will be a maximum of 120 people onsite at any given time.

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Table 13-13: Typical Mining Labour 

Management / Technical Support		Total Qty.*
Mining Manager		1
Mine Superintendent		1
Maintenance Superintendent		2
Chief Engineer		1
Geotechnical Engineer		1
Long Term Mine Planner		1
Short Term Mine Planner		1
Project Engineer (ventilation, water, construction)		2
Chief Geologist		1
Resource Geologist		1
Grade Control Geologist		2
Administrative / Mine Clerks		1
Chief Surveyor		1
Mine Surveyor		3
Material Handling / Shaft Shift Supervisor		2
Mine Services Shift Supervisor - Construction		1
Maintenance Shift Supervisor - Fixed Equipment		1
Electrical General Foreman		1
Total Management / Technical Support		24
Rotating Crews	Per Rotation Qty.*	Total Qty.*
Shaft Services	2	4
Hoistperson	2	4
Deckman	2	4
Skip Tender / Crusher	2	4
Safety Technician / Trainer	2	4
Development / Production Shift Supervisor	2	4
Vertical Development Crew	2	4
Blasting/Powder Crew	4	8
Blasting/Powder Crew Helper	4	8
Jumbo Operator	4	8
Longhole Drill Operator	3	6
LHD Operator	7	14
Haul Truck Operator	8	16
Bolter Operator	8	16
Cable Bolter Operator	2	4
Nipper	4	8
Shift Supervisor - Logistics	1	2
Utility / Construction Crew	4	8
Grouting Lead	1	2

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Grader Operator	1	2
Conveyor Attendant	2	4
Diamond Driller	4	8
Maintenance Supervisor - Mobile Fleet	2	4
Mine Electrician	7	14
Heavy Equipment Mechanic	12	24
Welder	2	4
Instrumentation Technician	2	4
Total Rotating Crews	96	192
Grand Total		216

Source: Nordmin, 2019 

*This value represents peak contractor and ownership workforce.

13.5.8  Equipment

The underground equipment used, shown in Table 13-14, is typical for a sublevel stoping mining method with the number of pieces of equipment calculated from the production rates and typical availabilities for equipment in underground mines.

The estimate uses typical availabilities and utilization rates for mining equipment used for this mining method. Each shift of 12 hours is reduced by 2.25 hours to represent shift change, breaks, lunch, fuel/grease/inspection time and travel to and from work areas. This provides an equivalent working day of 19.5 hours or 9.75 hours per shift. The resulting reductions result in 5,931 productive hours per year of mining time. It should be noted that the layout of this mine and mining on multiple levels requires the addition of equipment to reduce equipment move time. This reduces the overall utilization of the equipment fleet.

Table 13-14 summarizes the mine equipment totals for peak production.

Table 13-15 summarizes the major fixed equipment for the mine.

Table 13-14: Mine Mobile Equipment 

Type of Equipment	Quantity*
Drill Jumbo	3
Haul Truck (40 t)	5
LHD (6.2 m3)	4
Longhole Drill	2
Cable Bolter	1
Bolter	4
Grader	1
Personnel Carrier	2
Pick-up Trucks	6
Utility Vehicle	3
Boom Truck	2
Scissor Lift	2

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Shotcrete Sprayer	2
Anfo / Emulsion Loader	3
Development Emulsion Loader	1
Production Emulsion Loader	1
Portable Grout Unit	2
Blockholer	1
Exploration Drill	2

Source: Nordmin, 2019 

* This value represents peak production mine mobile equipment fleet.

Table 13-15: Mine Fixed Equipment 

Type of Equipment	Quantity*
Production Shaft Skip Hoist	1
Production Shaft Service Cage Hoist	1
Production Shaft Auxiliary Cage Hoist	1
Ventilation Shaft Auxiliary Cage Hoist	1
Mine Ventilation Supply Fans (65 kW)	2
Mine Ventilation Exhaust Fans (790 kW)	2
Auxiliary Ventilation Fans (112 kW)	6
Auxiliary Ventilation Fans (56 kW)	4
Natural Gas Mine Air Heaters	2
Surface Apron Feeders	2
Surface Conveyors	3
Service Cage	1
Auxiliary Cages	2
Skips	2
Rock Breakers	4
Grizzly Feeder	1
Apron Feeder	1
Jaw Crusher	1
Crusher Discharge Conveyor	1
Ore Bin Conveyor	1
Loading Pocket Conveyor	1
Vibratory Belt Feeder	2
Belt Magnets	3
Chain Gates	5
Crusher O/H Crane	1

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Electric Battery Stations	3
Underground Shop O/H Cranes	3
Main Dewatering Pumps	6
Primer Pumps	2
Submersible Pumps	10
Main Air Compressors	2

Source: Nordmin, 2019 

* This value represents peak production mine fixed equipment.

13.6 Ventilation

The ventilation system design was aligned with the mine design and production schedule described in previous sections. The backbone of the design includes the main production shaft as a fresh air intake airway, and the ventilation shaft as an exhaust airway. Both have a finished internal diameter of 6.0 m. Air is circulated through the two shafts each having a set of two parallel surface fans, which are connected to a plenum arrangement that in turn is connected to the shaft collar area. The two intake fans are low pressure and solely responsible for delivering air through the mine heaters and into the intake production shaft at a slightly higher volume than the mine requires, the excess ventilates the headframe building. The two exhaust fans are responsible for drawing air through the whole mine. Under this pull arrangement the intake fans would each have a power rating in the order of 75 kW and the exhaust fans would each have a power rating in the order of 800 kW.

The volume of air travelling through the mine is currently identified as 283 m³/s. When surface temperatures are lower than 4° C, heaters in front of the two intake surface fans are activated.

13.6.1  Airflow Requirements

The airflow requirements were primarily based on the engine rating for the diesel-powered mobile equipment planned to be operated, within the entire mine. The fresh air will then be distributed to meet demand based on this criterion within any given segment of the mine. Whereas the mucking and haulage fleet will be diesel powered, the drill carriers and utility vehicles will be electrically powered by batteries. A factor of 0.063 m³/s per kW of engine power has been used to determine the air quantity required for adequate dilution and dissipation of diesel engine emissions. Minimum airflow requirements for the mine were estimated based on the mobile mining equipment list and estimated utilization. The design airflow of 283 m³/s was deemed sufficient for the mining activity at maximum productivity. The validity of this air quantity with respect to currently unaddressed factors and their influence on fan and heater duties will need to be addressed at the next stage of the study. These further design considerations include adversely hot conditions underground in the summer, the potential need to manage harmful radio-active elements (radon and its progeny), and air velocity in the shafts.

The mine-wide airflow is directed, via airlock doors, regulators and auxiliary ventilation fans connected to ducting, to the areas where it is required based on specific equipment needs, and where it is needed to maintain a minimum airflow in working areas in the absence of diesel equipment. This assumes the use of a ventilation management system (ventilation on demand) to distribute the air in line with mining activities to maximize the air utilization and ensure system efficiency.

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The introduction of ore and waste passes has resulted in fewer haulage trucks being operated in the mine, which in turn has lowered the air quantity requirement. An opportunity may exist to make further reductions with the introduction of battery-powered electric mucking and haulage equipment as the mine progresses beyond the initial development phase. Alternatively, the use of cleaner diesel engine technology, to reduce diesel particulate and gaseous emissions may also permit a reduction in air quantities. However, all potential reductions will need further review to ensure other ventilation design needs are covered.

Air velocity limitations vary according to airway type and their activities. The minimum recommended air velocity in a drift, for perceptible air movement, for workers is 0.3 m/s. In areas such as return airways and shafts where personnel are not expected to work, higher velocities are acceptable. Table 13-16 contains relevant selections from a table on page 9-13 of Subsurface Ventilation Engineering by Malcolm J. McPherson (1993). It provides the threshold airflow velocities for various airway types. These maximum velocities were considered for shaft and development drift profiles.

McPherson also recommends upcast shafts avoid velocities in the range of 7 to 12 m/s, to avoid water droplet stagnation which could impact fan performance. This will need further consideration.

The present airflow allocation has not considered the consequences to ventilate for radio-nuclides if present, and heat management requirements during the summer months when surface temperatures can exceed 30C. The decay of thorium and uranium can produce radon daughters that have specific ventilation requirements.

Table 13-16: Maximum Airflow Velocities (m/s) 

Area	Velocity (m/s)
Working Faces	4
Conveyor Drifts	5
Main Haulage Routes	6
Shafts with Hoisting	10

Source: Nordmin, 2019

13.6.2  Ventilation Controls

Fixed Facilities Controls

Figure 13-17 demonstrates the suggested controls for the garage and shaft bottom area. Air from the garage flows directly to a return air raise that ventilates directly into the ventilation/exhaust shaft. The current allowance for such fixed facilities, and their accommodation in the overall mine airflow requirement will need to be reviewed at the next stage of the study.

In the event of a fire in the workshop/garage area, the controls prevent smoke and fumes from contaminating the mine. Remotely activated fire doors will be installed in all fuel bays and the garage to prevent the spread of fumes.

Interlocked equipment doors with adjustable regulators/louvers forming the airflow distribution management system are used to control the quantity of air delivered to various levels of the mine.

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Level Ventilation Controls

Air enters the mining levels from the fresh air raise system, the volume is regulated depending on the air requirement at that stage of production. Portions of the air travelling within the flow-through route on a level are picked up and used in production headings via auxiliary ventilation fans and ventilation ducting. Once used in a heading, the air then re-enters back to the flow-through route to be carried to the exhaust raise system and into the ventilation/exhaust shaft. Enough air is pulled from the intake raise system a allow a small quantity of fresh air to flow out to the ramp. This system allows mining on more than one level, without contaminating downstream work areas with upstream diesel exhaust and dust (see Figure 13-18). Further review of the airflow being supplied to the ramp will be necessary considering vehicle type and the distribution of activity over mining levels.

The construction of temporary bulkheads is required when stopes are open between levels, to maintain control of the ventilation system and prevent short-circuiting of air. These temporary bulkheads can be a simple curtain type bulkhead made from brattice material.

 

Source: Nordmin, 2019

Figure 13-17: Garage Area Ventilation Controls (Looking Northwest)

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Source: Nordmin, 2019

Figure 13-18: Level Ventilation Controls (Plan View)

When mining of the level is complete, the regulators are sealed off with door access bulkheads, which will allow re-entry for inspection if required. These bulkheads will prevent short-circuiting of air into these mined out levels.

13.6.3  Ventilation Model

Ventsim Design simulations were used to generate the ventilation model using pre-set airway resistance factors programmed into the software. Several staged ventilation models were generated to simulate the significant phases of the mining stages.

Staged Modelling Fan Results - Main Fans

Table 13-17 provides the main surface fan duties. The installed powers were calculated with an assumed fan efficiency of 70% and the required air power. The assumed air density at the fans is 1.18 kg/m³.

Table 13-17: Duties of Main Surface Fans 

Description	Pressure (kPa)	Quantity (m3/s)	Air Power (kW)	Motor Power (kW)1	Total Motor Power (kW)1
2 Intake Fans (in parallel, per fan)	0.3	300	45	65	130
2 Exhaust Fans (in parallel, per fan)	3.90	283	550	790	1580

Source: BBE QP review 

(1) Assume 70% fan efficiency

The resulting operating pressures should be considered “applied pressures”, that is, these pressures may not fully account for losses associated with fan housings, ducts, plenums, or diffusers. The required pressures will be the subject of further review at the next stage of study.

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Staged Modelling Fan Results - Surface Heating Fans

The surface intake fans are low-pressure, high-volume fan(s), as required for the air heaters. These fans produce just enough pressure to overcome the losses from the heater, fan, and plenum so that there is a slight positive pressure within the production headframe. The total airflow through the intake fans is slightly greater than through the underground workings because additional air is used to slightly upcast air in the main production shaft in order establish a positive pressure in the production headframe.

Table 13-18 provides an indication of the total airflow distribution to the main mining areas. These will be subject to further review with respect to heat management and radon exposure requirements, if applicable, facility allocations, leakage allowances and ramp vehicle movement requirements.

Table 13-18: Total Mine Airflow 

Mining Area	Airflow (m3/s)
Upper Mining Area	142
Lower Mining Area	142
Total	283

Source: Nordmin, 2019

13.6.4  Auxiliary Ventilation

In areas that are not in the path of flow-through ventilation, including the production area to the west of the fresh air raises as well as the stope access crosscuts, auxiliary ventilation is used. The typical auxiliary system consists of a fan and attached ducting, that is run to the mining face, draw point or drill drift. Fans and ducting are selected to deliver enough air to provide 0.063 m³/s for every 1 kW of the maximum diesel-powered equipment that will operate simultaneously in each such area.

13.6.5  Recommended Ventilation Infrastructure

Sensors

Several different types of remote sensors are recommended for controlling and managing the operation of the ventilation system within the mine. These sensors can help predict fan condition, alarm in the event of a fire, indicate low or high temperatures, or detect harmful gasses. These can be tied into the ventilation management system and the modelling software. Bundled air quality and quantity sensors are recommended for each fan installation, intake shaft, fixed facilities and each working level. These include fan monitoring, air quality, air quantity, and psychrometric sensors.

Regulators

Drop board regulators consist of large wooden or steel boards which are slotted from the ground up. Placing more boards results in a smaller opening and consequent generation of higher airway resistance and less airflow. Regulators can also form the frame for a bulkhead. Drop board type of regulators are recommended to be placed at the intake and exhaust of each level.

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Slide regulators are typically a piece of steel on a slider that can be adjusted manually to increase/decrease the area of an opening in a door or a regulator.

Louvred regulators are manufactured items that consist of steel slats that rotate on a horizontal axis within a frame and are controlled electronically. These regulators can be controlled from the surface and can form part of a ventilation on demand (VOD) system.

The level of ventilation distribution management required, and the variability of specific flows will ultimately dictate the type of regulator needed at each location

Bulkheads

Temporary bulkheads for stopes with an open brow are constructed from flexible plastic PVC brattice material. They do not experience high pressures and serve to prevent loss of ventilating air from stopes while in an LHD mucking phase.

Permanent bulkheads such as for a level no longer in use consist of a shotcrete wall with a steel personnel door, to allow worker access.

Equipment Doors

Pneumatically operated, steel equipment doors allow passage of vehicles and materials, and control the volume of airflow to or from an area. Dual air locking doors in series are used where the pressure is highest. Personnel doors are installed beside the main equipment doors to safeguard workers against being struck by a closing door.

Air Heaters

Natural gas air heaters are used at the surface intake shaft. Low-pressure high-quantity fans are used with the air heaters. These fans produce just enough pressure to overcome the losses from the heater, fan, and plenum so that there is a slight positive pressure in the production headframe.

13.6.6  Ventilation Power Consumption

Based on the fan operating points, the total motor power of the main exhaust fans is estimated to be 1,580 kW. The monthly power consumption averages to 1,153,400 kWh/month.

13.6.7  Air Heating

Winter ice buildup can cause airways to be restricted and lead to hazardous conditions. An intake air heating system, as described below, is recommended that will mitigate these conditions. Based on the average temperatures shown in Table 13-19, air heating is required for seven months out of the year. A 7.9 MW heater provides 74,930,000 MJ per year, based on average temperatures and an airflow of 300 m³/s. Instrumentation for this heater includes a thermostat in mixed air an appropriate distance downstream of the heater, as well as a carbon monoxide monitor with an alarm and automatic fuel cut-off to the heater. A higher duty heater would need to be considered to cover daily extremes.

Table 13-19: Surface Temperatures Near the Elk Creek Mine 

	Jan	Feb	Mar	Apr	Oct	Nov	Dec
AVG °C	-10.6	-8.3	-2.2	3.9	3.9	-2.2	-8.9

Source: Nordmin, 2019

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13.6.8  Thermal Exposure

Summer surface air temperatures can reach in excess of 30°C. The temperature of the air descending the intake shaft will increase though auto compression. Machinery and the strata can be additional inputs to the final working conditions at the work face. To maintain a suitable working environment, it may be necessary to provide additional air or cooling to avoid heat stress. More detailed climatic modelling will be required at the next stage of study.

13.7 Mine Infrastructure & Services

13.7.1  Material Handling System

The underground material handling system is designed for both waste and ore, to provide surge and storage capacity underground, to size ore, and to be an efficient, automated system from underground mining areas to the mineral processing plant on the surface via the production shaft and surface conveyors.

During underground operations, mined waste and ore will be dumped through a typical 300 mm x 300 mm (12” x 12”) scalping grizzly complete with rock breaker and will be stored in either a waste pass or either of two ore passes. The three passes will serve the majority of mining levels, (-215 El. and above) and will lead to either a waste or an ore storage bin. The ore then passes over an ore handling apron feeder which will feed the ore sizing and storage circuit. The waste from the storage bin feeds onto the loading pocket conveyor.

The apron feeder will supply ore from the ore pass to a grizzly feeder, which will allow undersized material to be removed from the crushing circuit prior to the crusher. The oversized material will continue to the single C series jaw crusher, which will size the ore to the mineral processing plant required 115 mm (4½”). All ore will then be passed through 48” belt width conveyor systems and transfer cars to either the dual use waste or ore storage bin or the single-use ore storage bin. Each of the underground storage bins are designed to hold up to 2,500 tonnes (2,756 tons).

Ore or waste from the loading pocket conveyor is forwarded via conveyor and transfer car to the twin 11 tonnes (12 tons) weighted flasks at the periphery of the production shaft. The flasks are weighed, and the feed and discharge system is controlled via a PLC/PC interface. Control monitoring around the facility will use CCTV cameras located at strategic points, load sensors, bearing and motor monitors, etc. Rock breakers will have the option to be remotely operated by the hoist operators via joystick and camera, or by an operator located underground at the rock breaker. The entire system will be monitored by the hoisting personnel.

Once each flask is adequately filled and the corresponding skip is in place (determined by skip load cells), arc gates will release ore or waste to the skip within the shaft. Loaded skips will be hoisted to the headframe on the surface and dumped into a chute which will direct the material to either a 1,000 tonne (1,100 tons) waste storage bin or to a 2,000 tonne (2,200 tons) ore storage bin via a transfer car. The waste will be stored in the bin and further loaded into haul trucks to be brought to the dedicated waste storage area. The stored ore will pass onto a conveyor system, complete with manual loading area, and will report to the mineral processing plant.

13.7.2  Mine Dewatering System

Mine dewatering at the Project is designed to accommodate groundwater inflows from the shaft and mine workings, along with inflows from drills and other underground operating equipment. Life of mine average mine inflow is estimated to be approximately 32 L/s (500 US gpm), with a peak mine inflow limited by grouting to 63 L/s (1,000 US gpm). Total mine dewatering flows

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(which include mine inflow plus a relatively small amount of service water), can generally be accommodated with one main dewatering station pump in operation. To accommodate for peak mine inflows (expected during the early mine life), the design capacity of the underground dewatering system can accommodate an elevated flow rate of 63 L/s (1,000 US gpm), per Section 7.4.2 with two station pumps in operation, and a maximum flow rate up to 95 L/s (1,500 US gpm) with all three station pumps in operation.

The system design incorporates two main pumping stations, one on the -15.4 m (-50.5 ft) level and one on the -335.4 m (-1,100 ft) level, that work in series to lift mine water from the lowest depths of the mine to the surface via the production shaft. Each station is comprised of a single 150 m³ (5,297 ft³) conical settling sump (borehole) allowing clear water to pass along to three positive displacement (PD) pumps (GEHO ZPM800) which lift the water from the station to the next available level through 8” pipes within the shaft. Additionally, during upset conditions and when required, the PD pumps will allow operations to lift dirty water from the settling sumps up through the shaft piping to surface. Figure 13-19 demonstrates a typical pumping system.

 

Source: Nordmin, 2019

Figure 13-19: Typical Pumping System (-15.4m and -335.4 m)

The cone bottom settling sumps will also integrate a flush water port to be utilized to assist draining of settled solids to a desired location on the pumping level. The shaft will house three 8”

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diameter carbon steel schedule 80 pipe columns. Water from the underground workings will be forwarded to the water treatment center on the surface, where it will be purified to supply the hydromet and pyromet and underground with process water and offices, mine dry and other buildings with potable water.

13.7.3  Compressed Air System

Compressed air will be supplied to the production surface mining structures (headframe and hoist house) via twin compressors located within the production shaft hoist house. Each compressor will be rated for 0.94 m³/s (2,000 scfm) at the intake with a discharge pressure of up to 862 kPa (125 psig), which will ensure that the compressed air delivered underground remains above 690 kPa (100 psig). An ASME Section VIII air receiver (complete with water purge valve and safety relief valve) located within the hoist house will also ensure that the compressors are not required to cycle on and off excessively. Surface branch connections will include the hoist house and headframe where a requirement for air driven hand tools exits. No requirement for instrument air, requiring a drying system exists within the production hoist house.

An 8” diameter carbon steel pipe will deliver the compressed air from the hoist house to the headframe via an underground services trench. This pipeline will then proceed to the production shaft, where it will form part of the services hung from the shaft steel, down to the working levels, including crusher station, conveyance level, loading pocket and spill pocket. Each level will have an individual line branched from the shaft line, that will service the level. No additional underground storage will be required, as all main lines serving the levels will be large enough to accommodate the usage and will supply additional storage.

The requirement for compressed air within the backfill plant and the ventilation shaft and headframe will be filled with small portable compressors housed within the structures.

13.7.4  Underground Water Supply

Industrial or process water supply for drilling and dust control will be supplied from the water treatment center via a 4” nominal carbon steel line to the production headframe and down the production shaft. Individual branch connections will report to each of the levels that require process water, including those that require water for the dust suppression system. As the static pressure increases, the deeper the line extends (to the lower shaft stations), pressure reducing stations will be utilized.

Pressure reducing stations are strategically located down the shaft, at the crushing station and loading pocket levels. Typically, a 4” diameter Pressure Reducing Valve (PRV) with either flanged or grooved ends will be employed at each station. The PRVs are used to reduce supply pressures from a maximum of 2758 kPa (400 psig) to the desired discharge service pressure. The PRV stations will be located as close as possible to the shaft.

13.7.5  Underground Fuel Storage and Distribution

Fuel from the surface storage facility will be delivered to the underground storage system via a 2” diameter fuel transfer pipeline within the production shaft. The fuel line will run from the surface, down to the underground shops level where the line will be routed to a storage area at a fuel bay for fueling vehicles. The fuel pipe feeds fuel to either of two 3,785 L (1,000 US gallons) storage and distribution systems, located within a cut out on the south of fuel bay, via a motorized three-way valve.

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Each storage and distribution station will be a bladder type, with up to 150% containment, complete with the following safety functions: 4 hr rated UL approved roll-up door, thermally activated fuel shut off valve to the dispensing system, anti-syphon valve, and a dry chemical automatic fire suppression system with detection and actuation. Each station will be individually alarmed, by means of a PLC with level alarms, and a level switch.

Additionally, fusible link fire doors are also included in the underground layout, these twin fire doors, upon actuation, will isolate the fueling area from the main shops.

13.7.6  Workshop, Maintenance Bays, and Warehouse

The maintenance area consists of nine large bays of approximately 16 m (52’) long by 7 m (23’) wide to accommodate vehicular traffic. One wash bay is included in the workshop layout. A drainage trench with covering grating runs the length of the bay to carry water to a nearby oil capture sump. Grading of the area will reduce the possibility of oil contamination. Three maintenance bays are equipped with an overhead crane to facilitate the maintenance work on vehicles.

Warehouse and tool cribs are included within the maintenance area.

Airlock doors separate the maintenance area from the rest of the mine. An office is located at the end of a drift located in the maintenance area.

13.7.7  Explosives Storage

The mine design includes underground powder and primer magazines. The mine explosives are stored off-site at a vendor location and deliveries are on as needed basis with the underground magazines providing the capacity required for production needs. The explosives pricing includes the contractor storage and supply totes, as per the manufacturer’s recommendation, all of which are included in the capital estimate.

13.7.8  Refuge Stations/Chambers

Two mobile refuge chambers have been included within the underground mine design. Each refuge chamber will be sufficiently equipped to house 12 or more persons, depending on location and unit size, for up to 36 hours. The stations are self-sufficient in that they include seating, a chemical toilet, emergency food and water, back-up power, lighting, and communications via external antenna and 12V power supply. The breathable air system that is incorporated within the refuge chambers includes a standard compressed air line tie in, oxygen cylinders connection, as well as an oxygen candle. Each chamber can be located at the most strategic location as dictated by the mining operation and underground workings. The chambers are easily transported by forklifts or LHD units.

In addition to the two mobile refuge stations, there will be a permanent refuge station located on the 530 Level and 650 Level. Both permanent refuge stations will be equipped in a similar fashion to the two mobile refuge stations, but with a capacity of 30 persons per station.

13.7.9  Hoist House Substation Surface Electrical Distribution

Electrical power will be supplied to the hoist house substation via a 44 kV overhead line from the main substation. Power will be stepped down to 13.8 kV by two 20/25 MVA Delta/Wye resistance grounded transformers, supplying a main-tie-main primary distribution switchgear.

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The E-House in the hoist house substation will distribute power at 13.8 kV to the production hoist and service hoist drive transformers, one underground feeder (3C 4/0 AWG), the hoist house 4160 V power distribution (via two 13.8 kV-4160 V step down transformers) and an overhead line. The overhead line will supply power to the surface infrastructure (offices, administration, dry) and ventilation shaft power distribution switchgear. The second underground circuit is fed from this ventilation shaft switchgear unit.

Two backup diesel generator units are included in the power system for use during a utility outage. The first backup generator is connected to the hoist house switchgear and includes a regenerative load bank. The second backup generator is connected to the ventilation shaft switchgear, at 13.8 kV, and includes a regenerative load bank.

Critical loads will include both auxiliary hoists, auxiliary hoist motor control centers (MCCs), downcast fan #1, and the underground mine feeder #1. The underground feeder #1 will provide power to critical underground ventilation during a utility outage.

Additional loads on the 4160 V hoist house switchgear include downcast fan #2, compressors and 480 V services.

The 480 V unit substation will supply power to all the 480 V MCCs which will supply power to all other electrical loads in the hoist house, headframe/collar house, hoist house substation and downcast fan heater building. The hoist house MCC will service 480 V loads, ventilation, lighting, and low voltage services for the hoist house. The production hoist MCC will supply all loads to all ancillary equipment for the operation of the production hoist. The headframe/collar house MCC will be in the collar house and will service all 480 V loads in the headframe/collar house, including ventilation, lighting loads, and skip handling equipment.

13.7.10  Underground Electrical Distribution

Two 13.8 kV shaft feeders will supply power underground. Underground feeder #1 will originate in the hoist house substation, traverse through the hoist house, to the headframe. The cable will then be hung vertically in the production shaft with the cable turning into selected levels to distribute power. Underground feeder #2 will originate at the ventilation shaft switchgear and will be hung vertically in the ventilation shaft.

At each supplied level, there will be a dual load break switch to select which feeder is used to supply power to the level. A 15 kV junction box will be used to provide a junction point to supply the necessary equipment and provide a point for expansion at each level.

13.7.11  Overhead Pole Line Electrical Distribution

A 13.8 kV overhead pole line requires construction to distribute power to temporary loads during shaft freezing and sinking. The overhead line will originate at a temporary generator farm, which will be used to supply power during the initial stages of construction. The line will remain in service after temporary power has been removed and will be permanently connected to the 13.8 kV main substation. This overhead distribution will supply power to the ventilation shaft switchgear unit, and the office/administration/dry structure.

13.7.12  Hoisting Plants

The hoisting plants are designed to serve as both an efficient means of hoisting ore and waste to the surface and of lowering and lifting labour forces, materials, and equipment between the surface and the underground working levels of the mine.

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The production shaft hoisting plant (see Figure 13-20) is comprised of three hoists:

● the service cage double drum (DD) single-clutched hoist in a balanced condition,

● the skipping DD single-clutched hoist in a balanced condition, and

● the auxiliary cage single drum (SD) hoist.

The hoist duty calculations conducted in the 2019 Feasibilty Study resulted in the service cage hoist being the largest of the three at 5 m (16’) in diameter and supports both a 3500 mm x 1900 mm (138” x 75”) double deck main service cage with a service capacity of 20,000 kg (44,092 lb) and a counterweight. The skipping DD hoist is the next largest hoist at 4 m (13’) in diameter and supports twin 11,000 kg (24,250 lb) balanced payload bottom dump skip conveyances. The smallest of the three production hoists is the 3 m (10’) diameter SD auxiliary cage hoist which supports a 1500 mm x 900 mm (59” x 35”) double deck cage with a total payload capacity of 5,000 kg (11,000 lb).

New hoisting duty calculations have been conducted in 2021, re-affirmed as part of this study with minor adjustments, in order to assess the opportinuty of optimizing the hoist plants. The results of this review have confirmed that the 2019 hoists are generously sized to meet the revised mine plan and resulting hoist duty.

Section 23.1.8 discusses the results of the optimization review and how the hoist plants can be reduced, all while maintaining the identified hoisting rates. This optimization will result in smaller equipment, a reduction in installed motor power and lowering of the capital cost.

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Source: Nordmin, 2019

Figure 13-20: Production Shaft Hoisting Plant

All three hoists will be powered through the main mining substation, which is dedicated to mine infrastructure both on surface and underground. Additionally, a 2 MW diesel powered generator

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set will supply back up power to the auxiliary hoist, ventilation system and one air compressor feeding compressed air to the underground workings.

Similarly to the production shaft, the 2019 Feasibility Study work resulted in the ventilation shaft hoisting plant (see Figure 13-21) being comprised of one 3 m (10’) diameter SD auxiliary cage hoist which supports a 1500 mm x 900 mm (59” x 35”) double deck cage with a total payload capacity of 5,000 kg (11,000 lb). The auxiliary hoist and cage are duplicates of those installed within the production hoisting plant and shaft, thus ensuring a common platform for both systems. Opportunities for optimization are similar to those for the production hoist and further discussed in Section 23.1.8.

As within the production shaft plant, the auxiliary cage hoist in the ventilation shaft is powered both by the main surface infrastructure and a secondary back up diesel generator. This ensures not only a secondary means of mechanical egress from underground but twinned mechanical egresses from either the production shaft or the ventilation shaft.

 

Source: Nordmin, 2019

Figure 13-21: Ventilation Shaft Hoisting Plant

13.7.13  Dust Suppression System

A multi-zoned underground dust suppression system will aid in reducing the amount of air born dust created by the ore and waste handling systems. Each ore and waste transfer point within the underground material handling system will have a set of air and water nozzles fed from both the

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compressed air and process water systems. The compressed air and process water are piped to a regulating station cabinet where the fluids are cleaned (filter and strainer) and pressure regulated.

From the cabinet, the individually cleaned, and regulated process water and compressed air are fed to a set of strategically placed nozzles which combine the streams and produce a light mist which is enough to drive airborne dust down. Regulating stations will be located at all material handling open draw areas including truck dump sites, apron feeder area, vibratory feeders, conveyor systems, transfer car chutes, the crushing station, the transfer area downstream of the crusher, bin conveyor feed area, bin discharge areas, and the flask infeed area.

13.7.14  Communications System

The mine will be equipped with a leaky feeder system that will allow internet, phone, and radio communications underground. The mine will have standard underground call phones with intercom. A control system will allow remote operation of the rock breaker and CCTV system to monitor dump points, crusher, and key material handling locations.

13.7.15  Safety and Health

Mine Safety and Health Administration (MSHA) safety standards are incorporated in the mine design and include dual secondary means of mechanical egress, backup power for both auxiliary hoists, partial ventilation system and one air compressor which feeds compressed air to the underground. Twelve-person mobile refuge chambers are included and will be in active working areas over the LOM. In addition, there is a cut-out on both the 530 Level and 650 Level to facilitate the installation of two permanent 30-person refuge chambers.

The mine will have a communications system that has both mine phones and wireless communication through a leaky feeder system. A mine rescue team will be required to support the mine’s underground operation. The mine safety program will integrate with local providers in case of any mine emergency. Additionally, a stench gas emergency warning system will be installed in the mine’s intake ventilation system. This system can be activated to warn underground employees of a fire situation or other emergency whereupon emergency procedures will be followed. The shop areas and underground fueling station will be equipped with automatic closure doors that will operate in case of fire.

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14.   PROCESSING AND RECOVERY METHODS

14.1   Process Plant Design Criteria

14.1.1   Surface Crushing, Ore Storage & Mineral Processing Plant

The primary driver of the comminution circuit design is the dry processing of ore, which will be used to avoid an expensive drying operation prior to acid leaching.

The process design relies upon two things; receiving a primary crusher product with a characteristic particle size of (P₈₀) 115 mm at the comminution circuit feed bin and producing feed material for the downstream hydrometallurgical processing at a characteristic particle size of (P₈₀) 1.1 mm.

The primary crusher product will be fed to the secondary cone crusher system, operating in closed circuit with a double deck screen. The screen undersize from the cone crusher system will be fed to an HPGR unit, operating in closed circuit with another double deck screen. The HPGR screen undersize is the comminution product that will report to the hydrometallurgical process. The process design criteria are provided in Table 14-1.

14.1.2    Hydrometallurgical Plant

The purpose of the Hydromet Plant is to extract the pay metals while separating them from the impurities. The process involves a series of successive leach and purification steps. The hydrometallurgical process design criteria has been established based on bench and pilot scale test work, conducted by SGS, Hazen and KPM, as well as similar projects, and standard industry practices. The process design criteria is provided in Table 14-2.

14.1.3    Pyrometallurgical Plant

The purpose of the Pyromet Plant is to reduce the niobium pentoxide in the Hydromet feed by converting it into a saleable ferroniobium metal. The Pyromet also plays an important role in the purification of the FeNb by removing excess Ti in the slag portion of the smelting. Since niobium is commonly alloyed with various high-grade steels to significantly increase their mechanical properties, producing ferroniobium metal is an attractive and suitable option to be created for use in the steel industry.

The pyrometallurgical process design criteria was established based on thermodynamic calculations, inspired by test results completed by KPM and supported by the literature available on the aluminothermic reduction as well as on the niobium pyrometallurgy. Table 14-3 presents the pyromet process design criteria.

The aluminothermic reduction has been selected as the technology to convert the hydrometallurgical Nb₂O₅ precipitate into a FeNb metal. Aluminum shots and iron oxide pellets will be introduced on a continuous basis along with the fluxing agents to initiate and complete the exothermic chemical reduction of the Nb₂O₅. This reduction is performed in a single electrical arc furnace with a continuous feed of precipitate, additives and fluxes to produce a saleable FeNb metal alloy.

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Table 14-1: Process Design Criteria 

Description	Value	Unit
Throughput and Operational Time
Non-operational Time	0	h/a
Planned Down Time	252	h/a
Unplanned Down Time	1,276	h/a
Available Time	7,232	h/a
Availability	85	%
Annual Design Throughput	1,008,129	t/a
Process Plant Throughput	125	t/h
Ore Characteristics
Average Specific Gravity	2.96	-
Moisture in Ore	5	%
Bulk Density	1.8	t/m3
Angle of Repose	37	degrees
Angle of Reclaim	60	degrees
Test Work Parameters
JK Drop Weight Test
A x b - Maximum	58.4	-
A x b - Minimum	44.3	-
SMC Test
A x b - Maximum	56.4	-
A x b - Minimum	34.9	-
M,a - Design	19.7	kWh/t
Mih - Design	15.0	kWh/t
Crushability and Grindability Tests
Cwi	12.0	kWh/t
Rw, - Design	17.9	kWh/t
Bw, - Design	15.4	kWh/t
A, - Design	0.112	g
Crushing Circuit	139	t/h
Feed Rate to Secondary Crusher
Primary Crusher Product Size (Pao)	115	mm
Primary Crusher Product Size (Ploo)	203	mm
Crushed Ore Bin Reclaim Feeder Type	Vibrating Feeder
Design Feeder Capacity (Total)	160	t/h
Number of Feeders	3	-
Secondary Crusher Screen
Screen Type	Double Deck Vibratory
Number of Screens	1	-
Fresh Feed Throughput	139	t/h
Secondary Crusher Recycle Throughput	171	t/h
Total Screen Feed	311	t/h
Number of Decks	2	-
Top Deck Opening Size	50	mm
Bottom Deck Opening Size	25	mm
Product Sze (Pao)	22.4	mm
Screen Size - Area	18	m2
Secondary Crusher
Crusher Type	Cone
Average Throughput	171	t/h
Number of Units	1	-
Feed Size - Maximum (Firm)	203	mm
Feed Size (Fao)	115	mm
Close Side Setting	25	mm

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Product Size (Pao)	26	mm
Selected Crusher Size	HP300 or Equivalent	-
Crusher Motor Size	200	kW
HPGR Circuit
Crusher Type	HPGR
Feed Size (Fao)	22.4	mm
Fresh Feed Throughput	139	t/h
Total Throughput	198	t/h
Number of Units	1
Specific Energy Consumption	4.18	kWh/t
Selected Size	POLYCOM 14/08 - 02 or Equivalent
Installed Power	1,000	kW
Product Size (Pao)	1.1	mm
HPGR Screen
Screen Type	Double Deck Vibratory
Number of Screens	1	-
Screen Throughput	198	t/h
Screen Recycle Throughput (to HPGR)	59	t/h
Top Deck Opening Size	6	mm
Bottom Deck Opening Size	3	mm
Product Size (Pao)	1.10	mm
Screen Size - Area	18	m2
Fine Ore Bin
Fine Ore Bin - Storage Time	48.0	h
Crushed Ore Bin - Live Capacity	6,000	t
Fine Ore Bin Reclaim Feeder		-
Feeder Type	Vibrating Feeder
Design Feeder Capacity (Total)	144	t/h
Number of Feeders	3	-

Source: Tetra Tech, 2017

Table 14-2: Hydrometallurgical Processing Design Criteria 

Description	Value	Unit	Source
HCI Leach Unit
General feed characteristics
Temperature	AMB		HMB
Percent Solids	95%
Mass flow rate	2764	dmt/d
WPL Feed Composition
Nb2O5	0.81	%w/w	Mine Plan
TiO2	2.86	%w/w	Mine Plan
Sc2O3	100.75	ppm	Mine Plan
Al203	2.24	%w/w	PEA II - PDC
BaO	4.41	%w/w	Mine Plan
CaO	16.98	%w/w	Mine Plan
FeO	7.36	%w/w	Mine Plan / Test Work Data
Fe2O3	9.60	%w/w	Mine Plan / Test Work Data
K2O	1.60	%w/w	PEA II - PDC
MgO	8.66	%w/w	Mine Plan
MnO	0.62	%w/w	PEA II - PDC
Na2O	0.24	%w/w	PEA II - PDC

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P205	0.83	%w/w	PEA II - PDC
Si02	9.68	%w/w	Mass Balance
SrO	0.27	%w/w	PEA II - PDC
ZrO2	270.00	ppm	Test Work Data
CO3	34.00	%w/w	v01 Mine Plan
ThO2	472.0	ppm	Mine Plan
UO3	58.5	ppm	Mine Plan
REE2O3	0.363	%w/w	Engineering Design
Leach Conditions
Temperature End	40	°C	Test Work/Eng. Design
Residence time	3.3	h	Test Work/Eng. Design
Acid Bake Unit
PUG Mill
Residence Time	1.0	H	Test Work/Eng. Design
Acid Addition Rate	925	kg/mt
Temperature	220	°C	Test Work/Eng. Design
Solids Fraction at Discharge	81%	wt%	Mass Balance
Hollow Flight
Residence Time	1.5	H	Test Work/Eng. Design
Temperature	300	°C	Test Work/Eng. Design
Water Leach Conditions
Temperature	35	°C	Test Work/Eng. Design
Water addition rate	3	kg/kg WPL(s)	Test Work
Solids Fraction at Discharge	14.5%	wt%	Mass Balance
Iron Reduction Tank
Temperature	Variable	°C	Test Work
Residence Time	1	H	Test Work
Iron Addition Rate	0.375	Stoich Fe+Ti	Test Work
Nb Precipitation Conditions
Temperature	90-100	°C	Test Work/Eng. Design
Residence Time	4	H	Test Work
Solids Fraction at Discharge	1.2%	wt%	Mass Balance
NbP Calcination Conditions
Temperature	950	°C	Test Work/Eng. Design
Nb Caustic Leach Conditions
Temperature	105	°C	Test Work
Caustic Addition	1.00	kg/kg	Test Work
Caustic Solution Strength	35.00	wt%	Test Work
Dilution Ratio (Water to NaOH Solution)	5.23		Test Work
Solids Fraction at Discharge	12.6%	wt%	Mass Balance
TiP Neutralization
Temperature	Ambient	°C	Test Work/Eng. Design
Target Acidity	15.00	gpl	Test Work
Ti Precipitation
Temperature	100	°C	Test Work/Eng. Design
Residence Time	2	H	Test Work
TiP Calcination Conditions

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Temperature	950	°C	Test Work/Eng. Design
Sc Precipitation Iron Reduction
Temperature	75	°C	Test Work
Residence Time	0.25	H	Test Work
85% Phosphoric Acid Addition	4.5	kg/m3	Mass Balance
Iron Powder Addition	2.9	kg/m3	Mass Balance
Sc Precipitation
Temperature	75	°C	Test Work
Target pH	3.25		Test Work
Reagent Used	15.30	kg/m3	Mass Balance
Sc Releach
Temperature	85	°C	Test Work/Eng. Design
Target pH	0.37		Test Work
Hydrochloric Acid Strength	20%	wt%	Test Work
Sc Solvent Extraction Circuit
Solvent Conditioning Acid Preparation Tank
HCI Addition Ratio	0.185	m3(a) / m3(o)	Test Work
HCI Concentration	36%	wt%	Engineering Design
Scandium Extraction Mixer-Settlers
Mixer Retention Time	3.4	min	Test Work
Overall O:A Ratio	1:8		Mass Balance
Internal O:A Ratio	1:1		Engineering Design / Test Work
Settler Retention Time	6.8		Test Work
Solvent Wash Mixer-Settler
Hydrochloric Acid Wash Solution Strength	36%	wt%	Engineering Design / Test Work
Mixer Retention Time	19	min	Test Work
Overall O:A Ratio	1:6
Internal O:A Ratio	1:1
Settler retention time	68	min
Scandium Scrub Mixer-Settlers
Mixer retention time	16	min	Test Work
Overall O:A Ratio	1:4		Mass Balance
Internal O:A Ratio	1:1		Engineering Design / Test Work
Settler retention time	43	min	Test Work
Scandium Stripping
Retention time	50	min
Scandium Hydroxide Leach
Sulphuric Acid Concentration	96	wt%
Target Residual Sulphuric Acid Concentration	50	gpl
Scandium Oxalate Precipitation
Temperature	75	°C	Test Work
Residence Time	1	H
Sulphate Conversion Conditions
Temperature	1050	°C	Test Work
Calcium Loop Calciner Sulphate Conversion Conditions

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Temperature	1050	°C	Test Work
Tailings Neutralization
Temperature	Ambient	°C	Test Work/Eng. Design
Residence Time	1	H
Tailings Calcination Conditions
Temperature	1050	°C	Test Work/Eng. Design

Source: Tetra Tech, 2017

Table 14-3: Pyrometallurgical Processing Design Criteria 

Section	Description	Value	Units
Nb2O5 Precipitate Pelletized	Nb2O5 Precipitate Feed Rate (Dry Basis)	2.94	t/h
		64.7	t/d
	Moisture Content (After Pelletizing)	<1	%
	Nb Precipitate Pellets d80	8	mm
Niobium Pentoxide Precipitate Composition	Nb2O5	30.5	%w/w
	TiO2	63.5	%w/w
	P2O5	0.4	%w/w
	Al2O3	0.4	%w/w
Nb Precipitate Pellets	Pellets feed rate	2.94	t/h
	Number of bins	1	#
	Storage time	4.6	days
	Capacity	324	t
Aluminum (Al) pellets	Aluminum (Al) feed rate	0.52	t/h
	Number of bins	1	#
	Storage time	13	days
	Capacity	162	t
Hematite (Fe2O3) Pellets	Hematite (Fe2O3) feed rate	0.48	t/h
	Number of bins	1	#
	Storage time	13	days
	Capacity	150	t
Sodium Dioxide (Na2O)	Feed rate	0.03	t/h
	Super sacks rack (1 Tm or 2Tm)	1	#
Limestone (CaCO3)	Limestone feed rate	0.174	t/h
	Super sacks rack (1 Tm or 2Tm)	1	#
FeNb Furnace – Aluminothermic Reduction	Total Feed to FeNb Furnace	4.18	t/h
	Operating Temperature	1900	°C
FeNb Furnace Power	Electric Arc Furnace	754	kWh
	Power Consumption Per Tonne Precipitate Pellets	182	kWh/t
	Furnace Thermal Efficiency	60.0	%
	Furnace Design Power	1000	kW
	Nb Recovery	96	%
Furnace Cooling system	Water Flow Rate	64.7	m³/h
	Cooling Tower	1	#
FeNb Furnace - FeNb Alloy Composition	Nb	63.3	%w/w
	Fe	33.2	%w/w

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	Ti	0.9	%w/w
	P	0.3	%w/w
	Al	1.4	%w/w
FeNb Alloy Tapping	FeNb Alloy Flowrate	0.917	t/h
	FeNb Alloy Tapping Schedule	2	taps/12-hour shift
		4	taps/day
	FeNb Alloy Tapping Time	10.0	min/tap
	FeNb Alloy Tapping Flowrate	5.5	t/tap
		22.0	t/d
	FeNb density	8.2	t/m3
Furnace Slag Rate	Slag Average Production Rate	2.96	t/h
Furnace Slag Composition	Nb2O5	1.0	%w/w
	Fe2O3	1.1	%w/w
	FeNb (particles)	0.3	%w/w
	Fe2O3	0.8	%w/w
	TiO2	62.3	%w/w
	Al2O3	30.9	%w/w
	CaO	3.2	%w/w
	Na2O	1.5	%w/w
	P2O5	0.02	%w/w
	Slag density	4.0	t/m³
	Slag Flowrate	2.96	t/h
	Slag Tapping Schedule	18	taps/12-hour shift
		36	taps/day
	Slag Tapping Time	15.0	min/tap
	Slag Tapping Flowrate	3.94	t/tap
		71.04	t/day
FeNb Furnace Off-gas Handling	Dusts all recycled to the furnace: Dust loss	0	%
	Generation of CO2 (use of limestone)	0.07	t/day
FeNb Pelletizing system	Cooling water	15.1	m³/h

Source: Metallurgy Concept Solutions, 2019

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14.1.4    Acid Plant

The sulphuric acid plant will primarily be used to regenerate off gas coming from the calciner (SO₂). The design basis is described below.

Calciner-Off Gas

Hot gas from the calciner contains SO₂ and H₂SO₄ which is removed in a gas cleaning system consisting of venturi scrubbers in series. The gas is also cooled to remove water from the gas. The gas conditions at the inlet to the drying tower (battery limits) are:

● Temperature                  50°C

● Pressure           -            4 in. WC

● Gas Composition           SO₂          14.51 vol%

● (Wet Basis)                    O₂              0.66 vol%

N₂              63.12 vol%

H₂O          10.69 vol%

CO₂           11.02 vol%

Overall Plant Conversion

Overall SO₂ to SO₃ conversion 99.7%.

Autothermal Limit

The minimum SO₂ concentration that can be handled by the acid plant is 5 vol%.

Turndown

The plant will be capable of operating at 50% of the design capacity.

Product Acid

The product acid produced by the plant will meet the following criteria at the acid plant battery limits:

Temperature                       40°C maximum

Concentration                     96 wt% H₂SO₄ +/-0.5 wt%

Fe Content                         50 ppm maximum

Cooling Water

Cooling Water Supply         30°C

Cooling Water Return          40°C

Instrument Air (oil-free quality)

Pressure                              800 kPa

Dewpoint                             -40°C

Site Barometric Pressure    97.65 kPa

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14.2   Flowsheets and Process Description

14.2.1   Surface Crushing, Ore Storage & Mineral Processing Plant

The ROM ore will be crushed underground in a primary crusher, and the crushed product with a top size of 203 mm and characteristic size (Pao) of 115 mm, will be delivered to the crushed ore bin located at the surface. The ore from the bin will be reclaimed by three vibrating feeders with a total capacity of 136 t/h and passed on to the secondary crusher circuit via the secondary crusher screen feed conveyor.

At the secondary crushing stage, the ore will be sized on a dry, double deck screen with a top deck aperture size of 50 mm and bottom deck aperture size of 25 mm. The screen oversize from both decks will report to the secondary crushing stage. The screen undersize will be conveyed to the HPGR circuit.

The screen oversize fractions will be crushed in a single secondary cone crusher operating with a closed side setting of 25 mm. The secondary crushed product will be sized by the same double deck screen with the primary crusher discharge ore.

The screen undersize, at an approximate characteristic particle size (Pao) of 22 mm, will be further crushed in the HPGR circuit. The HPGR circuit will consist of a single HPGR crusher, with a separate double-deck vibrating screen with top and bottom deck aperture sizes of 6 mm and 3 mm, respectively. The recirculating load of the HPGR circuit is expected to be in the range of 30 to 40% of the circuit new feed.

The HPGR screen undersize will be the final comminution product and is expected to have a characteristic particle size (Pao) of 1.1 mm. The ore will be stored in a fine ore bin, then reclaimed by a vibrating feeder with a design capacity of 132 t/h, and then passed on to the acid leach circuit via the acid leach feed conveyor for further processing. The HPGR conceptual block flow diagram can be seen in Figure 14-1.

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Source: Tetra Tech, 2017

Figure 14-1: HPGR Conceptual Block Flow Diagram

14.2.2   Hydrometallurgical Plant

The role of the Hydromet Plant is to separate the three pay elements Nb, Ti, Sc, from the crushed ore while utilizing processes to minimize the operating cost of the plant. This requires a large amount of acid, both Hydrochloric (HCl) and Sulphuric (H₂SO₄). The Hydromet Plant includes acid recovery processes to lower the operating expense of the process by requiring a small amount of fresh acid and sulphur to be brought onsite. The HCl and H₂SO₄ recoveries are 99% and 85% respectively. The other operating cost reduction comes from utilizing impurities in the ore separated out in the process as reagents in the process, which minimizes the need for fresh reagents brought onsite. The added benefit to utilizing impurities as reagents reduces the amount

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of tailings from the process that needs to go to the Tailings Storage Facility (TSF), reducing the overall size of the storage area.

The hydrometallurgical process is divided into fifteen units:

1. Hydrochloric Acid Leach (605)

2. Sulphuric Acid Bake (610)

3. Water Leach (615)

4. Iron Reduction (620)

5. Niobium Precipitation and Phosphorus Removal (625)

6. Scandium Precipitation (628)

7. Sulphate Calcining and Mixed Oxides Handling (630)

8. Titanium Precipitation (635)

9. Scandium Solvent Extraction (640)

10. Scandium Refining (645)

11. Product Handling and Packaging (650)

12. Sulphuric Acid Plant (655)

13. Hydrochloric Acid Regeneration (660)

14. Tailings Neutralization (665)

15. Tailings Filtration (670)

The majority of the unit processes selected for the hydrometallurgical flowsheet have been extensively reported on in literature and are predominately proven and existing processes. The plant consists of multiple buildings that will house 15 separate physical and chemical processes required to separate the niobium, scandium and titanium that are contained in the ore and to regenerate and recover reagents for reuse. More details can be found in Figure 14-2.

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Source: SMH, 2017

Figure 14-2: Simplified Sheet

Hydrochloric Acid Leach (605)

The Hydrochloric Acid Leach unit is designed to leach the majority of the impurities and the scandium present in the feed material to reduce the size of subsequent process equipment. The feed coming from the Mineral Processing Plant at a rate of 2,764 t/d (115 t/h) is fed to the HCl Leach Feed Bin. The crushed ore is then distributed using screw conveyors to two parallel trains each consisting of a primary HCl leach tank followed by two secondary HCl leach tanks. Hydrochloric acid from the Hydrochloric Acid Regeneration unit is combined with the crushed ore in each train and reacts at a controlled temperature. The discharge slurries (9 wt% solids) from each train of hydrochloric leach tanks are combined and fed to a dewatering and washing circuit consisting successively of a thickener and four parallel filter presses. The solids are washed in a series of counter current washing stages to ensure removal of the residual chloride ions that may be present in the cake moisture. The filtrate and wash liquors are combined along with the thickener overflow and sent to the PLS Aging section ahead of the Scandium Solvent Extraction unit and the Hydrochloric Acid Regeneration unit. The filter cake is sent to the Acid Bake and Water Leach unit.

The PLS aging tank receives the combined thickener overflow, leach filtrate and wash liquor along with the secondary scandium re-leach liquor. Titanium and other minor elements contained in the PLS are oxidized with the use of an oxidizing agent, precipitated and are further separated by a clarifier. The clarified PLS is sent to the Scandium Solvent Extraction unit for scandium recovery while the solids are sent to the Titanium Precipitation unit.

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Sulphuric Acid Bake (610)

The Acid Bake unit is used to convert all of the unleached metal content into sulphate compounds. The Hydrochloric Acid Leach cake is combined with pre-heated sulphuric acid and mixed in a pair of pug mills before being fed into a hollow flight screw and maintained at reaction temperature. The hollow flight provides the necessary reaction time at elevated temperature to convert the metal compounds to sulphate compounds. Off-gas from the Acid Bake is sent to a condensing column where the sulphuric acid is condensed and sent to the Hydrochloric Acid Regeneration unit. The Acid Bake discharge is continuously fed through a discharge lump breaker to the Water Leach unit.

Water Leach (615)

The Water Leach unit is used to solubilize all soluble sulphates while separating non-soluble impurities. The circuit is composed of a series of three cascading agitated tanks discharging to centrifuges. The feed is delivered to the cascading agitated tanks where it is combined with leach water. The discharge from the last cascading tank is pumped to dewatering centrifuges. The centrate, which contains the soluble sulphates, is sent to the Iron Reduction unit while the cake is transported via screw conveyor to a three-stage counter-current washing process. The Water Leach residue from the wash centrifuge is transferred by conveyor to the Paste Backfill Plant.

Iron Reduction (620)

The Iron Reduction unit is used to reduce iron (III) sulphate (Fe₂ (SO₄)₃) present in the solution to iron (II) sulphate (FeSO₄). The titanium (IV) oxysulphate TiOSO₄ is believed to also be reduced to titanium (III) oxysulphate (Ti₂O(SO₄)₂). Addition of iron solids to the solution at room temperature reduces iron and titanium compounds. In this unit, the acidic Water Leach discharge is received into the Iron Reduction column where it is contacted with iron. From this reduction column, the liquid is gravity fed to the agitated reduction tank where the reduction is completed with the addition of more iron. The discharge of the Iron reduction tank is sent to the Niobium Precipitation unit.

Niobium Precipitation and Phosphorus Removal (625)

The Niobium Precipitation unit uses water dilution (RO water) to selectively hydrolyze niobium sulphate and precipitate it as niobium oxyhydroxide. The Iron Reduction discharge is diluted with hot water, acidified with sulphuric acid, and cascaded through a series of agitated tanks. The dilution water to feed volume ratio is 0.6:1, while the sulphuric acid addition is adjusted to provide the required precipitant acid concentration. The precipitation reaction temperature is maintained at or near boiling by steam jacketed agitated tanks. The discharge of the Niobium Precipitation tanks is pumped to a clarifier. The overflow liquid is directed through a polishing filter before being forwarded to the Titanium Precipitation circuit. The solids slurry from the clarifier is pumped to two centrifuges for further dewatering. The centrate is sent to the Titanium Precipitation unit while the cake is sent to the direct fired Niobium Calciner where water is driven off, and niobium oxyhydroxide is oxidized to niobium pentoxide (Nb₂O₅) at 950°C while converting any sulphate trapped in the precipitate to an oxide thus liberating SO₂ in the off-gas. The calcination also converts a portion of the Phosphorus content to a leachable form. The calcined material is cooled in a rotary cooler and then pneumatically transported and fed via a rotary feeder into a series of cascading agitated caustic leach tanks. A Sodium hydroxide solution is added to the calcined niobium concentrate to leach Phosphorus to an acceptable residual concentration. The caustic leach discharge is pumped to another series of tube presses for

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dewatering and washing with RO water. The filtrate and wash liquor are combined and sent to the Tailings Neutralization unit, while the cake is sent to a pelletizer before being fed into a sintering kiln. The discharge from the sintering kiln is then conveyed to the Pyrometallurgical plant for further processing.

Titanium Precipitation (635)

The Titanium Precipitation is achieved through hydrolysis of the titanium oxyhydroxide using heat at a reduced free acid content. The titanium rich solution from the Niobium Precipitation along with the scrubbing liquor from the Scandium Solvent Extraction are partially neutralized using fresh and recycled CaO from the Calcium Loop. The gypsum precipitate containing scandium is filtered on a vacuum belt filter and sent to the rotary kiln where it is calcined back to oxides before being recycled back to the neutralizing tanks. A portion is purged from the loop to maintain impurity levels and sent back to the Hydrochloric Acid Leach unit where the scandium and any trapped titanium is recovered. The off-gas from the Calcium Loop containing the SO₂ is combined with the off-gases from sulphate calciners, is cleaned and sent to the Acid Plant for sulphur recovery.

The titanium rich filtrate from the vacuum belt filter is heated with steam directly injected in a series of agitated tanks where titanium hydrolyzes and precipitates as titanium oxyhydroxide. The slurry is then pumped to a clarifier. The thickened slurry is then fed to tube presses for additional dewatering and washing with RO water. The overflow from the clarifier along with the filtrate and wash liquor from the tube presses are combined and sent to the Scandium Precipitation circuit. The solids from the tube presses are calcined to drive off any remaining sulphur and water to convert the titanium to TiO₂. The titanium dioxide is then sent to the packaging area where it will be loaded into super sacks and/or plastic-lined steel drums according to the client’s specifications.

Scandium Precipitation (628)

The Titanium Precipitation filtrate is fed to the Scandium Precipitation unit where it is first mixed with Phosphoric acid (H₃PO₄). Iron is also added. Magnesium carbonate is used to adjust the pH to ensure the precipitation of the scandium. The slurry is pumped through a clarifier to a filter press where the liquids are separated and recycled to the clarifier. The clarifier overflow is sent to the Tailings neutralization circuit. The cake is conveyed to the Scandium Re-Leach tank where hydrochloric acid is added to re-leach the scandium. The resulting PLS is sent back to the PLS aging tank in the HCl leach circuit before being treated in the Scandium Extraction circuit.

Sulphate Calcining and Mixed Oxides Handling (630)

The sulphate calcining unit recovers sulphur from the different cakes formed throughout the process. The feed comes from two sources: the tailings neutralization filter cake and the HCl Regeneration filter cake. The cake from the Tailings Neutralization unit is processed similarly but separately from the remaining sulphate cake to maintain the desired neutralizing potential of the resulting mixed oxides.

The initial stage (Primary Kiln) of the calcination of the Acid Regeneration cake operates at a lower temperature and recovers the free H₂SO₄ acid, which is not associated with any other elements. All of the water content in the sulphate cake is also driven off in the off-gas stream. This gas stream is condensed in a nearby three-stage condensing and scrubbing circuit. The first stage condenses as much as 98% of the H₂SO₄ and recovers it in a solution containing more than 96% sulphuric acid. The second stage completes the condensation of the H₂SO₄ as 80% sulphuric acid. The last stage finalizes the scrubbing to minimize the H₂SO₄ release. The combined 96% sulphuric

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acid produced in the condensing circuit is sent back to the HCl Regeneration unit. The last stage of the two-stage calcining process of the Acid Regeneration cake reaches elevated temperatures to decompose all of the solid sulphates present in the kiln. Sulphur is added to the sulphate cake at this stage to compensate for sulphur losses and is burned to SO₂ during the decomposition of the sulphates. This decomposition releases SO₂ and H₂O. The sulphur dioxide gas from this kiln is combined with that of the Tails Neutralization sulphate cake calciner and scrubbed in a three-stage scrubbing circuit. The scrubbing circuit is identical to the first stage scrubbing circuit and condenses H₂SO₄ and water from the SO₂ gas before sending it to the Acid Plant for conversion back to H₂SO₄. The mixed oxides produced are pneumatically conveyed to the Tailings Neutralization as a neutralizing reagent.

The calcination of the Tailings Neutralization sulphate cake only has one stage as it contains no free H₂SO₄ acid that is not associated with any other elements. The calcination also releases SO₂ and H₂O. The sulphur dioxide gas from this kiln is combined and scrubbed as described above. The mixed oxides are sent to Tailings Neutralization via screw conveyor. Calcium loop calcined oxides are pneumatically conveyed back to Ti Neutralization and to the HCl Leach. The mixed oxides produced are conveyed to the Paste backfill Plant via belt conveyor. A portion, adjusted to minimize sulphur losses to the wastewater, is recycled back to the Tailings Neutralization for neutralization.

Scandium Solvent Extraction (640)

The Scandium Solvent Extraction unit is a four-stage D₂EHPA solvent extraction circuit followed by a wash stage, a three-stage scrubbing circuit and two-stage Stripping Circuits used to selectively recover scandium from the leach solution. The extracting organic solution is prepared in an agitated tank. The Barren organic from the stripping section is conditioned with HCl to remove any Hydroxides and to convert it from the Na+ form to its H+ form before being recycled to the Extraction section with fresh organic (small amount as required from time to time to adjust volume). The barren organic flows through the extraction mixer-settlers in series. The aqueous feed from PLS Aging is fed countercurrent to the barren organic through the extraction mixer-settlers. The scandium loads to the organic along with titanium and other elements in small amounts. The raffinate is sent to the HCl Regeneration unit while the loaded organic moves on to a single stage wash with an HCl solution to finalize the separation of aqueous impurities from the organic. The titanium and other impurities that loaded on the organic are then scrubbed in a scrubbing section with a solution containing H₂O₂ and H₂SO₄. The Sc rich, loaded organic is then sent to the Stripping Circuit. The scrubbed organic solution is combined with NaOH and agitated in a series of cascading tanks which strips the scandium of the organic and precipitates it. Separation is accomplished by a 3-phase settler where the organic, aqueous and solids are separated. The aqueous phase containing excess NaOH is moved to a holding tank and is recycled back to be re-used as strip solution. The barren organic is sent to a coalescer and is recycled to the conditioning step upstream the Extraction. The settler underflow is further refined in the Scandium Refining unit.

Scandium Refining (645)

Scandium Refining involves removing Zirconium and other impurities such as titanium and niobium from the scandium produced. The process is operated in batches. The first step is to re-dissolve the scandium rich solids in H₂SO₄ and dilute this solution down with RO water. The scandium rich solution is then contacted with a mix of two organics and diluent in a single mixer-settler tank. The Zr, Ti and Nb are loaded onto the organic, which is decanted and moved to a

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stripping stage while the decanted Sc rich raffinate is sent to the oxalate crystallization step. There Zr, Ti and Nb are stripped from the loaded organic with H₂SO₄ in a single mixer-settler tank. The stripped organic solution then moves on to the organic regeneration circuit to be recovered for re-use in another cycle, and the stripping solution is sent to Tailings Neutralization. The Sc rich raffinate is mixed with oxalic acid to form scandium oxalate crystals that are filtered and washed on a belt filter. The filtrate and wash liquor are sent to Tailings Neutralization while the cake is calcined to convert the solids to Sc₂O₃. The calciner discharge is transferred by screw conveyor to a screen where the final product is weighed and bagged for sale.

Sulphuric Acid Plant (655)

Gas from the calciner is delivered to the acid plant drying tower cleaned and cooled to a temperature of 50°C. The gas is deficient in oxygen, so ambient air is added to the clean gas leaving the gas cleaning system to provide sufficient oxygen for the conversion of SO₂ to SO₃ in the acid plant contact section. A minimum O₂:SO₂ ratio of 1:1 is required for efficient reaction.

The diluted gas still contains water which must be removed before the gas enters the contact section. Drying of the gas is done in the drying tower by counter current contact with concentrated sulphuric acid in the packed section. Concentrated sulphuric acid readily absorbs water from the gas. The dry gas leaves the top of the packed section and passes through a mesh pad mist eliminator before leaving the tower.

The dry gas enters the main acid plant blower, which compresses the gas for delivery through the acid plant contact section. The cold gas leaving the blower must first be heated to the catalyst ignition temperature. This is done in a series of gas-to-gas heat exchangers which transfer heat from the hot gas leaving the catalyst beds.

The contact process consists of multiple catalyst bed conversion stages with interstage gas-cooling heat exchangers, followed by two absorption stages. The conversion stages convert SO₂ to SO₃, while the absorption stages capture the SO₃ to produce concentrated sulphuric acid.

Primary conversion is obtained in the first three catalyst beds with the cooling of the process gas between each bed. The gas leaving the third catalyst bed is cooled prior to entering the intermediate absorber tower. In the absorber tower the SO₃ formed up to this stage is absorbed by counter current contact with concentrated sulphuric acid in the packed section. The gas leaving the absorber tower passes through a set of high-efficiency mist eliminators before being reheated to the Bed 4 inlet temperature.

The gas undergoes the final stage of SO₂ to SO₃ conversion in Bed 4. The removal of SO₃ in the

intermediate absorption tower results in a higher overall conversion rate in the final bed. The gas leaving the final catalyst bed is cooled before entering the final absorber tower. In the final absorber tower, the SO3 formed in Bed 4 is absorbed by counter current contact with concentrated sulphuric acid in the packed section. The gas leaving the absorber tower passes through a set of high-efficiency mist eliminators before being discharged to the atmosphere through the acid plant stack.

The acid plant is designed for an overall conversion rate of 99.7% and a product acid concentration of 96%.

Hydrochloric Acid (HCl) Regeneration (660)

In this area, chlorides are recovered in the form of Hydrochloric Acid (HCl) for reuse in the Hydrometallurgy Process. The scandium raffinate stream contains significant concentrations of

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dissolved metal chloride compounds in water. Metal chlorides react with sulphuric acid to produce metal sulphates and HCl gas. The HCl gas is vented to absorber columns to recover the HCl in solution. Sulphate compounds are precipitated as solids in a sulphuric acid solution and recovered by filtration.

The process contains a sulphuric acid recirculation loop. Sulphuric acid is added to the Precipitators and collected in Filtrate Tanks after the solids are filtered. The sulphuric acid from these tanks is heated and returned to the HCl Regen Reactors

Water interferes with precipitation of sulphate solids and must be removed along with HCl. Two water removal steps are provided with an initial feed vacuum flash and a second post-reaction vacuum flash unit.

Tailings Neutralization (665)

The Tailings Neutralization unit is fed by the centrate and filtrate from the Titanium Precipitation unit as well as other acidic tailings streams. The combined feed is reacted with recycled mixed oxides in a series of agitated tanks in order to raise the pH to around 9.5. The discharge slurry is pumped to Tailings Handling.

Tailings Handling (670)

The discharge slurry from Tailings Neutralization is successively dewatered in a thickener and filter presses. The filtrate is returned to the thickener, the filter cake is sent to the sulphate calciner, and the thickener overflow is sent to water treatment.

14.2.3   Pyrometallurgical Plant

The high-level Pyromet Plant flowsheet is presented in Figure 14-3. The selected process is based on the aluminothermic reduction of niobium pentoxide (Nb₂O₅) present in the hydromet precipitate. The dry Nb₂O₅ precipitate pellets are fed by conveyor to the Furnace Feed Preparation Area (FPA), and stored in a closed bin, giving a total of five days storage time. Aluminum grains and hematite Fe₂O₃ pellets to supply iron units are also stored in feed bins with a six-day storage capacity for each component. The additives and reductants complete for the aluminothermic reaction recipe, where super sacks will be used to handle these consumables. The three usage bins are loaded by conveyor while an overhead crane will be used to replace the empty super sacks.

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Source: Tetra Tech, 2017

Figure 14-3: Pyrometallurgical Processing Simplified Flowsheet

Six flow control bin/sacks discharge for aluminum, hematite, limestone and sodium oxide provide measured feed to the FeNb Furnace. According to the need, FeNb metal off-spec material can also be recycled along with the Nb₂O₅ feedstock from the hydromet. All the conveyors under the feed material storage units are on load-cells, as part of the furnace feed preparation mass measurement system which is automatically controlled via PLC.

The furnace feed preparation is performed as a continuous process with specified mass measurement of the Nb₂O₅ precipitate pellets with the required aluminum, hematite, and fluxes to satisfy a “recipe” to produce on-spec FeNb alloy (ratio Fe/Nb = 0.35/0.65). Each ingredient is fed onto the furnace feed conveyor at a pre-determined rate to provide a continuous charge to the furnace.

This allows tight control on continuous feed of the mixed charge into the furnace, to maintain furnace levels of slag and metal alloy. Furnace feeding will be stopped briefly for the tapping of both molten slag and FeNb alloy, according to levels of slag and metal in the furnace.

The tapping of slag and FeNb metal is scheduled over two 12-hour shifts:

1. Slag: 18 taps x per 12-hour shift, 15-minute tapping duration. 1.81 t per tap.

2. FeNb metal: 2 taps x per 12-hour shift, 10-minute tapping duration. 5.5 t per tap.

A tapping drill and clay gun unit is used to open each slag and metal tap-hole and plug each tap-hole with clay after the tap is complete. A molten heel or pool of metal is left remaining in the furnace, with some slag layer covering the metal. This is carried out according to measured furnace levels with the slag and metal masses, providing ongoing control and continuous operation of the furnace. The FeNb furnace will be operated at a temperature in the range of 1850 to 1900°C.

Electrical energy is supplied to the furnace to initiate and maintain heat input into the furnace to complete the reduction of Nb₂O₅ and Fe₂O₃. Aluminum is the primary reductant and on oxidation

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to Al₂0₃ forms a large part of the slag system with TiO₂, fluxed with limestone (CaCO₃) and sodium oxide (Na₂O).

The Al₂0₃-TiO₂-CaO type slag produced in the furnace is tapped into steel molds multiple times (18 times/shift) each shift where it can cool before being moved to a storage bunker area. The steel molds are sectioned so that the slag is easily removed from the mold and is in small, easily crushed pieces. The molds are reused. The slag is moved from the loadout bunker with a front-end loader (FEL) to the vibrating feeder that feeds the Slag Jaw Crusher. The crushed slag is then treated by gravity separation to recuperate the FeNb particles stuck in the slag. The remaining slag is transferred to the tailing’s impoundment.

The FeNb alloy metal is tapped via a short launder into the FeNb pelletizing pan where the molten droplets will solidify in a cold-water basin to form particles ranging in size from approximately 6 mm to 15 mm. The cooled FeNb pellets will then be removed from the basin by a pocket conveyor and transferred to a rotary dryer where the moisture will be driven off prior to screening and packaging. Undersize FeNb pellets are collected and sent to the off-spec feed bin to be reintroduced into the EAF.

Dust from the FeNb Furnace Feed Preparation Area are captured via ducting through a dry cyclone — bag-house system. All dust from this area is returned to the FPA and placed in a separate bin. As required, according to the furnace charge mix recipe, these fines are bled back into the furnace charge for smelting.

The FeNb furnace off-gas, since it contains a fraction of SO2, will be sent to a scrubber for treatment. The slag and metal tapping fumes, and casting fumes above each mold are captured and ducted to the furnace off-gas baghouse. Baghouse dust is recycled to the EAF with undersized FeNb pellets.

Both the Feed Preparation and dust collectors cleaned air exhausts are ducted respectively to their own exhaust stack. Each air exhaust duct may be monitored by sampling to meet environmental regulations.

14.2.4   Acid Plant

Gas from the calciner is delivered to the acid plant drying tower cleaned and cooled to a temperature of 50 °C. The gas is deficient in oxygen, so ambient air is added to the clean gas leaving the gas cleaning system to provide sufficient oxygen for the conversion of SO₂ to SO₃ in the acid plant contact section. The diluted gas still contains water which must be removed before the gas enters the contact section. Drying of the gas is done in the drying tower by counter current contact with concentrated sulphuric acid in the packed section. Concentrated sulphuric acid readily absorbs water from the gas. The dry gas leaves the top of the packed section and passes through a mesh pad mist eliminator before leaving the tower.

The contact process consists of multiple catalyst bed conversion stages with interstage gas-cooling heat exchangers, followed by two absorption stages. The conversion stages convert SO₂ to SO₃, while the absorption stages capture the SO₃ to produce concentrated sulphuric acid.

The acid plant is designed for an overall conversion rate of 99.7% and a product acid concentration of 96%.

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14.3   Process Equipment

14.3.1   Surface Crushing, Ore Storage & Mineral Processing Plant

The primary equipment list (see Table 14-4) and the ancillary equipment list (see Table 14-5) for the comminution area were prepared based on the process design criteria. The installed power of the major equipment determined during the process design is shown in Table 14-4.

Although the ancillary equipment list for the comminution area is shown in this report for completeness, the associated installed power is not determined as part of the process design process. The installed motor power of ancillary equipment is reported as provided by the Qualified Person for materials handling design.

Table 14-4: Primary Equipment List

Comminution Circuit Primary Equipment	No. of Units	Unit Installed Power (kW)	Total Installed Power (kW)
Double Deck Vibrating Screen	1	5.5	5.5
Secondary Crusher (Metso HP300 or equivalent)	1	200	200
High Pressure Grinding Rolls (Polycom 14/08 - 02 or equivalent)	1	1,000	1,000
HPGR Product Double Deck Screen	1	37.5	37.5

Source: Tetra Tech, 2017

Table 14-5: Ancillary Equipment List

Comminution Circuit Ancillary Equipment	No. of Units	Unit Installed Power (kW)	Total Installed Power (kW)
Crushed Ore Bin Vibrating Feeder	3	3.75	11.25
Secondary Crusher Screen Feed Conveyor	1	75	75
Secondary Crusher Recycle Conveyor	1	15	15
HPGR Feed Conveyor	1	11.5	11.5
HPGR Screen Feed Conveyor	1	11.5	11.5
HPGR Recycle Conveyor	1	22	22
Fine Ore Bin Feed Conveyor	1	30	30
Fine Ore Bin	1	Not Applicable	Not Applicable
Fine Ore Bin Vibrating Feeder	3	3.75	11.25
Fine Ore Conveyor	1	30	30
Conveyor Scales	1	Not Determined	Not Determined
Tramp Magnet	1	3.75	3.75

Source: Tetra Tech, 2017

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14.3.2   Hydrometallurgical Plant

The equipment list of the Hydrometallurgical Plant was developed based on the design criteria and using the mass balance provided by the METSIM model. Table 14-6 provides a summarized list of equipment for reference in this report. A more detailed list and sizing were used in the capital cost estimate.

Table 14-6: Summarized List of Equipment

Equipment Number	Equipment Name	Qty
600-BOL-001	BOILER	1
600-OCR-001	BRIDGE CRANE 20 TON CAPACITY	2
605-BIN-001	HCI LEACH FEED BIN	1
605-BIN-002	DISCHARGE BIN	1
605-CVO-001	ACID MIXER BIN FEED CONVEYOR	1
605-CVO-002	HCI LEACH RESIDUE FEED CONVEYOR	4
605-FPR-001	HCI LEACH RESIDUE FILTER PRESS	4
605-HTX-001	HEAT EXCHANGER	1
605-SCC-001	HCI LEACH FEED SCREW CONVEYOR	2
605-SLP-001	HCI LEACH RESIDUE THICKNER PUMP	4
605-SLP-003	HCI LEACH RESIDUE PUMP	3
605-SLP-005	CLARIFIER PUMP	2
605-SOP-003	INTERMEDIATE WASH PUMP	1
605-SOP-005	FLOCC MIX TANK PUMP	1
605-SOP-006	INTERMEDIATE WASH PUMP	1
605-SOP-103	INTERMEDIATE WASH PUMP	1
605-SOP-105	FLOCC MIX TANK PUMP	1
605-SOP-106	INTERMEDIATE WASH PUMP	1
605-TAK-001	HCI LEACH TANK	6
605-TAK-005	HCI LEACH RESIDUE SURGE TANK	1
605-TAK-006	INTERMEDIATE WASH TANK	2
605-TAK-009	HCI RESIDUE FEED TANK	1
605-TAK-010	PLS AGING TANK	1
605-TAK-011	TANK CLARIFIER	1
605-THK-001	HCI LEACH RESIDUE THICKNER	1
610-BIN-001	PUG MILL FEED BIN	1
610-BOL-001	PUG MILL HEATER	1
610-CON-001	ACID CONDENSING COLUMN	1
610-HTX-002	ACID PLATES HEAT EXCHANGER	1
610-PUG-001	PUG MILL	3
610-SCC-001	PUG MILL FEED SCREW	2
610-SCR-001	ACID CONDENSING SCRUBBER	1
610-SLP-001	PUG MILL ACID FEED PUMP	2
610-SOH-001	ACID FEED TANK HEATER	1

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610-SOP-002	SPENT SCRUBBER SOLUTION PUMP	2
610-SOP-005	ACID CONDENSING COLUMN DISCHARGE PUMP	2
610-TAK-001	ACID BAKE ACID FEED TANK	1
615-CEN-001	WL CENTRIFUGE	2
615-CEN-003	WL RESIDUE WASHING CENTRIFUGE	3
615-PBO-001	WL CENTRATE PUMPBOX	1
615-SCC-001	WL CENTRIFUGE SCREW CONVEYOR	2
615-SCC-003	WL RESIDUE SCREW CONVEYOR	3
615-SLP-001	WL RESIDUE PUMP	2
615-SLP-003	WL RESIDUE WASHING PUMP	6
615-SOP-001	WL CENTRATE PUMP	2
615-TAK-001	WATER LEACH TANK	2
615-TAK-003	WATER LEACH TANK DECANT FEED TANK	1
615-TAK-004	WL RESIDUE WASHING REPULP TANK	3
620-BIN-001	Fe POWDER BIN	1
620-PAC-001	IRON REDUCTION COLUMN	1
620-SLP-003	IRON REDUCTION SOLUTION TRANSFER PUMP	2
620-TAK-003	IRON REDUCTION TANK	1
625-BIN-001	NbP ROTARY COOLER DISCHARGE BIN	1
625-BIN-003	NbP PELLET STORAGE	1
625-BIN-004	NbP ROTARY COOLER FEED BIN	1
625-CAL-001	NbP CALCINER	1
625-CEN-001	Nb PRECIPITATION CENTRIFUGE	2
625-CVO-001	NbP TUBE PRESS DISCHARGE CONVEYOR	2
625-DIL-001	DILUTION SKID	1
625-HPP-001	TUBE PRESS HYDRAULIC POWER PACK	2
625-HPP-002	TUBE PRESS HIGH PRESSURE PUMP SYSTEM	2
625-HVS-001	TUBE PRESS HYDRAULIC VACUUM SYSTEM	2
625-ILF-001	NbP CLARIFIER 0/F AUTOMATIC BACKWASH FILTER	1
625-KLN-001	NbP SINTERING KILN	1
625-LPP-001	TUBE PRESS LOW PRESSURE PUMP SYSTEM	2
625-PBO-001	NbP CLARIFIER 0/F PUMPBOX	1
625-PBO-003	TUBE PRESS FILTRATE PUMPBOX	1
625-PEL-001	NbP PEL MIXER	1
625-RTC-001	NbP ROTARY COOLER	1
625-SLP-001	NbP CLARIFIER FEED PUMP	2
625-SLP-002	NbP CLARIFIER U/F PUMP	2
625-SLP-003	NbP CLARIFIER 0/F PUMP	2
625-SLP-005	NbP TUBE PRESS FEED PUMP	2
625-SLP-006	NbP CAUSTIC LEACH TUBE PRESS FEED PUMP	2
625-SLP-110	NbP CENTRIFUGE FEED PUMP	1
625-SOP-001	TiP NEUTRALISATION FEED PUMP	2
625-SOP-004	NbP CAUSTIC LEACH TUBE PRESS FILTRATE PUMP	2

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625-TAK-001	NbP TANK	4
625-TAK-005	NbP CLARIFIER FEED TANK	1
625-TAK-007	NbP TUBE PRESS FEED TANK	1
625-TAK-008	NbP CAUSTIC LEACH TANK	2
625-TAK-010	NbP CAUSTIC LEACH FEED TANK	1
625-TAK-011	NbP CLARIFIER	1
625-TAK-012	TiP CENTRIFUGE FEED TANK	1
625-TAK-013	NbP TANK	2
625-TAK-015	NbP SOLUTION BUFFER TANK	1
625-TUF-001	NbP TUBE PRESS	3
625-TUF-004	NbP CAUSTIC LEACH TUBE PRESS	3
628-BIN-001	NgCO3 Bin	1
628-CVO-001	Sc FILTER PRESS COLLECTION CONVEYOR	1
628-FPR-001	Sc FILTER PRESS	1
628-SLP-001	Sc RE-LEACH PUMP	1
628-SLP-002	PO4 ADJUSTMENT PUMP	1
628-SLP-003	Sc PRECIP PUMP	2
628-SLP-005	Sc PRECIP CLARIFIER U/ PUMP	2
628-SLP-006	FILTER FEED TANK PUMP	2
628-SLP-101	Sc RE-LEACH PUMP	2
628-SLP-102	PO4 ADJUSTMENT PUMP	2
628-TAK-003	Sc PRECIP TANK	1
628-TAK-004	Sc PRECIP SLURRY TANK	1
628-TAK-005	TANK	1
628-TAK-006	TANK	1
628-TAK-007	Sc PRECIP CLARIFIER	1
628-TAK-008	FILTER FEED TANK	1
630-BIN-001	DISCHARGE BIN	1
630-BIN-002	DISCHARGE BIN	1
630-BIN-003	DISCHARGE BIN	1
630-BIN-005	DISCHARGE BIN	1
630-CON-002	ACID CONDENSING COLUMN	3
630-CON-004	VENTURI SCRUBBER	1
630-CYC-001	CYCLONE	3
630-CYC-004	TAILS NEUT KILN CYCLONE	1
630-HTX-001	ACID PLATES HEAT EXCHANGER	4
630-ILF-001	VENTURI SCRUBBER FILTER	1
630-KLN-002	SECONDARY KILN	3
630-KLN-007	TAILS NEUT KILN	1
630-SCC-001	SCREW CONVEYOR DRYER	3
630-SCC-005	DISCHARGE BIN SCREW CONVEYOR	1
630-SCC-006	MIXED OXIDE SCREW CONVEYOR	3
630-SCR-001	ACID CONDENSING SCRUBBER	2

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630-SOP-005	ACID CONDENSING COLUMN DISCHARGE PUMP	2
630-SOP-008	ACID CONDENSING SCRUBBER DISCHARGE PUMP	4
630-SOP-009	VENTURI SCRUBBER DISCHARGE PUMP	2
630-SOP-010	ACID CONDENSING COLUMN DISCHARGE PUMP	4
635-BIN-003	KILN DISCHARGE STORAGE BIN	1
635-BIN-004	WPL FEED STORAGE BIN	1
635-BIN-005	LIME POWDER FEED BIN	1
635-BIN-001	PRODUCT STORAGE BIN	1
635-BIN-002	TiP NEUT FEED BIN	1
635-CVO-001	TiP TUBE PRESS DISCHARGE CONVEYOR	1
635-CVO-002	TiP NEUT PRESS DISCHARGE CONVEYOR	1
635-CYC-001	CYCLONE	1
635-DRM-001	OVERSIZE DRUM	1
635-DRY-001	TiP DRYER	1
635-DUC-001	DUST COLLECTOR	1
635-FPR-001	TiP NEUTRALIZATION FILTER PRESS	1
635-HPP-001	TUBE PRESS HYDRAULIC POWER PACK	1
635-HPP-002	TUBE PRESS HIGH PRESSURE PUMP SYSTEM	1
635-HVS-001	TUBE PRESS HYDRAULIC VACUUM SYSTEM	1
635-ILF-001	TiP CLARIFIER 0/F AUTOMATIC BACKWASH FILTER	1
635-KLN-001	TiP NEUT KILN	1
635-LPP-001	TUBE PRESS LOW PRESSURE PUMP SYSTEM	1
635-PBO-001	TiP CLARIFIER 0/F PUMPBOX	1
635-PNC-001	PNEUMATIC CONVEYING SYSTEM	2
635-SCC-002	KILN DISCHARGE SCREW CONVEYOR	1
635-SCC-001	TiP ROTARY COOLER DISCHARGE SCREW CONVEYOR	1
635-SCN-001	SWECO SCREEN 80 MESH	1
635-SLP-001	TiP CLARIFIER FEED PUMP	2
635-SLP-002	TiP CLARIFIER U/F PUMP	2
635-SLP-005	TiP CLARIFIER 0/F PUMP	2
635-SLP-007	TiP TUBE PRESS FEED PUMP	2
635-SLP-008	TiP NEUT TANK PUMP	2
635-SOP-001	Sc PHOSPHATE PRECIP FEED PUMP	2
635-SUP-001	TiP SUMP PUMP	2
635-TAK-001	TiP NEUTRALIZATION TANK	2
635-TAK-002	TiP TANK	2
635-TAK-007	TiP CLARIFIER FEED TANK	1
635-TAK-010	TiP TUBE PRESS FEED TANK	1
635-TAK-011	TiP CLARIFIER	1
635-TAK-013	TiP SOLUTION BUFFER TANK	1
635-TUF-001	TiP TUBE PRESS	3
640-CLS-001	STRIPPED ORGANIC COALESCER	1
640-SEX-001	EXT MIXER SETTLER	4

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640-SEX-005	ORG WASH MIXER SETTLER	1
640-SEX-006	SCRUB MIXER SETTLER	3
640-SLP-001	AQUEOUS PUMP	2
640-SLP-002	NaOH PRECIPITATION PUMP	2
640-SLP-003	3-PHASE SETTLER UNDERFLOW PUMP	2
640-SLP-004	SCANDIUM PRECIPITATION PUMP	2
640-SLP-007	COALESCER TANK PUMP	2
640-SOP-001	BARREN ORGANIC PUMP	2
640-SOP-002	ACID MIX TANK PUMP	2
640-SOP-003	AQ TANK PUMP	3
640-SOP-007	Sc SX FEED PUMP	2
640-SOP-009	AQUEOUS PUMP	2
640-SOP-010	STRIPPED ORGANIC PUMP	2
640-SOP-012	ORG TANK PUMP	2
640-SOP-014	COALESCER PUMP	2
640-SOP-112	ORG TANK PUMP	2
640-TAK-001	BARREN ORGANIC HOLDING TANK	1
640-TAK-002	EXT ACID MIX TANK	1
640-TAK-003	AQ TANK	1
640-TAK-007	Sc SX FEED TANK	1
640-TAK-008	ORG TANK	1
640-TAK-009	COALESCER TANK	1
640-TAK-010	SCANDIUM PRECIPITATION TANK	2
640-TAK-012	STRIPPED ORGANIC HOLDING TANK	1
640-TAK-013	NaOH PRECIPITATION TANK	1
640-TPS-001	3-PHASE SETTLER	1
645-BEF-001	Sc OXALATE BELT FILTER	1
645-BIN-001	PRODUCT STORAGE BIN	1
645-CAL-001	Sc CALCINER	1
645-CVO-001	Sc OXALATE TRANSFER CONVEYOR	1
645-DRM-001	OVERSIZE DRUM	1
645-PSC-001	PLATFORM SCALE	1
645-SCN-001	SWECO SCREEN 80 MESH	1
645-SLP-002	Sc PRECIPITATION FEED PUMP	2
645-SLP-003	Sc PRECIPITATION TRANSFER PUMP	2
645-SLP-004	Sc PRECIPITATION WASH PUMP	2
645-SLP-005	Sc FILTER FEED PUMP	2
645-SLP-006	FILTRATE PUMP	2
645-SLP-008	Sc DISSOLUTION LIQUOR PUMP	1
645-SLP-009	ZR STRIPPING FEED PUMP	1
645-SLP-010	ORGANIC REGEN FEED PUMP	1
645-SLP-011	ORGANIC REGEN DISCHARGE PUMP	1
645-SLP-012	ORGANIC BARREN PUMP	1

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645-SLP-013	Sc LIQUOR PUMP	1
645-SLP-014	STRIP LIQUOR PUMP	1
645-SLP-015	Na2CO3 MIXING TANK PUMP	1
645-TAK-002	Sc PRECIPITATION FEED TANK	1
645-TAK-003	Sc OXALATE PRECIPITATION TANK	1
645-TAK-004	Sc OXALATE FILTER FEED TANK	1
645-TAK-009	Sc DISSOLUTION LIQUOR TANK	1
645-TAK-010	ZR LOADING FRP TANK	1
645-TAK-011	ZR STRIPPING FRP TANK	1
645-TAK-012	ORGANIC REGEN TANK	1
645-TAK-013	ORGANIC/AQU SEPERATION TANK	1
645-TAK-014	Na2CO3 MIXING TANK	1
660-TK-90 A-C	HCI REGEN FEED TANKS	3
660-H-90	SALT HOPPER	1
660-TK-100 A-B	FEED WATER FLASH TANK	2
660-AG-100 A-B	FEED WATER FLASH TANK AGITATORS	2
660-R-110 A-C	HCI REGEN REACTORS	3
660-AG-110 A-C	HCI REGEN REACTORS AGITATORS	3
660-TK-115 A-C	VACUUM FLASH TANKS	3
660-AG-115 A-C	VACUUM FLASH TANKS AGITATORS	3
660-TK-120 A-C	PRECIPITATORS	3
660-AG-120 A-C	PRECIPITATORS AGITATORS	3
660-TK-125	FILTRATE TANK	1
660-TK-105	WEAK HCI DISTILLATE DRUM	1
660-TK-135	CONCENTRATED HCI RECEIVER	1
660-C-105 A-B	WEAK HCI ABSORBER COLUMN	2
660-C-105 A-B	WEAK HCI ABSORBER INTERNALS	2
660-C-130 A-B	CONCENTRATED HCI ABSORBER COLUMN	2
660-C-130 A-B	WEAK HCI ABSORBER INTERNALS	2
660-HX-100 A-B	FEED FLASH HEAT EXCHANGERS	2
660-HX-120 A-C	PRECIPITATOR COOLERS	3
660-HX-125 A-D	SULPHURIC ACID RECYCLE HEATERS	4
660-HX-105 A-R	WEAK ACID CONDENSERS	18
660-HX-110 A-B	HCI REGEN CONDENSERS	2
660-HX-130 A-H	CONCENTRATED ACID CONDENSERS	8
660-P-90 A-B	HCI REGEN FEED PUMPS	2
660-P-100 A-B	FEED FLASH PUMPS	4
660-P-110 A-C	HCI REGEN REACTOR PUMPS	6
660-P-115 A-C	VACUUM FLASH PUMPS	6
660-P-120 A-C	PRECIPITATOR PUMPS	6
660-P-125 A-B	FILTRATE PUMP	2
660-P-106 A-C	WEAK HCI ABSORBER COLUMN PUMPS	3
660-P-105 A-B	WEAK HCI ABSORBER COLUMN REFLUX PUMPS	2

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660-P-130 A-C	CONCENTRATED HCI ABSORBER PUMPS	3
660-P-135 A-B	CONCENTRATED HCI RECEIVER PUMPS	2
660-F-125 A-B	SOLIDS FILTER	2
660-VP-105 Al, A2	VACUUM PUMP	2
660-ED-90	SALT EDUCTOR	1
660-CVO-1	CONVEYOR #1	1
660-CV0-2	CONVEYOR #2	1
660-CV0-3	CONVEYOR #3	1
660-CV0-4	CONVEYOR #4	1
665-BIN-001	TAILS NEUTRALIZATION FEED BIN	1
665-BIN-002	TAILS NEUTRALIZATION FEED BIN	1
665-ROV-001	ROTARY FEEDER	1
665-ROV-002	ROTARY FEEDER	1
665-SLP-002	TAILS NEUTRALIZATION PUMP	2
665-TAK-002	TAILS NEUTRALIZATION FEED TANK	1
665-TAK-003	TAILS NEUTRALIZATION TANK	2
670-CVO-003	TAILS FILTER PRESS DISCHARGE CONVEYOR	2
670-FPR-001	TAILS FILTER PRESS	2
670-SLP-001	TAILS THICKENER 0/F PUMP	2
670-SLP-002	TAILS THICKENER U/F PUMP	2
670-SLP-003	TAILS FILTER PRESS FEED PUMP	2
670-TAK-001	TAILS THICKENER 0/F TANK	1
670-TAK-002	TAILS FILTER PRESS FEED TANK	1
670-THK-001	TAILS THICKENER	1

Source: Tetra Tech, 2017

14.3.3   Pyrometallurgical Plant

Based on the design criteria and mass balances, major process equipment and some minor equipment has been sized. These pieces of equipment were used to determine the capital and operating costs of the Pyrometallurgical Plant.

An allowance was made for some minor equipment and facilities where required. The major equipment items are listed in Table 14-7.

Table 14-7: Pyrometallurgical Processing Major Equipment List

Equipment Name	Qty	Description/Size/Model
FeNb Furnace Feed Preparation
Nb2O5 Pellets Bins (5 days)	1	4.88 m dia. x 9.76 m height
Aluminum Shot Feed Bin (6 days)	1	4.88 m dia. x 9.76 m height
Fe2O3 Pellet Feed Bin (6 days)	1	4.27 m dia. x 5.48 m height
Nb2O5 Pellet Operations Bin	1	1.00 m dia. x 1.68 m height
Aluminum Shot Operations Bin	1	1.00 m dia. x 1.68 m height
FeNb Off Spec Operations Bin	1	1.37 m dia. x 2.28 m height
Fe2O3 Pellets Operations Bin	1	0.76 m dia. x 0.91 m height

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FeNb Furnace
FeNb Furnace	1	Electric Arc Furnace, 1000kW
FeNb Pelletizing Basin	1	6.00 m3
Rotary Dryer	1	0.61 m dia. x 4.60 m len., 741 kBTU/hr
Slag Jaw Crusher	1
Screening system	1
Cooling Tower	1
Dust Collection	1

Source: Tetra Tech, 2017

14.3.4   Acid Plant

Based on the design criteria and mass balances, major process equipment has been sized and listed in Table 14-8.

Table 14-8: Acid Plant Equipment List

Equipment	Qty	Description	kW
Contact Section
Blower	2	2 operating, single-stage centrifugal, electric motor, lube oil system	5500
Cold Heat Exchanger	1	Shell and tube, CS
Hot Heat Exchanger	1	Shell and tube, SS
Converter	1	4 bed, 304 SS
Intermediate Reheat Exchanger	1	Shell and tube, SS
Cold Reheat Exchanger		Shell and tube, CS
Strong Acid
Drying Tower	1	Packed tower, CS shell, acid resistant brick lining, membrane, ceramic packing, mesh pad mist eliminator, acid distributor
Drying Acid Pumps	2	1 op/1 stdby, vertical submerged centrifugal	240
Drying Acid Cooler	1	Shell and tube, anodically protected
Drying Acid Pump Tank	1	Horizontal, CS shell, acid resistant brick lining
Intermediate Absorber Tower	1	Packed tower, CS shell, acid resistant brick lining, membrane, ceramic packing, high-efficiency mist eliminator, acid distributor

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Final Absorber Tower	1	Packed tower, CS shell, acid resistant brick lining, membrane, ceramic packing, high-efficiency mist eliminator, acid distributor
Absorber Acid Pumps	3	2 op/1 stdby, vertical submerged centrifugal	240
Absorber Acid Cooler	1	Shell and tube, anodically protected
Absorber Acid Pump Tank	1	Horizontal, CS shell, acid resistant brick lining
Product Pumps	2	1 op/1 stdby, vertical submerged centrifugal	25
Strong Acid Area Sump Pump		Vertical centrifugal, UHMWPE	4

Source: Tetra Tech, 2017

14.4   Power Requirements

14.4.1   Surface Crushing, Ore Storage & Mineral Processing Plant

The estimated power requirements for the mineral processing are shown above in Table 14-4 and Table 14-5.

The Mineral Processing total installed power is 4000 kVA. After applying the power factor and a 90% utilization rate, the installed operating power requirement is 2,700 kVA, which gives a total annual electrical energy consumption 23,622 MVAh/y.

The power requirement is estimated from vendor data for major power users such as the crusher, HPGR, vibrating screen and conveyor motors. Building lighting and other smaller users are also included in the total.

14.4.2   Hydrometallurgical Plant

The total installed power for the Hydrometallurgical Process Plant is 25,546 kVA. After applying the power factor and a 92% utilization rate, the installed operating power requirement is 17,244 kVA, which gives a total annual electrical energy consumption 151.053 MVAh/y. A summary unit breakdown is shown in Table 14-9.

Table 14-9: Hydromet Power Requirements

Processing Unit	Units	Value
Sulphuric Acid Plant	kVA	10,133
HCl Regeneration	kVA	3,074
Hydromet Total (all other processes)	kVA	12,339
Total	kVA	25,546

Source: Tetra Tech, 2017

14.4.3    Pyrometallurgical Plant

For the Pyrometallurgical Process Plant, the total installed power will be 5,200 kVA (including the furnace). After applying the power factor and a 90% utilization rate, the installed operating power

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requirement is 3,500 kVA, which gives a total annual electrical energy consumption 30,724 MVAh/y.

The power requirement was estimated based on scoping test work and from calculations from previous FeNb test work (XPS, KPM, and Hazen). Furnace equipment / technology vendors also confirmed the estimated power requirement for the FeNb Furnace, as summarized in Table 14-10.

Table 14-10: FeNb Furnace Power Requirements

Furnace Power Parameter	Units	Value
Electrical Power per ton Furnace Feed	kWh/t	334
Furnace Efficiency	%	60
Total Peak Power Input	kW	950
Furnace Design Power	kW	1,000

Source: Tetra Tech, 2017

14.4.4   Acid Plant

The estimated power requirements for the Acid Plant are shown above in Table 14-8.

14.5   Plant Water

14.5.1   Water Treatment Plant

Water used for all on-site for all process needs and activities will be supplied from mine dewatering activities, local groundwater wells and from a local water utility (Tecumseh Board of Public Works). Mine water will be pumped to the Water Treatment Plant (WTP) that will produce approximately 2,908 gpm of treated water. Approximately 2,154 gpm of water will be produced from the Reverse Osmosis and Evaporation/Crystallization units, and 754 gpm of water will be produced from the Cooling Tower Makeup (CTMU) system.

The Water Treatment System is designed to reduce the hardness, metals, and dissolved solids of the process wastewater, cooling tower blowdown, well/utility and mine water streams. The system consists of precipitation softening, clarification, pH adjustment, multimedia filtration (MMF), and reverse osmosis (RO). Concentrated brine from the RO system will be sent to a thermal evaporator and crystallizer to produce a salt cake for disposal with the distillate being returned and combined with RO permeate for reuse.

The Process Water Treatment System includes the following major equipment units:

1. Process Water Influent Equalization Tank

2. Softening Reactor

3. Clarifiers

4. pH Adjustment Reactor

5. Multimedia Filters

6. Reverse Osmosis Units

7. Sludge Holding Tank

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8. Filter Presses (shared with CTMU system)

9. Evaporator/Crystallizer System

10. Crystallizer Solids Dewatering System

11. Chemical Feed Systems

The following Table 14-11 was used as the design basis.

Table 14-11: Design Requirements

Parameter	Quantity (gpm)	Notes
Plant Source Water	3,375	From mine dewatering and water wells
Pyromet Feed Make-up (2 points)	5	From RO Units
	5	From RO Units
Hydromet Feed	200	From RO Units (Additional Hydromet Feed Water from the Acid Plant (450 gpm) and Potable Water Wells (1575 gpm) are untreated)
Acid Plant	40	From RO Units
Hydromet Cooling Tower	1,415	From RO Units & CTMU System
Pyromet Cooling Tower	1,158	From RO Units & CTMU System
Hydromet Return Water	750	Constituents: Na, Cl, Ca, SO4 and Fe

14.5.1.1   Flow Equalization

Process wastewater and underground mine water will be pumped into an equalization tank. Cooling Tower Blow Down (CTBD) will also be added to this tank since it will contain elevated total dissolved solids and hardness. The tank will also receive intermittent return flows from MMF backwash and the sludge dewatering system. The equalization tank will allow for storage during a shutdown and to sustain consistent flow to the system. The combined process wastewater and mine water will be pumped from the equalization to the softening reactor at a controlled rate. In the case of a system shutdown, it was assumed there would be enough storage capacity to accommodate reduced or no flow of mine water to the treatment system.

14.5.1.2   Softening and Clarification

The combined streams will enter a Turbomix^® softening reactor where chemicals will be added for precipitation softening. The advantage of the Turbomix design is that it promotes precipitation/crystallization of the dissolved particles to maximize their size and density. This results in faster settling rates, improved sludge handling characteristics, and improved sludge thickening and dewatering rates. To enhance the crystallization reaction kinetics and to maximize the density of the settled sludge, a portion of the precipitated sludge collected in the downstream

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clarification process will be recycled back to the Turbomix draft tube. Hydrated lime and soda ash will be fed to the Turbomix based on the flow rate, hardness, and alkalinity of the incoming water. A coagulant also will be added.

The Turbomix reactor will overflow to two flocculating clarifiers to provide redundancy to allow one unit to be taken down for short durations for maintenance. The polymer will be added to the clarifier center well to promote flocculant growth and improve the settling characteristics of the precipitated solids. A rotating rake assembly including two long rake arms will move the settled solids to a center sludge discharge sump. The clarifier rake drive will be equipped with a high torque alarm and an automatic rake lift to raise the rotating rake mechanism should a torque overload condition occur.

The settled sludge will be withdrawn from the bottom of the clarifiers continuously by underflow pumps. The settled softening sludge is expected to have a solids concentration of close to 10%.

The clarifier effluent will be collected in a launder and will exit the clarifier through a drop box and be conveyed to the pH adjustment reactor tank ahead of the multimedia filters. The pH will be reduced to near neutral. This will allow any residual aluminum to precipitate for subsequent removal in the Multimedia Filter (MMF). An oxidant will also be added to this tank for ammonia removal. Water will be pumped from this tank to the MMF to further reduce the suspended solids prior to RO.

14.5.1.3   Multimedia Filtration

The effluent from the pH Adjustment Reactor (MMF Feed tank) is pumped to the MMF System. The goal of the filtration system is to reduce the inlet suspended solids concentration prior to RO. The vessels contain three separate layers of filtration media and a gravel support bed. The gravel supports the top three active filter layers consisting of anthracite, sand and fine garnet. This layered media profile provides a high sediment holding capacity as compared to conventional dual media/sand filters. The larger incoming particles are trapped on the upper layer of the media allowing the smaller particles to continue through the bed where they are trapped in the lower layers, producing a high-quality effluent. A filter aid will be added to the inlet of the MMF to enhance solids-liquid separation process and achieve deep bed filtration versus conventional surface filtration.

During operation, the softened water enters the multimedia filter vessel under pressure at the top and is distributed uniformly over the top layer of the media bed. After passing through the media bed, the filtered service water exits the vessel through the under-drain assembly at the bottom. As the water flows through the media bed, the suspended solids and turbidity present in the feed water will be removed. The filter media bed slowly exhausts from top to bottom. When the turbidity and/or the differential pressure from the media bed approaches a predetermined set point, the media bed is exhausted and is subjected to cleaning/backwash cycle.

14.5.1.4   Reverse Osmosis (RO) System

Filtered water from the MMF is collected in the RO Feed Tank and will be pressurized through a single pass RO system for removal of total dissolved solids. A small portion of the filtered water will be utilized for Multimedia Filter backwash purposes.

The RO process separates dissolved contaminants from the feed water by passing through a semipermeable thin film composite membrane. These membranes remove 95 ^~ 99% of the dissolved solids present in the feed water and essentially perform a complete removal of all particulate matter.

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During operation, the filtered water from the RO feed tank is pumped to the cartridge filter vessels. The water pressure forces the feed water through the filter elements while leaving any residual impurities behind on the filter element surface. The cartridge slowly exhausts, and when they are clogged with impurities, the pressure drop across the cartridge filter system exceeds the desired limit, and the dirty filter elements are taken out of service for replacement. An antiscalant will be added at the RO cartridge filter inlet to prevent any potential scaling issues across the downstream RO system.

The filtered water from the cartridge filter is then pressurized using the RO booster pump and is fed to the first stage membranes in the RO system. The concentrate from the RO system is routed to the RO Reject Tank prior to being discharged. The concentrate will be sent to the evaporation/crystallization process for further concentration. Permeate stream from the system is collected in the RO Product tank where it blends with the distillate from the evaporator and crystallizer and is pumped to the Hydromet process, cooling tower and other water users.

Over a period of time, the RO membrane elements will be subjected to potential fouling by suspended material or sparingly soluble material that may be present in the feed water. Upon an increase of the feed pressure or decline of permeate quantity/quality, the RO system will be taken offline, and the membranes will be cleaned.

14.5.1.5   Cooling Tower Makeup System (CTMU)

The CTMU Treatment System is designed to reduce iron and manganese in the groundwater supply for cooling tower makeup. Limited data was available on the groundwater; data from a nearby farmer’s groundwater well shows that manganese is present at 0.4 mg/L. Cooling tower suppliers typically recommend that manganese be reduced to <0.05 mg/L to prevent deposition and fouling on the cooling tower fill and cooling loop systems. Based on the final water balance, there will be excess RO permeate available to blend with the groundwater (40:60 blend). Based on the projected blended quality, it is expected that the cooling towers can be operated at up to seven cycles of concentration.

The CTMU system will consist of a separate second treatment system to reduce iron and manganese using a filter media for this process. The filter backwash from the CTMU system will be combined with the Process Water Treatment system.

The CTMU Treatment System includes the following major equipment units and redundancy:

1. Manganese Removal Media Filters

2. Sodium Hypochlorite Feed System

The untreated well water will flow through Manganese Removal filters. It is assumed that the well water pressure is adequate for feeding the filters without re-pumping. Normally all filters are online, except when one filter requires backwashing, where the remaining filters handle the design flow. Sodium hypochlorite is injected in-line prior to the filters. The filters operate like the MMF units described previously. Upon high differential pressure, each filter is taken off-line for backwashing. Backwash water will be pumped from the downstream CTMU tank to the filters. Dirty backwash water will be conveyed to the PW sump and sent through the sludge handling system.

The treated water will be collected in a CTMU Tank and pumped to the cooling towers. Excess RO permeate/distillate from the PW treatment system will also be used as CTMU when available. Chemical storage systems will be shared between the CTMU and PW treatment systems.

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14.5.1.6   Sludge Handling

Sludge from the PW will be collected in a sludge storage tank.

Intermittently the sludge from the storage tank will be pumped to the filter presses for dewatering. Pumps are provided to feed the filter presses. Filter press filtrate will flow by gravity to the building sump and then pumped to the Process Water Equalization Tank using sump pumps. The building sump will also receive filter backwash from the MMFs.

14.5.1.7   Evaporation and Crystallization System

Evaporator Brine Flow

The RO concentrate will be processed through an Evaporator/Crystallizer system to produce a salt cake for disposal. The RO concentrate contains a certain amount of alkalinity. In order to prevent calcium carbonate fouling of the Evaporator heat exchanger, it is important to eliminate all the carbonate alkalinity in the feed stream. This is accomplished in a three-stage process: feed acidification with sulphuric acid, feed preheating and feed deaeration/decarbonation. Feed acidification (via metered sulphuric acid addition) is performed within the Evaporator Feed Tank. The sulphuric acid converts the carbonate and bicarbonate ions to CO₂. The CO₂ is subsequently stripped out of the feed stream in the Feed Deaerator following heat recovery in the Feed Preheater. Brine from the Evaporator Feed Tank is pumped to the Feed Preheater where the temperature is increased by exchanging heat with the Evaporator and Crystallizer condensate. The feed then enters the Feed Deaerator where vapour and non-condensable gasses (NCGs) vented from the shell side of the Evaporator, heats the feed and allows for the release of CO₂ to the atmosphere. The feed then enters the Evaporator.

The purpose of the Evaporator is to remove the majority of the water in the most energy and cost-efficient manner prior to the crystallization system. The feed flow enters the vapour body and is pumped up through the center of the heater via Evaporator Recirculation Pump. The recirculating brine stream is introduced into a vertical heat exchanger tube bundle utilizing Veolia’s double distributor plate design. The brine falls down the inside of the heater tubes where it is heated by vapours condensing on the outside of the tubes, causing the brine to boil. The concentrated brine gathers in the vapour body below the heater, where it is recirculated again.

Antifoam can be added to the Evaporator on an as needed basis to ensure that no liquid is carried over through the mist eliminators. Caustic is added to the Evaporator to maintain the pH between 8.0 and 8.5 to ensure the system will not be susceptible to corrosion.

The concentrated brine leaves the Evaporator via a purge line off the discharge of the Evaporator Recirculation Pump and is pumped to the Crystallizer Feed Tank for further concentration.

Low-pressure steam is created by the auxiliary boiler. This steam is utilized for start-up purposes and as supplemental heat for the system when required.

14.5.1.8   Crystallizer Brine Flow

The concentrated brine from the Evaporator is pumped to the Crystallizer Feed Tank. Caustic is again added to the system at the Crystallizer Feed Tank. Caustic is needed at this point to make up for metal hydroxides that precipitate as the brine is concentrated. The target pH in the Crystallizer is 8.0-8.5. The Crystallizer is a forced circulation unit meaning the recirculation pump circulates the concentrated brine through the Crystallizer Heater, where heat is transferred through the tubes. The hydrostatic head from the level in the Crystallizer Vapor Body suppresses

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boiling in the tubes. This prevents scaling that may occur if dry spots form on the heater tubes (which can be the case if boiling occurs in the tubes).

Brine entering the Crystallizer Vapor Body from the heater flash boils and releases heat in the form of water vapour. The concentrated brine collects in the vapour body and is re-circulated through the heater again. As the evaporation process continues, the concentration of the brine contained in the vapour body increases. As the concentration increases, the solution becomes supersaturated, and salts precipitate from solution resulting in a brine slurry.

Antifoam can be added to the Crystallizer on an as-needed basis to ensure that no liquid is carried over through the mist eliminators into the Crystallizer First Stage Fan during upset conditions.

Slurry from the Crystallizer is removed from the vapour body and is pumped through a recirculation loop to the crystallizer Centrifuges by the Slurry Pump. The feed flow to each centrifuge is controlled to maintain the proper slurry density, ^~25 wt% suspended solids, in the recirculating brine. The slurry is pumped from the vapour body, and a slipstream is diverted to each centrifuge for dewatering while the remaining portion recirculates back to the Crystallizer. This recirculating slurry highway is utilized to maintain a relatively high fluid velocity to avoid any solids settling and plugging in the piping.

The centrifuges process the Crystallizer product slurry. The resultant wet-cake is discharged for on-site disposal. The centrate is sent to the Centrate Tank and returned to the Crystallizer.

Make-up steam can be added as necessary but is normally only needed during start-up.

Figure 14-4 is the block flow diagram of the proposed Process Water Treatment Plant

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Source: NioCorp, 2019

Figure 14-4: Process Water Treatment Plant Block Flow Diagram

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14.5.2   Process Water

Process water will be produced at the Water Treatment Plant. Plant process water will be required in the Hydromet Plant, Acid Plant, HCI Regeneration Plant, Paste Backfill Plant and the Pyromet Plant. Additional treated water will be required for both the Mine, as well as for site potable and fire water systems.

The vast majority of process water will be required in the Hydromet Plant. The Paste Backfill Plant will utilize RO permeate for backfill, as will the mine for underground operations. The remaining plants identified will require small quantities of make-up water primarily for cooling and chilling purposes.

Mineral Processing Plant

The water requirements for the Mineral Processing plant are minimal. Plant water will be available for use during the cleanup.

Hydrometallurgical Plant

The water requirement for the Hydrometallurgical Plant is to provide dilution, make up and wash water to various sections of the plant. Table 14-12 provides a summary of the water requirement.

Table 14-12: Summary of Hydrometallurgical Process Water Requirement

HCI Leach	542.5 t/d
Water Leach	2,914 t/d
Nb Precipitation	1,713.7 t/d
Nb Caustic Leach	63.5 t/d
Sc Precipitation and Refining	24.0 t/d
Ti Precipitation	20.9 t/d
Total	5,272.9 t/d
gpm	967

Source: Tetra Tech, 2017

Pyrometallurgical Plant

The water requirements (Table 14-13) for the Pyrometallurgical Plant provide make up water to supply the FeNb Furnace cooling systems and the FeNb Furnace Pelletizer basin for cooling and pelletizing the FeNb product.

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Table 14-13: Pyrometallurgical Water Requirements

Water Item	Units	Value
Furnace Cooling Water Flowrate	m3/h	64.7
Cooling Water Flowrate Addition for Pelletizing	m3/h	15.1
Water Volume Required	m3/tap	68.5
Steam Produced	%	20
Steam Flowrate	m3/min	0.05
Make-up water Required	m3/min	0.05

Source: Tetra Tech, 2017

14.5.3   Fire Water

The firewater system will be comprised of two 225,000 gal insulated fire water tanks and two independent fire water pumps capable of delivering 2,000 gpm for a minimum period of four hours. The primary pump will be electrically driven while the backup pump will be diesel powered. A fire water distribution system will be installed throughout the site. Dry and wet sprinkler systems, hydrants, hose reels and fire extinguishers will be utilized per the design.

All infrastructure facilities on the surface, except for the gate house, will include fire suppression systems. Process building fire suppression systems will include wet sprinklers in all office spaces and control rooms. Dry sprinkler systems will be utilized in the hydrometallurgical buildings within specified high hazard areas. The remaining open process/factory areas of these two process facilities, as well as the open areas of the mineral processing building, will utilize fire hose protection from outside hydrants, as well as interior located fire hose reels.

14.5.4   Potable Water

Potable water will be supplied from three possible available sources at an operational flow rate of 3500 gpm to dedicated potable water tankage. These possible sources with their expected flow rates include; a supply line furnished by the Tecumseh Board of Public Works (2,000 gpm), a well and supply line from the Landowner 1 property (500 gpm), and two (2) wells and a supply line from the Landowner 2 property (1,500 gpm). Potable water will be distributed to all site facilities via a dedicated pumping system at 50 psig pressure. The nominal flow rate will be 100 gpm for the entire facility, with a peak flow rate of 750 gpm during shower usage.

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15. INFRASTRUCTURE

Figure 15-1 shows a site plan layout for the Elk Creek Project.

Source: Nordmin, 2019

Figure 15-1: Elk Creek Project Site Plan Layout

15.1  Electrical Power

15.1.1   Electrical Power Line & Substation

The local power utility (Omaha Public Power District) will provide power to the site. This will require approximately 29 km (18 miles) of new transmission line be installed by the utility to provide power to the Project site main sub-station to meet the required power demand, which is estimated at a peak of 30 MW. The local power utility will also design and install the main substation that will be owned and maintained by the utility. This infrastructure will be paid back through rate charges on electrical usage.

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15.1.2  Electrical Power Distribution - Plant and Facilities

The main substation will feed the site distribution substation with 44 kV. A 44 kV pole line will be constructed on the Project site to supply main power throughout the site and to the mine sub yard. In addition, this substation will include two 20/25 MVA transformers to provide 13.8 kV for distribution through the above ground facilities with approximately 1,100 m (3,610 ft) of power cables in vaults, and approximately 1,600 m (5,250 ft) of overhead lines.

15.1.3  Electrical Power Distribution - Underground

The underground electrical distribution will be fed from both the production and ventilation shafts, at 13.8 kV. Duplex fused disconnect switches will be present at several levels to allow power to be selected from either 13.8 kV feeder, providing redundancy. Power for utilization is accomplished through portable mine power centers, located at each production level. The duplex fused switches are not on every level but are distributed to adjacent levels through medium voltage junction boxes and boreholes.

15.1.4  Emergency Power Generation

Independent emergency power generation at the hoist house and ventilation shaft switchgear will be provided for back-up generation for surface infrastructure. Ventilation and hoisting are all powered from the surface, and thus, no emergency power is fed to the underground electrical distribution. Emergency power generation for the hoisting and ventilation systems will be supplied with from two diesel-powered generators, one at the hoist house and one at the ventilation shaft.

15.2  Natural Gas

15.2.1  Natural Gas Pipeline to Site

Natural gas, to be used throughout the Elk Creek during the construction and operation phases of the project, will be brought to the site via pipeline from the local utility company. NioCorp has a natural gas transportation contract with Tallgrass Energy, which operates the Rockies Express (REX) pipeline. Tallgrass will construct a 45 km (28 mile) gas pipeline lateral from the main REX pipeline system in Kansas to the project site. The lateral will be sized to provide a minimum of 27.5 dekatherms of gas per day. Natural gas will be distributed to all on-site facilities utilizing buried height density polyethylene (HDPE) natural gas distribution pipe. Natural gas piping above ground and located inside of facilities will consist predominately of carbon steel pipe. Maximum on-site pipeline distribution pressure will be 100 psig. Natural gas will be used for facility heating, water heating, and for natural gas-fired process equipment.

15.2.2  Natural Gas Distribution on Site

Natural gas will be distributed to all on-site facilities utilizing HDPE natural gas distribution pipe. Natural gas piping located inside of facilities will consist predominately of carbon steel pipe. Maximum on-site pipeline distribution pressure will be 100 psig. Natural gas will be used for facility heating, water heating, and for natural gas-fired process equipment.

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15.3  Plant Water

15.3.1   Water Treatment Plant

Water used for all on-site for all process needs and activities will be supplied from mine dewatering activities, local groundwater wells and from a local water utility (Tecumseh Board of Public Works). Mine water will be pumped to the Water Treatment Plant (WTP) that will produce approximately 2,908 gpm of treated water. Approximately 2,154 gpm of water will be produced from the Reverse Osmosis and Evaporation/Crystallization units, and 754 gpm of water will be produced from the Cooling Tower Makeup (CTMU) system.

The Water Treatment System is designed to reduce the hardness, metals, and dissolved solids of the process wastewater, cooling tower blowdown, well/utility and mine water streams. The system consists of precipitation softening, clarification, pH adjustment, multimedia filtration (MMF), and reverse osmosis (RO). Concentrated brine from the RO system will be sent to a thermal evaporator and crystallizer to produce a salt cake for disposal with the distillate being returned and combined with RO permeate for reuse.

The Process Water Treatment System includes the following major equipment units:

12. Process Water Influent Equalization Tank

13. Softening Reactor

14. Clarifiers

15. pH Adjustment Reactor

16. Multimedia Filters

17. Reverse Osmosis Units

18. Sludge Holding Tank

19. Filter Presses (shared with CTMU system)

20. Evaporator/Crystallizer System

21. Crystallizer Solids Dewatering System

22. Chemical Feed Systems

The following Table 15-1 was used as the design basis.

Table 15-1: Design Requirements

Parameter	Quantity (gpm)	Notes
Plant Source Water	3,375	From mine dewatering and water wells
Pyromet Feed Make-up (2 points)	5	From RO Units
	5	From RO Units
Hydromet Feed	200	From RO Units (Additional Hydromet Feed Water from the Acid Plant (450 gpm) and Potable Water Wells (1575

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		gpm) are untreated)
Acid Plant	40	From RO Units
Hydromet Cooling Tower	1,415	From RO Units & CTMU System
Pyromet Cooling Tower	1,158	From RO Units & CTMU System
Hydromet Return Water	750	Constituents: Na, Cl, Ca, SO4 and Fe

Following is a summary description of the proposed Water Treatment Plant.

15.3.1.1  Flow Equalization

Process wastewater and underground mine water from NioCorp will be pumped into an equalization tank. Cooling Tower Blow Down (CTBD) will also be added to this tank since it will contain elevated total dissolved solids and hardness. The tank will also receive intermittent return flows from MMF backwash and the sludge dewatering system. The equalization tank will allow for storage during a shutdown and to sustain consistent flow to the system. The combined process wastewater and mine water will be pumped from the equalization to the softening reactor at a controlled rate. In the case of a system shutdown, it was assumed there would be enough storage capacity to accommodate reduced or no flow of mine water to the treatment system.

15.3.1.2  Softening and Clarification

The combined streams will enter a Turbomix^® softening reactor where chemicals will be added for precipitation softening. The advantage of the Turbomix design is that it promotes precipitation/crystallization of the dissolved particles to maximize their size and density. This results in faster settling rates, improved sludge handling characteristics, and improved sludge thickening and dewatering rates. To enhance the crystallization reaction kinetics and to maximize the density of the settled sludge, a portion of the precipitated sludge collected in the downstream clarification process will be recycled back to the Turbomix draft tube. Hydrated lime and soda ash will be fed to the Turbomix based on the flow rate, hardness, and alkalinity of the incoming water. A coagulant also will be added.

The Turbomix reactor will overflow to two flocculating clarifiers to provide redundancy to allow one unit to be taken down for short durations for maintenance. The polymer will be added to the clarifier center well to promote flocculant growth and improve the settling characteristics of the precipitated solids. A rotating rake assembly including two long rake arms will move the settled solids to a center sludge discharge sump. The clarifier rake drive will be equipped with a high torque alarm and an automatic rake lift to raise the rotating rake mechanism should a torque overload condition occur.

The settled sludge will be withdrawn from the bottom of the clarifiers continuously by underflow pumps. The settled softening sludge is expected to have a solids concentration of close to 10%.

The clarifier effluent will be collected in a launder and will exit the clarifier through a drop box and be conveyed to the pH adjustment reactor tank ahead of the multimedia filters. The pH will be reduced to near neutral. This will allow any residual aluminum to precipitate for subsequent removal in the Multimedia Filter (MMF). An oxidant will also be added to this tank for ammonia

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removal. Water will be pumped from this tank to the MMF to further reduce the suspended solids prior to RO.

15.3.1.3 Multimedia Filtration

The effluent from the pH Adjustment Reactor (MMF Feed tank) is pumped to the MMF System. The goal of the filtration system is to reduce the inlet suspended solids concentration prior to RO. The vessels contain three separate layers of filtration media and a gravel support bed. The gravel supports the top three active filter layers consisting of anthracite, sand and fine garnet. This layered media profile provides a high sediment holding capacity as compared to conventional dual media/sand filters. The larger incoming particles are trapped on the upper layer of the media allowing the smaller particles to continue through the bed where they are trapped in the lower layers, producing a high-quality effluent. A filter aid will be added to the inlet of the MMF to enhance solids-liquid separation process and achieve deep bed filtration versus conventional surface filtration.

During operation, the softened water enters the multimedia filter vessel under pressure at the top and is distributed uniformly over the top layer of the media bed. After passing through the media bed, the filtered service water exits the vessel through the under-drain assembly at the bottom. As the water flows through the media bed, the suspended solids and turbidity present in the feed water will be removed. The filter media bed slowly exhausts from top to bottom. When the turbidity and/or the differential pressure from the media bed approaches a predetermined set point, the media bed is exhausted and is subjected to cleaning/backwash cycle.

15.3.1.4   Reverse Osmosis (RO) System

Filtered water from the MMF is collected in the RO Feed Tank and will be pressurized through a single pass RO system for removal of total dissolved solids. A small portion of the filtered water will be utilized for Multimedia Filter backwash purposes.

The RO process separates dissolved contaminants from the feed water by passing through a semipermeable thin film composite membrane. These membranes remove 95 ^~ 99% of the dissolved solids present in the feed water and essentially perform a complete removal of all particulate matter.

During operation, the filtered water from the RO feed tank is pumped to the cartridge filter vessels. The water pressure forces the feed water through the filter elements while leaving any residual impurities behind on the filter element surface. The cartridge slowly exhausts, and when they are clogged with impurities, the pressure drop across the cartridge filter system exceeds the desired limit, and the dirty filter elements are taken out of service for replacement. An antiscalant will be added at the RO cartridge filter inlet to prevent any potential scaling issues across the downstream RO system.

The filtered water from the cartridge filter is then pressurized using the RO booster pump and is fed to the first stage membranes in the RO system. The concentrate from the RO system is routed to the RO Reject Tank prior to being discharged. The concentrate will be sent to the evaporation/crystallization process for further concentration. Permeate stream from the system is collected in the RO Product tank where it blends with the distillate from the evaporator and crystallizer and is pumped to the Hydromet process, cooling tower and other water users.

Over a period of time, the RO membrane elements will be subjected to potential fouling by suspended material or sparingly soluble material that may be present in the feed water. Upon an

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increase of the feed pressure or decline of permeate quantity/quality, the RO system will be taken offline, and the membranes will be cleaned.

15.3.1.5  Cooling Tower Makeup System (CTMU)

The CTMU Treatment System is designed to reduce iron and manganese in the groundwater supply for cooling tower makeup. Limited data was available on the groundwater; data from a nearby farmer’s groundwater well shows that manganese is present at 0.4 mg/L. Cooling tower suppliers typically recommend that manganese be reduced to <0.05 mg/L to prevent deposition and fouling on the cooling tower fill and cooling loop systems. Based on the final water balance, there will be excess RO permeate available to blend with the groundwater (40:60 blend). Based on the projected blended quality, it is expected that the cooling towers can be operated at up to seven cycles of concentration.

The CTMU system will consist of a separate second treatment system to reduce iron and manganese using a filter media for this process. The filter backwash from the CTMU system will be combined with the Process Water Treatment system.

The CTMU Treatment System includes the following major equipment units and redundancy:

3. Manganese Removal Media Filters

4. Sodium Hypochlorite Feed System

The untreated well water will flow through Manganese Removal filters. It is assumed that the well water pressure is adequate for feeding the filters without re-pumping. Normally all filters are online, except when one filter requires backwashing, where the remaining filters handle the design flow. Sodium hypochlorite is injected in-line prior to the filters. The filters operate like the MMF units described previously. Upon high differential pressure, each filter is taken off-line for backwashing. Backwash water will be pumped from the downstream CTMU tank to the filters. Dirty backwash water will be conveyed to the PW sump and sent through the sludge handling system.

The treated water will be collected in a CTMU Tank and pumped to the cooling towers. Excess RO permeate/distillate from the PW treatment system will also be used as CTMU when available. Chemical storage systems will be shared between the CTMU and PW treatment systems.

15.3.1.6  Sludge Handling

Sludge from the PW will be collected in a sludge storage tank.

Intermittently the sludge from the storage tank will be pumped to the filter presses for dewatering. Pumps are provided to feed the filter presses. Filter press filtrate will flow by gravity to the building sump and then pumped to the Process Water Equalization Tank using sump pumps. The building sump will also receive filter backwash from the MMFs.

15.3.1.7  Evaporation and Crystallization System

Evaporator Brine Flow

The RO concentrate will be processed through an Evaporator/Crystallizer system to produce a salt cake for disposal. The RO concentrate contains a certain amount of alkalinity. In order to prevent calcium carbonate fouling of the Evaporator heat exchanger, it is important to eliminate all the carbonate alkalinity in the feed stream. This is accomplished in a three-stage process: feed acidification with sulphuric acid, feed preheating and feed deaeration/decarbonation. Feed

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acidification (via metered sulphuric acid addition) is performed within the Evaporator Feed Tank. The sulphuric acid converts the carbonate and bicarbonate ions to CO₂. The CO₂ is subsequently stripped out of the feed stream in the Feed Deaerator following heat recovery in the Feed Preheater. Brine from the Evaporator Feed Tank is pumped to the Feed Preheater where the temperature is increased by exchanging heat with the Evaporator and Crystallizer condensate. The feed then enters the Feed Deaerator where vapour and non-condensable gasses (NCGs) vented from the shell side of the Evaporator, heats the feed and allows for the release of CO₂ to the atmosphere. The feed then enters the Evaporator.

The purpose of the Evaporator is to remove the majority of the water in the most energy and cost-efficient manner prior to the crystallization system. The feed flow enters the vapour body and is pumped up through the center of the heater via Evaporator Recirculation Pump. The recirculating brine stream is introduced into a vertical heat exchanger tube bundle utilizing Veolia’s double distributor plate design. The brine falls down the inside of the heater tubes where it is heated by vapours condensing on the outside of the tubes, causing the brine to boil. The concentrated brine gathers in the vapour body below the heater, where it is recirculated again.

Antifoam can be added to the Evaporator on an as needed basis to ensure that no liquid is carried over through the mist eliminators. Caustic is added to the Evaporator to maintain the pH between 8.0 and 8.5 to ensure the system will not be susceptible to corrosion.

The concentrated brine leaves the Evaporator via a purge line off the discharge of the Evaporator Recirculation Pump and is pumped to the Crystallizer Feed Tank for further concentration.

Low-pressure steam is created by the auxiliary boiler. This steam is utilized for start-up purposes and as supplemental heat for the system when required.

15.3.1.8  Crystallizer Brine Flow

The concentrated brine from the Evaporator is pumped to the Crystallizer Feed Tank. Caustic is again added to the system at the Crystallizer Feed Tank. Caustic is needed at this point to make up for metal hydroxides that precipitate as the brine is concentrated. The target pH in the Crystallizer is 8.0-8.5. The Crystallizer is a forced circulation unit meaning the recirculation pump circulates the concentrated brine through the Crystallizer Heater, where heat is transferred through the tubes. The hydrostatic head from the level in the Crystallizer Vapor Body suppresses boiling in the tubes. This prevents scaling that may occur if dry spots form on the heater tubes (which can be the case if boiling occurs in the tubes).

Brine entering the Crystallizer Vapor Body from the heater flash boils and releases heat in the form of water vapour. The concentrated brine collects in the vapour body and is re-circulated through the heater again. As the evaporation process continues, the concentration of the brine contained in the vapour body increases. As the concentration increases, the solution becomes supersaturated, and salts precipitate from solution resulting in a brine slurry.

Antifoam can be added to the Crystallizer on an as-needed basis to ensure that no liquid is carried over through the mist eliminators into the Crystallizer First Stage Fan during upset conditions.

Slurry from the Crystallizer is removed from the vapour body and is pumped through a recirculation loop to the Crystallizer Centrifuges by the Slurry Pump. The feed flow to each centrifuge is controlled to maintain the proper slurry density, ~25 wt% suspended solids, in the recirculating brine. The slurry is pumped from the vapour body, and a slipstream is diverted to each centrifuge for dewatering while the remaining portion recirculates back to the Crystallizer.

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This recirculating slurry highway is utilized to maintain a relatively high fluid velocity to avoid any solids settling and plugging in the piping.

The centrifuges process the Crystallizer product slurry. The resultant wet-cake is discharged for on-site disposal. The centrate is sent to the Centrate Tank and returned to the Crystallizer.

Make-up steam can be added as necessary but is normally only needed during start-up.

Figure 15-2 is the block flow diagram of the proposed Process Water Treatment Plant.

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Source: NioCorp, 2019

Figure 15-2: Process Water Treatment Plant Block Flow Diagram

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15.3.2  Process Water

Process water will be produced at the Water Treatment Plant. Plant process water will be required in the Hydromet Plant, Acid Plant, HCI Regeneration Plant, Paste Backfill Plant and the Pyromet Plant. Additional treated water will be required for both the Mine, as well as for site potable and fire water systems.

The vast majority of process water will be required in the Hydromet Plant. The Paste Backfill Plant will utilize RO permeate for backfill, as will the mine for underground operations. The remaining plants identified will require small quantities of make-up water primarily for cooling and chilling purposes.

Mineral Processing Plant

The water requirements for the Mineral Processing plant are minimal. Plant water will be available for use during the cleanup.

Hydrometallurgical Plant

The water requirement for the Hydrometallurgical Plant is to provide dilution, make up and wash water to various sections of the plant. Table 15-2 provides a summary of the water requirement.

Table 15-2: Summary of Hydrometallurgical Process Water Requirement

HCI Leach	542.5 t/d
Water Leach	2,914 t/d
Nb Precipitation	1,713.7 t/d
Nb Caustic Leach	63.5 t/d
Sc Precipitation and Refining	24.0 t/d
Ti Precipitation	20.9 t/d
Total	5,272.9 t/d
gpm	967

Source: Tetra Tech, 2017

Pyrometallurgical Plant

The water requirements (Table 15-3) for the Pyrometallurgical Plant provide make up water to supply the FeNb Furnace cooling systems and the FeNb Furnace Pelletizer basin for cooling and pelletizing the FeNb product.

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Table 15-3: Pyrometallurgical Water Requirements

Water Item	Units	Value
Furnace Cooling Water Flowrate	m3/h	64.7
Cooling Water Flowrate Addition for Pelletizing	m3/h	15.1
Water Volume Required	m3/tap	68.5
Steam Produced	%	20
Steam Flowrate	m3/min	0.05
Make-up water Required	m3/min	0.05

Source: Tetra Tech, 2017

15.3.3  Fire Water

The firewater system will be comprised of two 225,000 gal insulated fire water tanks and two independent fire water pumps capable of delivering 2,000 gpm for a minimum period of four hours. The primary pump will be electrically driven while the backup pump will be diesel powered. A fire water distribution system will be installed throughout the site. Dry and wet sprinkler systems, hydrants, hose reels and fire extinguishers will be utilized per the design.

All infrastructure facilities on the surface, except for the gate house, will include fire suppression systems. Process building fire suppression systems will include wet sprinklers in all office spaces and control rooms. Dry sprinkler systems will be utilized in the hydrometallurgical buildings within specified high hazard areas. The remaining open process/factory areas of these two process facilities, as well as the open areas of the mineral processing building, will utilize fire hose protection from outside hydrants, as well as interior located fire hose reels.

15.3.4  Potable Water

Potable water will be supplied from three possible available sources at an operational flow rate of 3500 gpm to dedicated potable water tankage. These possible sources with their expected flow rates include; a supply line furnished by the Tecumseh Board of Public Works (2,000 gpm), a well and supply line from the Landowner 1 property (500 gpm), and two (2) wells and a supply line from the Landowner 2 property (1,500 gpm). Potable water will be distributed to all site facilities via a dedicated pumping system at 50 psig pressure. The nominal flow rate will be 100 gpm for the entire facility, with a peak flow rate of 750 gpm during shower usage.

15.4  Roads

15.4.1  Main Access Road to Site

The primary access to the site will be from County Road 721. Access into the site will be controlled by security personnel. The site access road will be leading to the main access points to the mine, the administration building and the primary traffic destinations on the site.

15.4.2  Secondary Site Access Roads

A second, emergency access to the site will be connecting to Nebraska State Route 50. The entrance to the emergency access road will be secured with a locked gate.

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15.4.3  Secondary Site Roads (to tailings, etc.)

Secondary roads on site include haul roads connecting the plant site to TSF cells and light vehicle access roads connecting infrastructure throughout the site. Haul traffic is expected to include 40-tonne haul trucks delivering tailings and water treatment system residual salt to the active TSF and salt cells and support equipment for the haul fleet. Light vehicles include light-duty pickups and service vehicles supporting infrastructure.

Haul Roads

Haul roads are required to provide access between the plant site and TSF and salt cells. A haul road will provide access to Plant Site TSF Cells 2 and 3 and Salt Management Cell 1 (SMC-1), as well as the Area 7 TSF and salt management Cell 2 (SMC-2). The Area 7 haul road will require a connection between Highway 50 and the TSF area. Improvements to the public roadway may be required based on the haul fleet size delivering tailings and salt to Area 7. Highway-compliant tractor-trailer trucks may be used instead of off-road trucks.

Haul roads are designed to allow trucks to pass safely. The haul trucks assumed for this study are 40 ton articulated off-road trucks. The width of a truck is approximately 4 m wide. To provide safe passage for two-way traffic, the suggested width of the travel way (driving surface) is 3.5 times the width of the truck or 14 m. Haul roads are shown with a width of 20 m, allowing for up to 6 m for safety berms. The road widths and berm placement will be determined during the final engineering design. Much of the roadway may not require a safety berm due to height, but a running width of 20 m is used for estimating and preliminary design purposes. Haul traffic speed is generally slow (30 to 40 km/h); the design shown does not incorporate engineering controls for a specific design speed, but rather are shown for estimating and preliminary design purposes only.

Light Vehicle Access Roads

Light vehicle access roads are located throughout the site. They provide access to infrastructure such as ponds, embankment crest and toe fills.

Expected traffic on light duty roads includes light-duty pickup trucks, maintenance equipment, and the occasional haul truck. Light vehicle roads assume occasional use, single-lane traffic with areas to safely pull out of the traffic lane should vehicles meet. A typical light-duty vehicle is approximately 3 m wide. Road widths are designed at 6 m in width.

Speeds are expected to be slow (20 to 30 km/h); the design shown does not incorporate engineering controls for a specific design speed, but rather are shown for estimating and preliminary design purposes only.

Construction

Geotechnical information for soils underlying road alignments is not available at this time. The construction of the roadways assumes similar construction practices as defined for the TSF embankment construction, including removal of 1 m (+/-) of topsoil, replacement with suitable compacted sub-grade fill, and the provision of structural support for traffic with a durable gravel surface. Geotextile fabric will be installed at the base of the gravel layer to provide stability. A minimum of 0.5 m of compacted gravel is assumed for the driving surface.

All roadways will be designed to promote drainage off of the driving surface. This requires that the roadways be elevated slightly above the surrounding ground elevations and crowned, and/or a drainage ditch be provided as needed in areas of elevation transition from cut to fill. In areas where berms are required, notches in the berms should be provided at regular intervals to allow

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stormwater to discharge off of the roadways. In areas where safety berms are not required, shoulder slopes should not exceed 3:1 (horizontal to vertical), and 4:1 is preferred to reduce the chance of a vehicle rollover should they divert from the roadway.

15.5 Tailing Storage and Associated Facilities

Preliminary investigations performed by SRK included a comparison of potential TSF sites for both slurry and filtered (dewatered) tailings disposal options. This comparison considered potential engineering, strategic, permitting and closure issues, including:

● Engineering: Containment area, required reclaim for water balance on tailings impoundment, relative embankment heights, distance to plant, pumping head for slurry (plant to impoundment) and reclaim water (impoundment to plant), upstream stormwater management, major road crossings, potential residential relocations, and potential road relocations.

● Strategic: Proximity to major roadways, churches and cemeteries, visual embankment heights, and property ownership.

● Permitting: Major drainage crossings and major road encroachment.

● Closure: Closure cover areas and volumes, seepage potential, and mass stability.

Of eight potential sites, Area 7 and Area 1 ranked first and second for both slurried and filtered tailings, respectively. This evaluation included the development and implementation of a preliminary foundation characterization plan for both Area 1 and Area 7 and development of preliminary water balance spreadsheets for both slurried and filtered tailings options for both sites.

Following the development of the 2015 PEA, the decision was made to only generate dry tailings, by calcining and filtration processes, and a more detailed foundation characterization investigation was performed for Area 7. Revised planning indicated that a significant portion of the filtered tailings would be used for underground backfill operations, limiting the total tailings tonnage to be disposed of in the TSF cells to around 1,070 dry t/d for a life of 36 years (from the original plan for 4,930 t/d for 30 years).

This significant decrease in deposition rate, as well as the finding that the calcined tailings material will be a dry “clinker” with a sandy gravel or gravelly sand gradation (i.e., well drained), led to NioCorp’s decision to evaluate the plant site (refer to TSF Cells 1, 2 and 3 in Figure 15-3) as feasible tailings storage and stormwater management locations for the first 19 years of operations, with the following significant advantages:

● No access roads or conveyors crossing Elk Creek.

● Shorter distance from Plant Area for tailings transport and reclaim water management.

● Reduction in stormwater management.

● Consolidation of disturbance into a much smaller area (without Area 7).

The plant area was therefore considered the best option for the first 19 years of management and storage of dry tailings (in three, State-approved “solid waste” disposal facilities or cells), and management of precipitation contacting the tailings via runoff and infiltration in separate double-lined leachate collection ponds. Once the Plant Site TSF cells are full, a new facility will be constructed at Area 7.

The feasibility design incorporates the following parameters and details:

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i. Topography: Feasibility design has been performed using 1 m contoured topography.

ii. Feasibility Design: Feasibility design of the TSF Cells is based on dam safety regulations, solid waste regulations (including tailings placement/compaction/interim covering), leachate water management regulations, and radioactive licensing regulations, all discussed in Section 17. The design is intended to demonstrate compliance with Nebraska industrial solid waste regulations for design, operation and closure and is based on a meeting held between NioCorp, SRK and the Nebraska Department of Environment and Energy (NDEE).

iii. Embankment Cross-Section: All TSF and LCP embankment sections will incorporate a 20 m crest width and 3 (horizontal) to 1 (vertical) sideslopes as shown in Figure 15-3 and Figure 15-4. Revegetation of embankment crests and downstream sideslopes will be provided for erosion protection, immediately after construction completion of each embankment.

iv. Leachate Collection Ponds (LCPs): Each of the three LCPs will be utilized for management of precipitation runoff and drainage from the tailings’ solids. The ponds will be lined with two layers of geomembrane liner (80-mil primary and 60-mil secondary), sandwiching a permeable spacer that allows evacuation of all leakage through the primary liner to be collected in a lined sump area, or leakage collection and recovery system, and pumped back into each leachate collection pond, thereby providing a means of long-term leakage control (refer to Figure 15-7).

v. Tailings Production Rate: The tailings production rate is an average of 2,460 dry t/d consisting of 1) 825 dry t/d of water leach residue tailings; 2) 1,588 dry t/d of calcined excess oxide tailings, and 3) 46 t/d of slag. Of this, an average of 1,390 dry t/d will be placed into the mine backfill (Section 13.5.4), and 1,070 dry t/d will be placed in the TSF. Testing performed on the excess, and insoluble oxides indicate that a loose (non-compacted) dry density of 1.6 t/m³ will be achieved without compaction, and that placement and spreading of the dry tailings will increase the density to between 1.7 and 1.8 t/m³.

vi. Growth Media Salvage: A minimum of 1 m of subbase soils will be removed prior to construction of each TSF cell and stockpiled.

vii. TSF Area, Storage and Time Characteristics: The TSF system includes four TSF cells, Plant Site Cell 1, Cell 2 and Cell 3 and Area 7 Cell 1. Table 15-4 provides a summary of TSF cell footprint areas, storage characteristics and time periods.

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Table 15-4: TSF Area, Storage and Time Characteristics

Cell No.	Approximate Footprint Area (Ha)	Storage Capacity (@ dry density of 1.7 t/ m3) (Mt)	Time Period (years after commissioning)
Plant Area TSF
1	8.8	1.5	3
2	14.1	3.1	4-10
3	13.9	3.1	10-18
Area 7 TSF
1	26.6	6.7	19-38

Source: SRK, 2019

viii. Leachate Collection Pond (LCP) Storage Characteristics: The LCP system includes three ponds; one for Plant Site Cell 1 (LCP-1), one for Plant Site Cells 2 and 3 (LCP-2) and one for Area 7 Cell 1 (Area 7-LCP). Table 15-5 provides a summary of LCP footprint areas and leachate and stormwater storage characteristics.

Table 15-5: LCP Area, Storage and Time Characteristics

LCP	Approximate Footprint Area (Ha)	Total Storage Capacity (1) (ns)
LCP-1	0.9	29,200
LCP-2	1.5	61,600
Area 7 LCP	4.3	240,000

Source: SRK, 2019

(1)   For operating and 100-year stormwater conditions.

ix. Structural Embankment Foundation Preparation: Foundation preparation for all TSF and LCP embankments will incorporate removal of a minimum of 0.5 m of native soils, and re-compaction in layers to form a non-settling structure as shown in Figure 15-5 and Figure 15-6.

x. Embankment Compaction: All TSF and LCP embankments will be constructed using soil borrowed from within the respective TSF basins and compacted in layers to form a non-settling structure, as shown in Figure 15-6 and Figure 15-7.

xi. Embankment Raises: All tailings embankments will be constructed to completion before each cell is commissioned as shown in Table 15-6, thereby eliminating the need for raising extension of the TSF liner and embankment drainage elements at any stage. This assists in preventing the facility from being affected by potential liquefaction of the tailings’ solids under seismic loads.

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xii. TSF Liner and Above-Liner Drainage System: The components of the facility liner/drainage systems are described below and shown in Figure 15-4 through Figure 15-7.

a. The TSF basins and inside embankment sideslopes will incorporate:

i. A minimum of 0.6 m of glacial till, amended if necessary, with bentonite, and compacted in layers to result in hydraulic conductivity of less than or equal to 1×10^-7 cm/s;

ii. An 80-mil high-density polyethylene (HDPE) geosynthetic liner placed over the prepared basin subgrade and inside embankment sideslopes; and

iii. For Plant Site TSF Cell 1, an above-liner centralized drain (i.e., from north to south) directing drain flows into an above-liner, double-lined, leak-detected sump that facilitates pumping into LCP 1.

iv. For Plant Site TSF Cells 2 and 3 and Area 7 TSF Cell 1, above-liner embankment toe drains (along entire inside perimeter), as well as centralized drains that gravitate drain flows into above-liner, double-lined, leak-detected sumps that facilitate pumping into LCP 2 and Area 7 LCP, respectively.

v. Typical drain sections are provided in Figure 15-5 for the centralized drain (Plant Site Cells 1, 2 and 3), and the perimeter inside toe drains (Plant Site Cells 2 and 3 and Area 7 Cell 1).

b. Lining for LCP Basins and Inside Sideslopes: LCP basins and inside embankment sideslopes, as well as the leak collection and recovery system (LCRS) sumps within the TSF cells, will incorporate:

i. A minimum of 0.6 m of glacial till, amended if necessary, with bentonite, and compacted in layers to result in hydraulic conductivity of less than or equal to 1×10^-7 cm/s;

ii. A 60-mil HDPE secondary liner incorporating an Agru DrainLiner^® system or geonet layer plus smooth geomembrane;

iii. An 80-mil HDPE primary liner;

The DrainLiner or geonet layer will gravitate into an LCRS sump that facilitates pumping of collected seepage water back into the LCPs via a submersible pump and riser pipeline arrangement. As shown in Figure 15-7, the riser pipeline will be contained in a “port” pipeline installed between the two liners. The LCRS sumps are gravel-filled collection areas between the primary and secondary liners, with a horizontal section of perforated pipe within the gravel for pumping.

xiii. Tailings Solids Transportation and Deposition: A cost trade-off study was performed to compare trucking to conveying costs and trucking as selected as the preferred option. Tailings solids will be trucked from the process plant directly to each planned deposition location at the TSF Cells, dumped, spread and compacted using a bulldozer, and graded to slope to facilitate control of surface water. Tailings will be placed in sections in the cells starting at the high point in the base grading and working toward the sumps. Each cell will be closed in phases every 3 to 4 years, once the full depth of tailings has been achieved in each section, as described in Section 17.4.

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a. For Plant Site TSF Cell 1 and Area 7 TSF Cell 1, tailings placement will be performed from south to north.

b. For Plant Site TSF Cell 2 and 3, tailings placement will be performed from north to south.

xiv. TSF Stage-Area-Capacity Data: Stage-capacity data is provided in Table 15-6 and summarizes TSF elevation, area, cumulative volume, capacity in tonnes, time in years and rate-of-rise in metres per year.

xv. Surface Water Management: Surface water management comprises both precipitation-induced contact water and non-contact water.

a. Surface water contacting the tailings will be managed via dedicated pump arrangements for all three cells that comprise of slotted HDPE riser pipes located above the liner system at the impoundment topographic low point, on the embankment inside slopes. Submersible pumps will be used to pump collected water into the LCPs. The submersible pumps will be maintained above the current tailings elevation at all times. The locations of the riser pipes are shown in Figure 15-6. Any infiltrating surface water will be collected in the TSF above-liner drainage system.

b. The average rainfall is shown in Table 15-7. Based on the average monthly precipitation, pump back from the TSF underdrainage system has been estimated:

▪ For Plant Site TSF Cell 1 at an average of 36 gpm varying between 64 gpm and 11 gpm,

▪ For Plant Site TSF Cell 2 at an average of 58 gpm varying between 104 gpm and 18 gpm, and

▪ For Plant Site TSF Cell 3 at an average of 57 gpm varying between 102 gpm and 18 gpm.

▪ For Area 7 TSF Cell 1 at an average of 105 gpm varying between 180 gpm and 35 gpm.

c. Storm-related precipitation depths are provided for 25-year and 100-year, 24-hour duration storms in Table 15-7. Based on 100-year precipitation depth, the pump back requirements for the 100-year condition is estimated to require 115 gpm from Plant Site Cell 1, 190 gpm from both Plant Site Cells 2 and 3, and 373 gpm from Area 7 Cell 1.

d. Non-contact surface water will be managed via channels, spillways, and culverts as shown in Figure 15-3. Spillways are sized to pass the PMF storm event, and all other stormwater controls are designed to accommodate 100-year, 24-hour storm event precipitation.

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Table 15-6: Tailings Storage Facility Stage-Area-Capacity Data

Cells	TSF Elevation (masl)	Area (m2)	Cumulative Volume (m3)	Capacity (1) (t)	Years	Rate of Rise (m/y)
Plant Site Cell 1	349	3,343	0	0	0.0	0.0
	350	9,570	6,456	10,330	0.0	36.6
	352	26,589	42,022	67,235	0.2	10.7
	354	36,889	107,528	172,045	0.5	6.7
	356	43,353	187,734	300,375	0.8	5.7
	358	50,109	281,160	449,856	1.2	4.9
	360	57,163	388,395	621,431	1.6	4.3
	362	64,516	510,036	816,058	2.2	3.8
	364	72,168	646,682	1,034,692	2.7	3.4
	366	80,119	798,932	1,278,291	3.4	3.0
	368	88,369	967,383	1,547,813
Plant Site Cell 2	360	25,179	21,391	34,225	3.8	14.0
	362	63,468	109,856	175,770	4.2	4.4
	364	74,603	249,913	399,861	4.8	3.2
	366	81,976	406,456	650,329	5.5	2.9
	368	89,630	578,027	924,842	6.2	2.7
	370	97,567	765,189	1,224,302	7.0	2.5
	372	105,786	968,507	1,549,611	7.8	2.3
	374	114,286	1,188,544	1,901,670	8.8	2.1
	376	123,069	1,425,864	2,281,382	9.8	2.0
	378	132,133	1,681,031	2,689,649	10.8	1.8
	380	141,480	1,954,609	3,127,374
Plant Site Cell 3	351	14,534	8,924	14,279	11.5	26.5
	352	32,842	32,613	52,180	11.6	10.0
	354	68,836	139,323	222,917	12.0	3.8
	356	76,009	284,360	454,976	12.6	3.2
	358	83,004	443,339	709,343	13.3	2.9
	360	90,251	616,564	986,502	14.0	2.7
	362	97,765	804,545	1,287,272	14.8	2.5
	364	105,554	1,007,830	1,612,527	15.7	2.3
	366	113,609	1,226,960	1,963,136	16.6	2.1
	368	121,919	1,462,457	2,339,931	17.6	2.0
	370	130,486	1,714,827	2,743,724	18.7	1.8
	372	139,324	1,984,603	3,175,365
Area 7 Cell 1	348	3,936	0	0	0.0	0.0
	350	39,195	43,131	73,323	19.4	11.0

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352	78,499	160,825	273,402	19.9	4.0
354	125,046	364,370	619,428	20.8	2.3
356	155,733	645,148	1,096,752	22.0	1.7
358	167,568	968,449	1,646,363	23.3	1.5
360	179,701	1,315,718	2,236,721	24.8	1.4
362	192,133	1,687,552	2,868,839	26.4	1.3
364	204,863	2,084,549	3,543,733	28.1	1.2
366	217,892	2,507,304	4,262,416	29.9	1.1
368	231,219	2,956,414	5,025,904	31.8	1.1
370	244,844	3,432,477	5,835,211	33.8	1.0
372	258,768	3,936,089	6,691,351	35.9	0.9
373	265,842	4,198,394	7,137,270	37.0	0.9

Source: SRK, 2019. (1) Tonnes of storage is based on an assumed dry density of 1.7 t/m³.

Table 15-7: Mean Monthly Average Precipitation

Station	Mean Monthly Precipitation	Mean Monthly Pan Evaporation	Mean Monthly Lake Evaporation	Annual Potential Evapotranspiration (PET)
	Tecumseh Station’ (mm)	Sabetha Lake Station2 (mm)	Sabetha Lake	Rainwater Basin
			Station2 (mm)	Station3 (mm)
Jan	21		-	30
Feb	28	-	-	32
Mar	49	-	-	66
Apr	72	131	98	84
May	111	167	126	98
Jun	117	186	139	98
Jul	99	210	158	102
Aug	97	190	142	87
Sep	89	138	103	86
Oct	58	103	77	81
Nov	39	57	43	58
Dec	26	-	-	29
Annual	805	1182	887	851
Seven-Year Wet-Cycle Total	6,662
Seven-Year Dry-Cycle Total	4,318

1. Tecumseh station data (WRCC, DRI) is considered the most representative based on elevation and proximity to the Project site.

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2. Data from Southwest Climate and Environmental Information Collaborative (WRCC, DRI); Sabetha Lake station data is considered the most representative based on elevation and proximity to the Project site.

3. RAWS Network (DRI), ASCE Standardized Reference ET Calculations

4. 5-year average from 2009 through 2013

5. Based on Lake Evaporation as 75% of Pan Evaporation

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Source: SRK, 2019

Figure 15-3: Tailings Storage Facility Layout Showing Plant Site Cells 1, 2 and 3 and Area 7 Cell 1

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Source: SRK, 2019

Figure 15-4: Tailings and Waste Rock Storage Area Embankment Cross-Section

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Source: SRK, 2019

Figure 15-5: Tailings Storage Facility Central and Toe Drain Details

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Source: SRK, 2019

Figure 15-6: Leachate Collection Pond Embankment Cross-Section

Source: SRK, 2019

Figure 15-7: Leachate Collection Pond LCRS System

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15.6 Salt Management Cells

The crystalline salt produced as a waste product by heating and evaporating brine from the Reverse Osmosis (RO) water treatment plant will be transported by conveyor to the temporary salt staging area within the aforementioned Sprung Structure over concrete containment. The salt will then be transported by truck to the dedicated Salt Management Cells (SMC).

The feasibility design incorporates the following parameters and details:

A. Feasibility design has been performed using 1 m contoured topography.

B. The salt management cell embankment sections will incorporate a 20 m crest width and 3:1 sideslopes. Revegetation of the embankment crests and downstream sideslopes will be provided for erosion protection immediately after the construction of each embankment.

C. A minimum of 1 m of subbase soils will be removed prior to construction of the SMCs and stockpiled for future use as growth media for site closure.

D. Foundation preparation for the embankments will incorporate removal of a minimum of 0.5 m of native soils, and re-compaction in layers. The embankments will then be constructed using soil borrowed from within the active SMC footprint and compacted in layers to form a non-settling structure.

E. SMC embankments will be constructed to their ultimate configuration before each cell is commissioned, as opposed to construction in phases.

● The SMCs will incorporate, as described from the sub-base vertically upwards, the following:

● A minimum of 0.6 m of glacial till, amended if necessary, with bentonite, and compacted in layers to result in hydraulic conductivity of less than or equal to 1×10^-7cm/s;

● A 60-mil HDPE secondary liner incorporating an Agru DrainLiner® system or geonet plus smooth geomembrane; and

● An 80-mil high-density polyethylene (HDPE) primary liner.

The DrainLiner or geonet layers will route collected leakage into an LCRS sump that facilitates pumping of collected water back into the SMCs or LCPs via a submersible pump and riser pipeline arrangement.

F. The salt production rate is anticipated to be 45,250 m³ per year with a total of 1.63 million m³ required for the life of the mine. The SMC system includes two cells, SMC-1 and SMC-2, adjacent to TSF Cell 2 and Area 7 TSF Cell 1, respectively. Table 15-8 provides a summary of SMC footprint areas and storage capacities.

Table 15-8: SMC Footprint Areas and Storage Capacities

LCP	Approximate Footprint Area (Ha)	Total Storage Capacity (m³)	Time Period (year after commissioning)
SMC-1	9.7	700,000	15.5
SMC-2	10.2	930,000	36

Source: SRK, 2019

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G. Stormwater will be managed within each Salt Management Cell by spray evaporating within the open portion of each cell until the useable area is too small or the volume of stormwater collected is too great. The base of each cell will be graded to an internal sand drain, and sump from which captured runoff will be pumped to the internal evaporator system or the nearest tailings leachate collection pond. Salt will be placed in each cell (via trucks), starting upstream and progressing downstream to the internal sump location. Temporary covers will be employed to minimize the exposed salt to water. As portions of each cell reach maximum design height, they will be progressively closed using geomembranes and soil cover to minimize exposed salt and contact water. When the salt placement sequence requires filling over the internal sump, a riser housing consisting of HDPE pipe laid along the cell sideslope will be installed to provide protection to the pump and discharge lines.

15.7 Paste Backfill Plant and Underground Distribution

15.7.1 Surface Plant

The Paste Backfill Plant at the Elk Creek Mine is designed to make a paste using waste products from the Process Plant to produce a paste backfill for use as backfill material for additional stability for the underground mining operation.

The Hydromet Process Plant produces mixed oxides and leach residue waste products which are typically destined for a tailings storage facility on a typical mine site. The paste backfill system utilizes a large portion of the mixed oxides and leach residue waste products and combines these with cement and water, within a concrete mixer, to produce the paste backfill product.

The mixed oxides waste exits the Hydromet Plant completely dry as it is the product of a calcination step undertaken at very high temperatures. This oxides waste will be crushed to approximately 2 mm in size. The leach residue will be a low moisture filter cake after pressure filtration (approximately 20% moisture content) with the consistency of a fine sand. Approximately half of both waste streams of material will be conveyed to the Paste Backfill Plant location.

The Paste Backfill Plant will be housed within a pre-engineered structural steel building cladded on the sides and roof built upon concrete foundations. The building is 20 m long, 15 m wide and 18 m in height with four floor levels. A 160 tonne capacity outdoor steel cement silo is located alongside the building. The internal floors, above the concrete ground floor, are constructed from structural steel. A stairwell accesses each floor level and personnel access is by grated walkways and platforms around equipment. A 5 tonne capacity overhead crane will provide maintenance and operational lifts from the ground floor to upper floors through a central, open-volume area.

The top floor houses the oxide and leach residue live hoppers, transfer conveyors, mixer feed conveyor, mixer feed chute and cement hopper. The oxide and leach residue hoppers are fed individually by conveyors. Transfer conveyors beneath these hoppers draw material at preset rates out of the hoppers and transfer them to the mixer feed conveyor that moves the products to the mixer. All conveyors have belt scales to measure out materials for the mixer. The cement hopper stores a minimal amount of cement transferred to it from the outdoor cement silo as required during the mixing operation. Cement is added to the mixer through the mixer chute from the cement hopper by way of a screw conveyor and weigh hopper.

The third floor houses the dual shaft paste mixer, a small 1.5 m³ mix water storage tank and feed pumps (duty and standby). The paste mixer receives the oxide and leach residue material from the

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mixer feed conveyor above, through an enclosed mixer feed chute overhead. Water is added by a pump into the mixer from the mixer chute above.

The second floor houses the live bottom concrete hopper. Once a paste is mixed on the third floor, the mixer will discharge the paste into the live concrete hopper. The live hopper acts as a surge area and keeps the paste moving by way of a helical screw shaft. A gate beneath the live concrete hopper empties the paste into a chute that positively feeds either of the paste pumps.

The ground floor houses the two positive displacement pumps and their hydraulic power units. The pumps are in a duty and standby arrangement and are fed by a chute overhead from under the live concrete hopper. The air compressor and dryer system, complete with receivers, supply both instrument air and fluidising air for the cement silo are also located on the ground floor, adjacent to the cement silo.

The cement blower is located alongside the cement silo. The emergency diesel flush pump, small diesel tank and 40 m³ water storage tank are located outside and adjacent to the building for use in the event of an extended power failure, to flush the paste lines. The water storage tank is continually kept topped up by level switches.

The office, control room, electrical room and testing laboratory are also located on the ground floor. An open area is provided in the centre of the building to act as a maintenance area. The area is accessible from two sides of the building using roller doors. The ground floor slab is bunded and shaped to falls, with sumps complete with sump pumps, for cleaning up spills.

15.7.2  Backfill Testwork

The backfill paste formulation and characterization tests were performed at SGS Canada in Lakefield, Ontario between March and July 2017. The mix designs were formulated to achieve 1 MPa at 28-days cure time, meeting the strength requirements for the mine operation. The advantage of sufficiently high early strength gain will be to allow for a flexible mining schedule.

Phase 1 of testing utilized a synthetic oxide material, not representative of the mixed oxides (MOs) expected to be produced in the calcination process, and issues regarding arose with expansion, exothermic reactions and cracking of test samples.

Phase 2 and 3 of testing was able to utilize mixed oxide material produced in a calcination pilot program carried out at Hazen Lab in Golden, Colorado. The result suggested that the formation of Srebrodolskite (Ca₂Fe₂O₅) significantly reduced the CaO content in the mixed oxide solids. This has allowed the backfill paste system to utilize the mixed oxide from the Hydromet Plant as a source of solids without having a self-heating issue due to the exothermic reaction from hydrating CaO.

Utilizing the more representative mixed oxide waste material, together with leached residue (LR), several mixes were tested using a 5% binder, including 100% cement, 100% fly ash, and mixtures of the two. As well, the ratios between mixed oxide and leach residue were tested at 75/25 and 60/40 mixes.

The resulting Phase 2 tests demonstrated that the formulations far exceeded the 1 MPa at 28-day mining requirement using a 5% binder. The mixes tested are presented in Table 15-9.

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Table 15-9: Paste Backfill Formulations for Phase 2 Testwork

Mix #	MO/LR	Excess Oxide	Leach Residue	Type GU* Cement	Fly Ash Binder	Note
1	75/25	71%	24%	5%	0%	Control
2	60/40	57%	38%	5%	0%	Design Production Rate
3	75/25	71%	24%	2.50%	2.50%	50/50 GU/FA
4	75/25	71%	24%	3.75%	1.25%	75/25 GU/FA

*GU Is “General Usage” Or Standard Portland Cement. MO is mixed oxide. LR is leached residue.

The subsequent results at 7 days cure time were very positive. All the mixes resulted in high strength fill (see Table 15-10).

Table 15-10: Results of Phase 2 UCS Testing of paste backfill samples after 7 days

Mix #	Cube 1 (MPa)	Cube 2 (MPa)	Cube 3 (MPa)	Average (Mpa)
1	5.9	6.0	6.3	6.1
2	8.3	8.0	8.6	8.3
3	4.9	4.9	5.0	4.9
4	5.6	6.2	6.4	6.1

The results of the test work proved that a) the use of the mixed oxides with the leach residue was suitable for backfill use and b) the resulting product would have sufficient strength. The comparison for Mix 1 and Mix 2 would also suggest that increasing the proportion of the LR relative to the MO might also provide higher results.

Based on these results, a Phase 3 of testing was conducted in July 2017 to attempt to optimize the recipe, as the total GU cement binder requirement to achieve 1 MPa at 28 days is clearly lower than 5%. The mixes tested are presented in Table 15-11.

Table 15-11: Paste Backfill Formulations for Phase 3 Testwork.

Mix #	MO/LR	Excess Oxide	Leach Residue	Type GU* Cement	Fly Ash Binder	Note
1	75/25	74%	24%	2%	0%	Control (GU cement)
2	75/25	74%	24%	0%	1%	1% fly ash
3	75/25	73%	24%	0%	2%	2% fly ash
4	75/25	73%	24%	0%	3%	3% fly ash

*GU Is “General Usage” Or Standard Portland Cement. MO is mixed oxide. LR is leached residue.

The subsequent results at 7 days and 28 days cure time were acceptable only for the cement binder used. All of the mixes utilizing fly ash resulted in results that are too low in strength to be adequate for the use as structural backfill (see Table 15-12).

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Table 15-12: Results of Phase 3 UCS Testing of paste backfill samples after 7 days

Mix #	7 day UCS* (MPa)	28 day UCS* (MPa)
1	1.39	2.42
2	0.14	0.18
3	0.13	0.21
4	0.14	0.25

*UCS Is “uniaxial compressive strength”.

The results of the test work suggested that a) the use of fly ash only as a binder in the paste mix was not suitable for structural backfill use and b) GU cement content alone, and as low as 2% was sufficient to reach sufficient strength. This would result in a savings in Capex and Opex by requiring only one GU cement silo on site.

Additional test work is proposed with regards to the MO/LR proportions and blend of cement and fly ash as binders. The Phase 2 test work strength comparison for Mix 1 and Mix 2 would also suggest that increasing the proportion of the leach residue (LR) relative to the mixed oxides (MOs) may also provide higher strength result.

It is recommended that further tests be carried out with binder blends to further reduce cement and increase fly ash content considering a local coal power generator should provide relatively cheap fly ash. When comparing Phase 2 Mix 3 result against the Phase 3 Mix 1 result, it would suggest as expected that using fly ash (additional binder) with GU cement provides a higher strength result at 7 days, albeit this strength increase results in additional Capex and Opex and may be unnecessary considering the aim of 1 Mpa at 28 days was achieved without fly ash added in the Phase 3 test 1 result. It is recommended that further tests be carried out with different blends to further reduce cement content and increase fly ash considering a nearby source of cheap fly ash to the site.

15.7.3  Paste Plant Process

The Paste Backfill Plant process is illustrated in Figure 15-8.

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Source: Optimize Group, 2022

Figure 15-8: Paste Backfill Plant process flow diagram

Along with slag material, the Hydromet Process Plant is expected to produce approximately 1588 t/d (66 t/h) of waste mixed oxide solids from the calcination process and 825 t/d (34 t/h) of waste leach residue filter cake.

The requirement to keep up with the voids created underground is approximately 900 m³/day of paste backfill. Accounting for the swell factor, this volume requires approximately 1400 t/d of solids. This relates to approximately 58% of these waste material streams being utilized as ingredients within the paste backfill material per day.

The paste backfill system has a nominal production of 37.8 m³/h (or 81.9 wet tonne/hr). This would be the rate required to keep up with the void created underground in a day. The design production rate is 62.4 m³/h (or 137.8 wet tonne/hr). This design rate allows a catch-up factor of 65% (over design divided by nominal rate). At the design rate, the backfill system can fill the void created underground in 14.5 hours per day. However, the system is expected to run continuously except for planned maintenance.

At the design rate of 62.4 m³/h, and a mix ratio of 60/40 mixed oxides to leached residue, the paste plant process consists of combining 58 t/h of mixed oxides and 38.6 t/h of leached residue for making paste backfill in the 14.5 h/d.

At the design rate of 62.4 m³/h, cement binder at 2% is expected to be consumed at 2.1 t/h (or 32.5 t/d). A 164-tonne silo will allow for approximately 5 days of production.

The mixed oxides are completely dry (due to the high-temperature calcination process) and are received crushed to minus 2 mm with a suitable particle size distribution, before used in the Paste Backfill Plant process. The leach residue filter cake is received in a filter cake form, expected to have 20% to 25% moisture content.

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Both waste materials will be fed from stockpiles by conveyors to live bins in the Paste Backfill Plant from where a pre-determined weight of material is drawn for the mixer. Water addition, adjusted for the leach residue moisture, is expected to be approximately 20 m³/h - 30 m³/h (or 290 m³/d - 435 m³/d).

The mixer receives the pre-weighed mixed oxide and leach residue material from the mixer feed conveyor above, through an enclosed chute. Cement is fed into the chute in pre-determined weight from the cement hopper by way of a screw conveyor and weigh hopper. Water is added by pump in predetermined volume measured by flow meter into the mixer from above.

All the ingredients are fed into the mixer simultaneously, prior to being mixed. This continuous mixer process is expected to have a residence time of 1.5 minutes to 2 minutes. The mixer will then mix the paste in a pre-set time with the mixer shaft drive resistance assisting to ensure consistency of the paste mix.

15.7.4     Underground Distribution of Paste Backfill

The paste is pumped underground from the Backfill Plant with positive displacement pumps (one in operation and one spare), through the production headframe, into the shaft via either of two 150 mm (6”) carbon steel schedule 80 pipes anchored to buntons, discharging on the appropriate level using HDPE pipes for the last sections to the discharge points. A short section of surface pipe will be required to broach the gap between the Paste Backfill Plant and the production headframe.

It is imperative that the paste backfill lines remain clean between each use. To ensure cleanliness, water is flushed through the system from the Paste Backfill Plant through to the underground before paste backfill starts to “slick” the pipelines. Additionally, at the end of a backfill operation, a post flush operation is undertaken with water diverted to an underground sump and pump. Additionally, cleaning pigs are to be used to remove any materials from the inside walls of the slicklines.

As detailed within Section 16.6.3, barricades are to be installed in the lower access drift to the stopes, development level pipe extensions are added to the shaft slicklines from the production shaft via the upper access drift into the stopes, backfill paste flows and fills the stope. Once the stope is filled the backfill is allowed to cure (28 days) to the design strength of over 1 MPa before blasting on the adjoining stope. This ensures the maximum loading on the barricade is kept under 200 kPa.

Rupture spools are used to manage any unexpected high pressure in the pipeline. The paste is directed from the rupture point to a sump for safety and easy clean-up. Pressure transmitters sense ruptures from pressure readings taken the control room in the Paste Backfill Plant. Water-proof cameras with night vision can also be used to monitor the underground backfill operation.

In the event of a prolonged power outage, a diesel-powered flush pump can be used for an emergency flush of the paste pipeline.

15.8  Freeze Plant

Key to the revised plan to develop the shafts for the mine access will be the installation of a Freeze Plant that will provide super-cooled brine to be utilized for freezing the ground from the surface through the limestone to the carbonatite interface. The use of this technology allows the project to complete these excavations without the need for an extensive pumping system. Ground freezing is an additional sealing measure without any structural requirements.

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The Freeze Plant will require a 4 MW cooling facility that will prepare and recirculate supercooled brine through a number of deep boreholes surrounding the two shafts. The boreholes, which will be 200 mm (8”) in diameter, will utilize insert pipes of a smaller diameter to push the brine down to the carbonatite and allowing it to recirculate to the surface and back to the Freeze Plant.

The plant itself will consist of compressor houses and cooling coil sets in gangs according to the final required capacity. A typical arrangement is shown in Figure 15-9.

 

Source: Nordmin, 2019

Figure 15-9: Typical Freeze Plant Configuration (with gangs of compressors and cooling coils in series to make up the total capacity of the plant)

The boreholes around each shaft are arranged radially around the planned perimeter of the excavation, as shown in Figure 15-10. The actual working diameter of the freeze hole perimeter and the number of holes is determined by geotechnical design.

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Source: Nordmin, 2019

Figure 15-10: Typical Layout of a Freezewall Borehole System

The red perimeter holes are used for freezing the shaft envelope, which is shown as the inner ring.

The stabilization of the shaft envelopes down to the carbonatite is critical to the progress of the project. To this end, the freeze will start three to six months prior to commencement of shaft sinking and will be left in place until one month after the shaft liner is socketed and sealed into the carbonatite.

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16. MARKET STUDIES

The market section of this TRS constitutes a review of current and historical market reports and offtake agreements, as may be relevant, for niobium, scandium, titanium, and rare earth elements (REEs). Dahrouge has reviewed these market reports as well as previous reports for the Project prepared under the Canadian NI 43-101 standard to gain an understanding of the markets for the Company’s proposed products. Dahrouge has adapted/modified/excerpted materials from these previous reports, most notably from the Company’s 2019 Feasibility Study for the project (Nordmin Resource & Industrial Engineering, (2019).

16.1 Market Studies

Understanding the markets for niobium (Nb), titanium dioxide (TiO₂) and scandium trioxide (Sc₂O₃) are an important part of the proposed Elk Creek Mine and end-product determinations. These products, especially niobium and scandium trioxide (scandium), are thinly traded and somewhat opaque markets without well-established publicly available pricing.

The rare earth elements (lanthanides plus yttrium), which are confined in this report to the Mineral Resource, comprise a wide variety of markets, some more thinly traded and opaque than others. However, the magnet feed rare earths (neodymium, praseodymium, terbium, and dysprosium), which carry the large majority of the rare earth value at Elk Creek, are more widely traded and therefore pricing may be readily acquired from several commercial services.

16.1.1   Niobium Market Overview

Niobium is a versatile element that adds value to a range of applications. Niobium improves material properties, which often leads to increased efficiency, safety, performance, and transform’s the properties of advanced steels, cast aluminum, glass, batteries and electronics. Ferroniobium in steelmaking consumes approximately 90% of the available world supply of niobium. The remainder goes into a wide range of smaller volume but higher value applications, such as high-performance alloys (which includes superalloys), carbides, superconductors, electronic components and functional ceramics.

Commercial trade of niobium occurs in several forms, the most common of which is ferroniobium. Ferroniobium is sold most commonly as steel grade (65% Nb content) as well as a higher purity technical grade.

16.1.1.1   Niobium Supply

The niobium market is generally described as an oligopoly with three major producers dominating supply. These three producers are Companhia Brasileira de Metalurgia e Mineração (CBMM), Magris Resources, and China Molybdenum Co. Ltd (CMOC). However, in practical terms, the market operates as a monopoly with a single company (CBMM) setting the price and the other operations acting as price takers. In addition, CBMM performs its own research and development activities to evaluate additional/increasing usage of niobium, which provides a significant benefit for other market participants. Over many decades, CBMM has become a very reliable producer and has significantly reduced supply disruptions and in return has increased supply to accommodate overall demand growth. In terms of ferroniobium production, Table 16-1 provides the reported annual production capacity from the three largest mine operations along with the project’s estimates.

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Table 16-1: Comparison of Project Versus Selected Niobium Producers

Mine/Project	Owner	Country	Reserves (Est.)	Annual Ferroniobium Production  (Est.)
Araxa (OP)	CBMM	Brazil	829 Mt @ 2.5% Nb2O5 (weathered)  936 Mt @ 1.57% Nb2O5 (fresh)1	110 kt/y2
Niobec (UG)	Magris Resources	Canada	Proven 19.9 Mt @ 0.51% Nb2O5 Probable 54.5 Mt @ 0.51% Nb2O53	9.2 kt/y3
Elk Creek (UG)	NioCorp	USA	Probable 36 Mt @ 0.81% Nb2O5	7.2 kt/y
Catalao (OP)	CMOC	Brazil	Area I 37.4 Mt at 0.97% Nb2O5  Area II 217.7 Mt at 0.34% Nb2O54	13.8 kt/y5

Source: NioCorp, 2019

¹CBMM, 2017, Sustainability Report. CBMM does not report reserves, only resources

²Roskill, 2018

³Roskill, 2017

⁴CMOC Annual Report, 2017

⁵Roskill, 2017

Niobium is not traded in public markets. Transactions generally occur directly between mine operators and downstream consumers. Trading firms also play a smaller role in the market as intermediaries. There are several quoted prices for various ferroniobium and niobium oxide products that are established based on these transactions with traders.

16.1.1.2   Niobium Demand

Based on market information provided by CBMM, Niobium demand showed an active profile of growth over almost 20 years from the early 1990s to late 2000s, greatly exceeding growth rates in steel demand (see Figure 16-1). CBMM expects growth to continue and is actively developing additional applications for niobium as a component of solid-state lithium-ion batteries, in collaboration with Toshiba. These batteries could potentially provide longer ranges in automotive application and provide charging times measured in minutes. CBMM has reportedly devoted a 50,000 tonne per year capacity expansion devoted to supplying this application for niobium (Fucuchima, 2022).

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Source: CBMM, 2013

Figure 16-1: CMBB Niobium Sales Versus Steel Demand and Niobium Intensity of Use

Pricing

Figure 16-2 demonstrates recent price trends for 65% ferroniobium (pricing basis anticipated for NioCorp).

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Source: Argus Media, 2022

Figure 16-2: Ferroniobium (65% - EU) Price Trends Previous Quarter

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Future prices of niobium are highly dependent upon the intention of CBMM. CBMM could flood the market with low-cost production, dropping the price and driving out its competitors, however to date, CBMM has shown a tolerance for other producers and has controlled its production levels to maintain a stable price.

Roskill’s Global Industry, Markets and Outlook 2018 (Roskill, 2018) has indicated that Niobium prices are historically very stable. They moved little in the period up to about 2006, when a producer-driven doubling in the pricing began and have remained stable at the higher benchmark. Ferroniobium prices, in particular, are an inelastic demand, with the 2009 slump in demand from the global steel industry having only minimal impact on pricing. The outlook for prices is one of a gentle but steady increase; spikes are unlikely. The economic analysis in this report used the US$ 47/kg Nb as the forward-looking price for steel grade (65%) ferroniobium.

16.1.2   Titanium Dioxide Market Overview

The global titanium dioxide market size is currently US$ 11.4 billion and is expected to grow steadily, owing to its growing demand in end-use industries such as plastics, coatings, paper, cosmetics and others. Furthermore, technological innovations in manufacturing processes, which have resulted in higher and good quality yield, positively impact the overall titanium market (see Figure 16-3, Adroit Market Research, March 2019).

 

Source: Adroit Market Research, 2019

Figure 16-3: Global Titanium Dioxide Market Value and Volume 2014-2025

TiO₂ is used extensively as paint pigment with some minor, though increasing, demand from the aerospace industry as an alloy in next-generation aircraft. Titanium oxide compounds are also being developed for deployment in the next generation of lithium-ion batteries, using solid state formulations that provide longer ranges in automotive applications and charging times measured in minutes (Sciencebriefs 2022). Average domestic US consumption in 2018 was 920,000 t and the USGS reports that imports supply approximately 90% of US demand. With the Project producing approximately 12,000 t TiO₂ per year during LOM, it is assumed that this annual production volume can easily be absorbed into the domestic market.

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16.1.2.1   Titanium Dioxide Demand

The competitive landscape of global titanium dioxide market is highly fragmented with a large number of global and regional players including Henan Billions Chemicals Co., The Chemours Company, Huntsman International LLC, NL Industries, Inc., Tronox Limited and others. These prominent players have always looked forward to implementing essential strategies through partnerships, agreements, collaborations and business expansions.

Formal market studies were not completed at this time as TiO₂ represents only 2% of the overall revenue in the economic analysis. All market information for titanium and titanium dioxide is derived from USGS Commodity Market Summaries (Bedinger, 2019) and Adroit Market Research (Johnson, 2019).

The economic analysis assumes a constant long-term price of US$ 0.99/kg, based on rutile concentrate FOB Australia benchmark with no discounts (see Table 16-2). Pricing has shown a significant rebound in the recent period. Although the market is well-established and mature, the key risk to maintaining this price is the domestic US and global economic growth.

Table 16-2: Titanium Mineral Concentrates Pricing History (Rutile Concentrate FOB Australia)

	2014	2015	2016	2017	2018	2019
95% TiO2 Price (US$ /kg)	0.95	0.84	0.74	0.74	0.99	0.99

Source: USGS MCS 2019, as presented in Nordmin Resource & Industrial Engineering, 2019

16.1.3   Scandium Trioxide Market Overview

The majority of scandium production origninates in China and is a by-product of iron ore and rare earth production. Scandium has critical utilization in areas such as the aerospace industry, solid fuel cells, electronics industry and is also used in metallurgical applications (Altinsel et al., 2018). Scandium is the 50^th most abundant element with a crustal abundance of 20 - 30 ppm. However, scandium does not have any identified single deposit type due to its natural occurrence being as a dispersed state. Due to the scarcity of high-grade scandium deposits, and high processing costs, scandium production rate is relatively limited. Scandium is generally produced as a co-product of primary metal processes, wastes and reprocessed tailings (Altinsel et al., 2018).

Given the relatively opaque nature of the scandium market, NioCorp engaged OnG Commodities LLC (OnG) to produce an independent market assessment and a report was provided to the Company in April 2017 (OnG 2017). This report was updated via a subsequent memo by OnG in 2019 and concluded that the scandium oxide demand and pricing environment remained robust (OnG 2019). Specifically, the updated market assessment noted that the realization of several projects targeting commercial production of scandium had slipped by about two (2) years. Therefore, OnG concluded that “forecasts for supply, demand, and pricing, in the period 2020 onwards, should be adjusted out by two years and are otherwise today [2019] an appropriate forecast for the scandium market” (OnG, 2019).

No additional formal market assessment update for scandium was completed for this current TRS; however, phone and email correspondence was carried out between Dahrouge and Dr. Andrew Matheson, author of the 2017 scandium market assessment report and the 2019 update by OnG (Pers. Comms with OnG, May 13, 2022). This market information from this correspondence has been incorporated in the text below.

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The OnG studies and related correspondence examine recent and current scandium production trends (~15 t/y in 2018 and ~30 t/y in 2021) from existing and emerging producers plus an outlook for supply to 2030. The outlook then reviews the current and emerging applications for scandium including fuel cells, aerospace, industrial and other uses plus an outlook for demand to 2030.

This study outlines that even though the 2018 global market for scandium was approximately 15 t/y in the form of Sc₂O₃, this relatively small amount of production is due to the market demand being relatively muted given the diminutive size of the global market along with a lack of stable supply. This conclusion remains unchanged in today’s market, where scandium oxide production is estimated at 30 t/y. The scandium supply is highly reliant on China as a co-product or by-product of rare earth mining along with increasing supply from the Russian Federations. Accordingly, the distribution of supply, as much as the amount of scandium available, should be seen as an impediment to scandium demand growth. Consequently, this lack of supply has been an inhibitor of demand growth and the lack of demand has depressed supply growth.

It is reasonable, to take the view that until 2010, scandium while promising in principle, was little more than an academic curiosity; due to the unwillingness of any large potential user to commit to developing supply. The situation changed exclusively by the actions of Bloom Energy in the production of solid oxide fuel cells for stationary power generation as well as to power ocean-going vessels. Bloom has contracts with numerous existing and emerging scandium suppliers and is constrained first by the availability of scandium and only second by the price of scandium.

16.1.4 Key Aspects of OnG Commodities Report

Scandium Trioxide Market Supply

Historically, the majority of scandium production has originated in China, as a by-product or co-product of rare-earth production. However, over the last 2-3 years new production has come on to the market. Over this period, the Sumitomo Taganito scandium plant in the Philippines began operations and is currently operating close to its nameplate capacity of 7.5 t/y scandium oxide (Pers. Comms with OnG, May 13, 2022). Additionally, in May 2022 Rio Tinto (Fer et Titane) announced that it had produced its first batch of scandium oxide at a commercial scale at its plant in Sorel-Tracy, Quebec, as a by-product of its iron and titanium operations. Rio Tinto is currently focused on ramping up production to its nameplate capacity of 3 t/y scandium oxide (Rio Tinto, 2022). Drawdowns of former USSR stockpiles and by-product recovery from uranium in situ leaching operations represent the balance of the current world supply. Current scandium oxide supply is estimated in the range of 30 t/y, approximately double that estimated for 2018.

OnG developed a detailed analysis of various production sources expected to come online in the next few years (from 2019). These include resources in Australia, the USA, Turkey, Canada, and India, in addition to the expansion of existing resources within China and Russia. Thorough analysis details the relative challenges entrants may face monetizing these resources given the new technologies being developed. OnG develops two primary forecast ranges for scandium oxide supply, differentiated by the inclusion of Russia and China (see Figure 16-4 and Figure 16-5).

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Source: OnG, 2019, Nordmin Resource & Industrial Engineering, 2019

Figure 16-4: High, Expected, and Low Case Forecasts for Scandium Oxide Potential Supply 2019 – 2030, Tonnes per Year

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Source: OnG 2019, Nordmin Resource & Industrial Engineering, 2019

Figure 16-5: High, Expected, and Low Case Forecasts for Scandium Oxide Potential Supply 2019 – 2030, Tonnes per Year, Excluding Russia and China

Scandium Trioxide Market Demand

OnG speculates that scandium has two primary applications (1) As an alloying agent in aluminum alloys (with aerospace the largest candidate market) and (2) in solid oxide fuel cells (SOFC). The SOFC market is currently the largest single consumer of scandium and is almost entirely constituted by Bloom Energy of the US. Also, fuel efficiency standards driven by an increasing focus on carbon emissions in the EU is anticipated to lead to a dramatic increase in scandium usage within the transportation sector (see Figure 16-6).

New and emerging developments for demand include Sc-Al alloys, with relatively high Sc content, used in 5G cell phones and network towers and is expected to be the preferred alloy in these applications. A current cell phone contains 10-12 antennae and represents a potentially new and significant source of demand. Additionally, scandium is being incorporated into an EV’s heat exchanger system resulting in more energy efficient cabin heating, as well as battery enclosures allowing for increased mechanical strength and minimization of structural stress (Pers. Comms with OnG, May 13, 2022).

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The market studies and subsequent correspondence with OnG conclude that the scandium market can absorb a significant increase in supply and will require additional supply to support the large-scale adoption of the current and emerging uses in SOFC, aerospace, and 5G markets.

 

Source: Sumitomo Metal Mining Co., 2019

Figure 16-6: Current/Potential Scandium Market

Solid Oxide Fuel Cells

OnG provides an overview of scandium use in the SOFC market and the technology that necessitates its use. Scandium is an essential component of Bloom Energy’s SOFCs and delivers high reliability and the ability to operate the fuel cell at a much lower temperature than competing SOFC technologies reliant on yttrium doped zirconia. The lower operating temperature simplifies construction and allows for less costly materials of construction. Using published data, SOFC manufacturing requires an estimated 150 kg of scandium oxide per MW of power. There are no true substitutes that deliver an equivalent level of performance.

According to OnG (Pers. Comms with OnG, May 13, 2022), the sector is making significant gains towards large scale adoption (e.g., Plug Power and non-public investments). Additionally, SOFC used in ocean going vessels and the development of green hydrogen demand presents a new and significant opportunity for Bloom Energy. This emerging demand provides a strong foundation for continued long-term growth in demand for the SOFC market.

In 2018, Bloom Energy became a public company and at this time stated its intention to maintain a growth rate of 40% of higher in systems installed. In 2018, system sales for SOFC by Bloom Energy were reported at 80.9MW. The company is well established in South Korea (a large fuel cell market) and has a growing foothold in the east coast of the United States

Aircraft Aluminum Alloys

The 2017 market study by OnG (and subsequent update in 2019) provides significant background on the Aluminum Scandium (AlSc) market opportunity. The aerospace industry was an early

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adopter of the alloy and has accumulated years of experience with the materials. The lack of a reliable supply has been the primary barrier to broad commercial market adoption.

Very little scandium is necessary for AlSc alloys, and less than 0.5% scandium is sufficient and loading as low as 0.1% can be adequate (although Airbus’ patented alloys can contain up to 1.3% scandium). A typical single-aisle jetliner, such as an Airbus A320 or a Boeing 737, has a dry weight of 45 t to 50 t, which is mostly (80% by weight) aircraft aluminum. According to Airbus, scandium alloys can reduce this weight by an estimated 15% to 20%, or by 6 t to 10 t.

Assuming the AlSc alloy is 1% scandium, each aircraft would require approximately 600 kg of scandium oxide, approximately US$ 2.1 million at 2017 market prices, while the lifetime value of fuel savings would total US$ 20 to 30 million.

The weight reductions come to some degree from the ability to use less aluminum alloy when it is alloyed with scandium. The majority of weight savings accrue from the ability to weld the airframe. Welding eliminates thousands of rivets currently needed to fasten an aluminum aircraft together. Welding also has the potential to save time and cost in aircraft assembly, offering further benefits to a switch to AlSc alloys.

OnG notes a key question for scandium demand growth will be driven by the pace of adoption by aerospace firms. Specifically referencing the A320, current production rates would necessitate at least 100 t/y of scandium oxide if only key components transitioned to AlSc alloys with 1% Sc content. That grows to nearly 250 t/y if all aluminum components were transitioned, and the buy-to-fly impact on required input scandium would increase both these quantities substantially (aluminum buy-to-fly ratios in civilian aerospace vary by component but can commonly reach 5:1 - as reported for Constellium’s Airware alloys (deployed in the Airbus A350) for example.

Widespread adoption will, therefore, take time. Under realistic supply-side scenarios, the early to mid-2020s is the earliest period when large-scale deployments of AlSc could be expected in passenger aircraft. This will be because of supply chain issues primarily because AlSc alloys are well characterized and understood for aerospace applications. OnG draws a link to the development of the Airbus A380 which required new LiAl alloys for wing main spars, new ingot casting techniques, and new manufacturing and assembly. The entire process from inception to launch required seven years. Scandium could potentially be adopted faster if the supply side is well established because the foundational alloy development and understanding has already been completed. Further, global capacity for lithium aluminum alloys is approaching 50,000 t/y, which if replicated for scandium would represent 250 to 500 t/y of scandium usage depending on the level of scandium doping.

Other Markets

OnG provides additional context on the broader adoption of AlSc alloys in the transportation sector, defence sector, and as a replacement for titanium. The potential also exists for growth in smaller existing markets, such as sporting equipment, stadium lighting, handguns, specialty alloys and lasers. These potentials are disregarded for price forecasting in the OnG analysis.

Figure 16-7 provides a summary chart of the aggregate demand by the differing sectors.

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Source: OnG, 2019, Nordmin Resource & Industrial Engineering, 2019

Figure 16-7: Supply-Demand Forecast for Scandium Oxide to 2032, Tonnes, Base Case

Each of the independent market segments is expected to drive significant demand growth over the next ten years. As supply begins to align with demand in 2025, a deficit again appears in 2028.

Market Pricing

OnG provides a forecast of market pricing and the context of current scandium pricing with the following statements:

● Price trends are more reliable than the actual quoted numbers. The general increase in scandium oxide pricing reported by the USGS since 2010, and the narrowing of the spread between low purity and high purity scandium oxide pricing, does reflect an increase in consumption (by Bloom Energy of California) as well as a willingness to purify lower grade scandium oxide through secondary reprocessing. OnG goes on to note that Bloom has, to date, been willing to purchase all the scandium available to it, has entertained long term supply agreements with many (if not all) of the existing and emerging scandium suppliers, and has managed to grow at rates exceeding 40% per year despite scandium oxide prices in the range of US$ 3,500 to 4,000/kg.

● Too much supply would inevitably depress prices in the long run. A substantial increase in supply from a more diverse set of countries and underwritten by well-capitalized mining operations could increase the size of the scandium market and support prices at today’s levels.

OnG presents two scenarios for market pricing driven by aggregate aerospace adoption of the increased supply of scandium (OnG, 2017, 2019; Nordmin Resource & Industrial Engineering,

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2019). Under the “base case” assumptions, scandium oxide prices will likely rise slowly from their current level of US$ 3,500/kg to US$ 4,000/kg by 2022 as demand begins to outstrip supply. With new Western operations beginning from 2023 – 2027, there is likely to be a period of moderate oversupply, causing a softening of prices to US$ 3,000/kg. This oversupply period is expected to support substantial growth in aerospace demand. By 2027, as demand begins to outstrip supply prices will likely rise.

This scenario is considered probable even if the market returns to undersupply in 2027 since suppliers will have entered into contracts as they commission plants and because market tightness will take time to manifest. From 2028 the market should recover strongly to a level of US$ 3,750/kg.

However, if aerospace demand is slow to materialize, prices may fall through 2027 to a level of around US$ 2,500/kg, before turning around in 2028 as delayed aerospace growth begins to tighten the supply of scandium oxide. Prices are unlikely to fall below this due to the relatively short periods of supply excess. Further, Bloom and industrial users are likely to make efforts to accelerate growth (see Figure 16-8).

 

Source: OnG, 2019, Nordmin Resource & Industrial Engineering, 2019

Figure 16-8: Scandium Oxide Pricing Outlook, US$/kg, 2019 – 2030

Summary

Based on these inputs, the following summaries of OnG pricing forecasts and global demand volumes by year to 2030 based on estimated production costs and supply-demand balances.

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These forecasts, plus the Project’s estimated annual scandium production volumes, are shown in Table 16-3 and Figure 16-9.

Table 16-3: Scandium Supply, Demand and Price Forecast Summary

Description	Price (US$/kg)	Est. Global Supply (kg)	Est. Global Demand (kg)	Project Annual Production (kg)	% of Est. Global Supply	% of Est. Global Demand
2019	3,600	8,000	21,168
2020	3,700	15,000	30,236
2021	3,800	29,000	45,530
2022	3,900	90,750	70,342	-	-	-
2023	4,000	183,810	114,731	-	-	-
2024	3,500	228,950	162,861	-	-	-
2025	3,000	316,000	260,705	47,750	15%	18%
2026	3,000	366,330	364,488	112,110	31%	31%
2027	3,200	443,120	465,755	108,500	24%	23%
2028	3,400	488,380	571,452	103,400	21%	18%
2029	3,600	595,430	664,029	95,330	16%	14%
2030+	3,750	631,670	771,577	96,120	15%	12%

Source: OnG, 2019, Nordmin Resource & Industrial Engineering, 2019 (modified with Elk Creek Project’s projected annual production and corresponding estimated % of global supply and demand)

 

Source: OnG, 2019,

Figure 16-9: Global Scandium Supply/Demand and Price Projections Summary

From an overall market standpoint, demand for scandium oxide is straining supply, and there are few other truly near-term opportunities to increase supply. So, for emerging larger scale producers such as NioCorp, a few extra tonnes of supply out of Russia will create the potential for

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further market growth as they develop their supply, just as the Sumitomo project will be beneficial for all.

Currently, a significant amount of scandium production is from China, which does not have transparency in reserves and cost reporting. If production from other parts of the world, outside of Russia, begins to take off as projected, it is not clear whether Chinese or even Russian production will increase and keep new entrants from entering the market. Conversely, if the entrance of a few new producers to the market stimulates demand, and the new entrants and existing producers cannot meet that demand, market pricing will adjust positively.

Dahrouge recommends a full update to the 2017 market assessment report for scandium be completed by OnG (OnG 2017, 2019) as a next step in assessment of the market and its potential impacts to the Elk Creek Project. As the last forecasts of the market (OnG, 2019) are now three (3) years old, an update is prudent. Moreover, there have been new entrants into the supply side of the scandium market since the last market update (OnG 2019), in addition to recent and major global events – most notably the Russian invasion of Ukraine and COVID pandemic – further supporting the need for a revised market assessment for what is a very opaque market.

16.1.5    Rare Earth Market Overview

In the mineral exploration and mining industry, the rare earth elements (REEs) are comprised of fifteen (15) elements – lanthanum (La), cerium (Ce), praseodymium (Pr), neodymium (Nd), samarium (Sm), europium (Eu), gadolinium (Gd), terbium (Tb), dysprosium (Dy), holmium (Ho), erbium (Er), thulium (Tm), ytterbium (Yb), lutetium (Lu), and yttrium (Y). Specifically, these are the lanthanide elements plus yttrium. Rare earths are not termed ‘rare’ because they are hard to find in nature – Ce has a similar crustal abundance to Cu – but rather they are rare because they are very difficult to find in a manner that is economic to extract. REEs always occur together and must be recovered together into an intermediate product before they may be separated into their individual oxide forms for dissemination into their respective downstream supply chains.

The rare earth elements comprise a wide variety of markets, some more thinly traded and opaque than others. The main uses of rare earths may be grouped into eight categories:

● Battery alloys: used in rechargeable batteries for hybrid electric vehicles, power tools, etc.

● Catalysts: used in catalytic converters, fuel cracking catalysts, etc.

● Ceramics, pigments and glazes: used in applications which necessitate high temperature stability

● Glass polishing powders and additives: used in optical glass to mobile phones and LCD screens

● Metallurgy and alloys: added to liquid steel during steelmaking

● Permanent magnets: for use in motors

● Phosphors: used in lamps and backlighting

● Other: uses in chemicals, materials and technologies such as communications, defense and healthcare.

Of these eight (8) categories, two (2) account for roughly 60% of the global market by volume (catalysts and permanent magnets). Although roughly 1/3 of the market by volume, the magnet feed rare earths (neodymium, praseodymium, terbium, and dysprosium) account for roughly 90% of the value of the overall rare earth market consumption (Adamas Intelligence 2019) (Figure 16-

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10). As such, these four REEs dominate the REE value within essentially all REE mineral projects, including Elk Creek.

 

Source: Adamas Intelligence, 2019

Figure 16-10: REE uses by volume and by value

Market Demand and Supply

Demand for the magnet feed REEs (Nd, Pr, Tb, and Dy) make up the vast majority of global REE value today and, in the years ahead, demand growth for these four REEs is expected to exceed demand growth for all other rare earth elements, challenging the ability of the supply-side to keep up. Adamas Intelligence forecasts a Compound Annual Growth Rate (CAGR) of 8.6% for NdFeB permanent magnets from 2022 through 2035, which translates into comparable demand growth for the magnet feed REEs (Adamas Intelligence 2022). This follows on the heels of a 9.3% COVID induced drop in global REE permanent magnet consumption in 2020 to 113,695 tonnes, which is forecasted to rebound sharply in 2021 (up 23.5% year-over-year) and in 2022 (up another 13% year-over-year) (Adamas Intelligence 2020). Thereafter, it is forecast that demand will increase at a CAGR of 7.6% through 2030, to 285,923 tonnes, on the back of strong demand growth in virtually all magnet-related end-use categories (Figure 16-11).

 

Source: Adamas Intelligence, 2020

Figure 16-11: Historical global consumption and forecasted demand for NdFeB magnets by end-use category

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This demand growth for rare earth permanent magnets (also termed ‘NdFeB magnets’) is supported by the continuing strong and forecasted long-term growth from EV traction motors and micromotors, wind power sectors, and other applications that require high-performance rare earth permanent magnets (consumer appliances, cordless power tools, industrial robots, speakers, etc.). Significant shortages of all four of these REEs – Nd, Pr, Tb, Dy – are projected over the next decade (Adamas Intelligence 2020). From 2020 through 2030, it is forecast that the greatest demand growth for rare earth permanent magnets will come from commercial EV traction motors (40.8% CAGR), passenger EV traction motors (25.7% CAGR), and consumer appliances (16.3% CAGR), among others. NdFeB magnet demand for other e-mobility applications, including electric bicycles, scooters, mopeds, quadricycles, motorcycles, and low-speed passenger EVs, is expected to increase as a CAGR of 13.2% from 2022 through 2030, while demand for NdFeB magnets for wind power generators will increase at a CAGR of 7.6% over the same period (Adamas Intelligence, 2020).

 

Source: Adamas Intelligence, 2020

Figure 16-12: NdFeB magnet demand forecast for passenger EV traction motors

As alluded to above, the driving force of demand for Nd, Pr, Tb, and Dy are high performance permanent magnets. There is a wide array of rare earth permanent magnet specifications, each designed to fit a specific high-performance application. In general, a Nd2Fe14B permanent magnet (often abbreviated to ‘NdFeB’) is composed of approximately one-third Nd+Pr (Nd>>Pr), two-thirds iron (Fe), and minor boron (B) to act as a binder agent. For operating at higher temperatures (typically >140°C), Dy and Tb are added, typically in the range of up to several percent.

NdFeB permanent magnet alloy is the strongest type of permanent magnet material commercially available today in terms of maximum energy product (i.e., magnetic flux output per unit volume, measured in megagauss-oersteds (MGOe) or Joules per cubic meter (J/m3)). As such, NdFeB magnets have largely supplanted SmCo, AlNiCo, and ferrite magnets in many size-and weight-sensitive applications since the 1980s, and simultaneously have enabled the conception and miniaturization of a wide array of ubiquitous gadgets and electronics that have pervaded modern society.

Higher performance coupled with miniaturization means more can be done with less - specifically micro-motors. An example would be the vibrate function of a cell phone, which is generated by a very small NdFeB magnet motor. If this motor was a ferrite magnet motor, the cell phone would be significantly larger and heavier by comparison which could materially impact the portability of

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the device. A major application where high-performance magnets are required are automobiles. A typical automobile may contain hundreds of small actuators and motors, in addition to the permanent magnet motor in the case of EVs, which is the primary use of NdFeB magnets in EVs today. The use of NdFeB permanent magnets in EVs allows for the light weighting of the vehicle through miniaturization of motors, as well as for the higher efficiency/performance that a permanent magnet traction motor provides compared to a non permanent magnet traction motor equivalent. Collectively, this allows for the EV to travel greater distances and at higher motor efficiency than otherwise possible and, effectively, supports increased adoption. China dominates rare earth magnet demand and therefore, by extension, dominates the demand for Nd, Pr, Tb, and Dy, and is forecasted to remain in this dominant position through 2030 (Adamas Intelligence 2020). China is followed by Japan, Europe, United States, and rest of the world (Figure 16-13).

 

Source: Adamas Intelligence, 2020

Figure 16-13: NdFeB magnet demand forecast for passenger EV traction motors

In addition to dominating demand, China also dominates the production of REEs. The largest source of rare earths globally is the Bayan Obo mine, in the Baotou region of China, and has historically been the dominate global producer of REEs since the 1990s (Figure 16-14). Rare earths are also produced from the Maoniuping mine and several other mines in China. Recently, within the last decade, additional production has been brought online, most notably from Myanmar (south Asian clays), Australia (Mt. Weld), and the United States (Mountain Pass).

Events over the last decade, coupled with the concentration of supply from China, and more recently Myanmar, has highlighted the need for security of supply and therefore, production outside of these jurisdictions. Currently, the only commercial mining production of REEs in the United States is from the Mountain Pass Mine in California and from mineral sands operations in the southeast portion of the country.

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Source: Adapted from Geology 2022

Figure 16-14: Rare earth oxide production by region

Pricing

In nature, all fifteen (15) REEs occur together in the same minerals. Therefore, during processing they must all be processed together until they are able to be separated into their individual oxide products where they can then be disseminated into their specific downstream value chains. This creates what has been termed a “balance” problem, whereby lower value REEs (e.g., Ce and La) must be processed and recovered in order to also recover the higher value REEs (e.g. Nd and Pr). This has significant implications to the cost of processing targeted individual REEs – most specifically, the magnet feed REEs. This creates an oversupplied market for certain REEs such as Ce and La and results in depressed market pricing for those respective commodities. Alternatively, this results in a relatively higher processing cost for the targeted REEs – Nd, Pr, Tb, Dy – and can also limit production capacity. However, this also has the affect of adding upward price pressure to these particular REEs (Nd, Pr, Tb, Dy) to compensate for this additional processing costs (Figure 16-15).

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Source: Adapted from Adamas Intelligence, 2019

Figure 16-15: Relative price forecast for individual REEs

Adamas Intelligence forecasts a CAGR of 8.6% for NdFeB permanent magnets from 2022 through 2035, which translates into comparable demand growth for the magnet feed REEs – Nd, Pr, Tb, Dy (Adamas Intelligence 2022). Additionally, from 2020 through 2030, Adamas Intelligence forecasts that “global annual demand for NdFeB will consistently and increasingly exceed global annual production, translating to shortages of 19,224 tonnes by 2025 and 48,007 tonnes by 2030” (Adamas Intelligence 2022). This strong demand growth will require significant added production capacity to come online and in turn is forecasted to underpin a positive long-term pricing environment.

Pricing forecasts by Adamas Intelligence for Nd oxide, Pr oxide, Tb oxide, and Dy oxide are presented in Figure 16-16 below (Adamas Intelligence, 2020). These pricing forecasts support the inclusion of REEs in the Mineral Resource for the Elk Creek Project. For additional reference, the current spot prices for Nd oxide, Pr oxide, Tb oxide, and Dy oxide, respectively, are approximately $142 USD/kg, $142 USD/kg, $2,167 USD/Kg, and $388 USD/Kg, highlighting the demand in a market that appears posed to be chronically undersupplied over the next decade (Baiinfo, 2022).

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Forecast China Domestic Prices of Nd Oxide, Pr Oxide and NdPr Oxide

 

Source: Adamas Intelligence, 2020

Forecast China Domestic Price of Dy Oxide

 

Source: Adamas Intelligence, 2020

Forecast China Domestic Price of Terbium Oxide

 

Source: Adamas Intelligence, 2020

Source: Adamas Intelligence, 2020

Figure 16-16: Pricing forecasts out to 2030 for the magnet feed rare earth oxides

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16.2 Contracts and Status

At the time of this report, NioCorp had entered into three offtake agreements covering ferroniobium and scandium trioxide production from the Project.

Each ferroniobium agreement has a ten-year term which, when combined, means 75% of the projected production is contracted at a 3.75% discount to the quoted Metal Pages. ⁴ price (unless a premium can be achieved by the offtake customers, which is uncertain).

The scandium trioxide offtake agreement is structured similarly. The agreement has a ten-year term and a minimum of 12 t/y. At that rate, approximately 10 - 15% of the projected annual production is contracted. Further, the customer may elect to take more material in any given year above the prescribed minimum quantity.

No offtake agreements have been executed at the time of the report for the titanium dioxide product from the Project. It is assumed this product and all other material not covered by an offtake agreement will be sold on a spot price, ex-mine gate basis.

In addition to the offtake agreements noted above, supply contracts for natural gas transportation from Tallgrass Energy and natural gas supply from Tenaska have been executed at the time of this report.

⁴ As of May 6, 2014, Metal-Pages Ltd. operates as a subsidiary of Argus Media Limited. https://www.argusmedia.com/metals/argus-metal-prices

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17. ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS

17.1 Environmental Studies

17.1.1 Soils

According to the Natural Resources Conservation Service (NRCS), soils in the vicinity of the Project are primarily comprised of clay, silty clay, silt loam, and clay loam within an ecological site that is typified as “Rangeland.” For all soil types, the depth to any soil restrictive layer is more than 200 cm below ground surface (bgs), and the infiltration is generally “slow” to “very slow.” Soils in the area are generally eroded and range in slopes from 2% to 30%, with the majority of the area having slopes of between 6% and 11%. (NRCS, 2015)

17.1.2 Climate/Meteorology/Air Quality

A dedicated meteorological station was installed at Elk Creek in July 2014. Parameter measurements included in the overall instrument package include:

● Wind Speed

● Wind Direction

● Temperature

● Temperature Difference

● Dew Point Temperature

● Precipitation

● Pressure

● Solar Radiation

The meteorological data thus far collected includes continuous monitoring that has been audited periodically by a third party and can subsequently be used in air quality modelling and permitting.

In September 2016, NioCorp met with the Nebraska Department of Environment and Energy (NDEE) regarding the on-site air monitoring program and the air quality permit application process. It was decided that the ambient monitoring program needed to include PM_2.5 data collection, in light of the attention that this parameter has been given recently by the U.S. EPA. Air quality monitoring was conducted from March 6 to August 20, 2017: the PM_2.5 monitoring was initiated at the Elk Creek site in February 2017; a PM₁₀, monitor was added in March 2017, along with co-located PM_2.5, and monitoring for four gasses including CO, NO_x, SO_x and ozone (O₃).

The ambient air quality monitoring results indicate values less than the appliable national ambient air quality standards. The ambient air quality monitoring results are also similar to values obtained from ambient air quality monitoring systems operated by the NDEE and the Lincoln Lancaster County Health Department during the same time period.

17.1.3 Cultural and Archeological Resources

There were at least 15 Native American tribes that have inhabited the Great Plains region now incorporated in the State of Nebraska, including the Kansa and Otoe tribes of southeastern

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Nebraska. Of these original inhabitants, there are five federally recognized Indian tribes that remain in Nebraska today, including:

● Omaha Tribe of Nebraska;

● Winnebago Tribe of Nebraska;

● Iowa Tribe of Kansas and Nebraska;

● Ponca Tribe of Nebraska; and

● Santee Sioux Tribe of Nebraska.

Reservations associated with these tribes are located in the northeastern part of the state, over 200 km to the north of Elk Creek.

The Otoe Tribe once lived south of the Platte River in the region of the proposed mine, but in 1881, sold all of their lands in Nebraska to the federal government and moved to Indian Territory (now Oklahoma). No direct tribal consultation appears to be necessary at this time.

In January 2017, Cultural Resources Consulting of Hickman, Nebraska (CRC) conducted archeological resources investigations within the proposed area of potential effect, including the proposed mine Project area and the waterline corridor to the Missouri River. The investigation was intended to determine if there are known archeological sites recorded, or currently unknown, but potentially significant cultural resources that may be impacted within the defined area of potential effect.

Items observed throughout the project area included broken farm machinery pieces such as portions of disks, rake teeth, cultivator tips and nuts and bolts. Observed items greater than 50-years in age also included various colors and types of bottleglass, whiteware, stoneware and tin-cans and can fragments. These historic materials have no known associate as a related assemblage, and no connection to a known person or single event and are considered to be isolated finds. As such the materials have no research potential, and do not meet criteria to be recorded as an archeological site. No special notations or collections were made. No evidence was found to suggest that archeological resources are present on the property proposed for construction of the mine and processing area.

As currently designed, no significant archeological resources will be impacted by construction of the Elk Creek Mine and processing area, the evaporation pond and tailings impoundment area, or installation of the waterline. It is recommended that no further Historic Preservation compliance actions are warranted, and the Project be allowed to proceed as currently planned. (CRC, 2017) During construction, the Project will still be subject to the provisions of the Nebraska Unmarked Human Skeletal Remains and Burial Goods Protection Act.

The waterline to the Missouri was eliminated from the scope of the Project after the archeological resource’s investigations were completed.

17.1.4 Vegetation

Cultivated cropland (principally corn, soy, and alfalfa) makes up the majority of the surface area within the Project boundary. Native and non-agricultural vegetation exist primarily in the form of hedgerows and windbreaks along field margins, and in riparian areas along surface water drainages. According to ecosite descriptions from the NRCS (2015), plant communities within the vicinity of Project consist of annual and perennial weedy forbs and less desirable grasses from abandoned farmland, as well as big bluestem (Andropogon gerardii), smooth brome (Bromus inermis), tall fescue (Schedonorus arundinaceus), switchgrass (Panicum virgatum), Indiangrass

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(Sorghastrum nutans), sideoats grama (Bouteloua curtipendula), little bluestem (Schizachyrium scoparium), Scribner’s rosette grass (Dichanthelium oligosanthes var. scribnerianum), porcupinegrass (Hesperostipa spartea), sedge (Carex), leadplant (Amorpha canescens), eastern redcedar (Quercus macrocarpa), honey locust (Gleditsia triacanthos), and smooth sumac (Rhus glabra).

17.1.5 Wildlife

According to Schneider et al. (2011), the Project is located in Nebraska’s Tallgrass Prairie Ecoregion which is home to more than 300 species of resident and migratory birds and 55 mammal species, most of which can also be found in central and western Nebraska. The small mammal fauna of the Tallgrass Prairie Ecoregion consists of species such as the plains pocket gopher (Geomys bursarius), prairie vole (Microtus ochrogaster), thirteen-lined ground squirrel (Spermophilus tridecemlineatus), and Franklin’s ground squirrel (Spermophilus franklinii). White-tailed deer (Odocoileus virginianus) are the common big game species in the region. The most abundant large predator of the region is the coyote (Canis latrans), but other predators such as the red fox (Vulpes vulpes) and American badger (Taxidea taxus) can be found in the Tallgrass Prairie Ecoregion as well. The bobcat (Lynx rufus), least weasel (Mustela nivalis), and American mink (Neovison vison) can be found in wooded areas, wetlands and along river valleys (Schneider et al. 2011).

17.1.6 Threatened, Endangered, and Special Status Species

The Project and surrounding areas lie in the Southeast Prairies Biologically Unique Landscape within the Tallgrass Prairie Ecoregion of Nebraska (Schneider et al., 2011).

According to the Nebraska Game and Parks Commission (NGPC) Conservation and Environmental Review Tool (CERT) (NGPC, 2022), one state-listed threatened and one federally- and state-listed threatened species have the potential to occur within the immediate vicinity of the Project. The USFWS IPaC tool (USFWS, 2022) identifies one additional candidate species (Candidate: species under consideration for official listing for which there is sufficient information to support listing) with the potential to be affected by activities in the Project location.

Aforementioned threatened, endangered and candidate species are listed below:

● Northern Long-eared Myotis (NLEB) (Myotis septentrionalis) – Federally and State-Listed Threatened

● Western Massasauga (Sistrurus tergeminus)

● Monarch Butterfly (Danaus plexippus)

The development of the NioCorp site may impact the NLEB, massasauga, or monarch butterfly. These impacts can be readily avoided or mitigated during construction by implementing standard conservation measures, including conducting certain activities during winter months when these species are not active. Risks associated with protected species are minimal.

17.1.7 Land Use

Since the settlement of Johnson County, farming for livestock, crops, and pasture has been the most important land use enterprise. Over the years, crop production has shifted from orchards, oats, barley, and rye to corn, soy, wheat, alfalfa, and grain sorghum. Livestock in the county generally consists of hogs, cattle, and milk cows (USDA SCS, 1984). About 4,046.86 hectares (10,000 acres) in Johnson County is irrigated cropland, while about 16,996.78 hectares (42,000 acres) is used for pasture. About 12,949.94 hectares (32,000 acres) of Johnson County is used for

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rangeland, which includes both native prairie that was never broken from sod and areas that were cultivated and then reseeded. Based on known soil types, land use in the vicinity of the Project is best suited for rangeland and native hay, introduced or domestic grasses for pasture and, if irrigated, corn, sorghum, and soybeans (USDA SCS, 1984).

17.1.8 Hydrogeology (Groundwater)

A hydrogeological characterization of the deposit was conducted during the core drilling program. The program included:

● 42 downhole packer-isolated injection and airlift testing in core holes.

● Installation of six, 50 mm (2”) PVC standpipe piezometers isolated in the carbonatite and open to large intervals of the deposit.

● Installation of two, nominal 50 mm (2”) PVC standpipe piezometers isolated in the 180 m (590 ft) thick Pennsylvanian aquitard above the carbonatite.

● Frequent measurement of water levels in open core holes and piezometers over six months.

The hydrogeological characteristics of the resource area were significant enough that a 10-day pumping test was conducted in the fall of 2014. During this initial test, an open borehole was pumped at 7.9 m³/h (35 gpm), and the response was observed in nearby piezometers. These data were used to establish the prospective mine water inflow prediction that appears in the 2015 PEA level documents. However, the hydrogeological issues associated with these initial findings were considered to be significant enough for a second test, conducted in May and June of 2015. For this second test, a large diameter injection well was installed in the approximate center of the deposit, and two additional distant monitoring piezometers were established. Water was injected at a rate of 22 to 30 L/s (350 to 480 gpm) over a nominal 30-day period, and the response was measured by a series of instrumented piezometers. Analysis and interpretation of the data from these testing programs have been completed, and a preliminary conceptual model developed.

In 2017, NioCorp engaged Adrian Brown, an expert mining geohydrological consultant, to re-analyze the data set generated during the previous investigations. He concluded that mine inflow control could be achieved at this project using ground freezing and grouting for the shafts and grouting in the mine development drives and stopes. This approach is designed to limit the peak mine water inflow to around 66 L/s (1,000 gpm), and the LOM average to 32 L/s (500 gpm). Water treatment of this flow can be effectively handled with Reverse Osmosis (RO) treatment. Treated water may be used in the process circuit (or discharged, as necessary – though not anticipated) and the brines from the RO will be evaporated/crystallized to form a solid salt; this salt, in turn, will be disposed of in the engineered and lined Salt Management Cells.

While water samples collected from these deep holes, NEC 14-014 and Met-1, and the follow-up investigation by Adrian Brown, indicate very similar quality, overall, water sampling results are variable across the site. This includes total dissolved solids which can range in concentrations of over 18,000 ppm, with the major contributors being sodium and chloride. Both of the wells noted above also exceed EPA primary Maximum Contaminant Levels (MCLs) with respect to the following:

● Arsenic;

● Gross alpha; and

● Ra-226 + Ra-228.

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Water from both of these wells also exceeds secondary MCLs with respect to chloride, fluoride, iron, manganese, sulphate, as well as total dissolved solids. NEC 14-014 also exceeds the secondary MCL for aluminum. There were no detectable pesticides or herbicides. Although the deep groundwater is not currently a drinking water source, concentrations were compared to drinking water standards as a reference to possible regulatory and management implications of groundwater disposal from future mine dewatering. Given the variability of water quality across the site, additional testing may be necessary to appropriately characterize the deep aquifer.

The deep groundwater chemistry data indicate a low-oxygen, chemically reducing groundwater system that is out of chemical equilibrium with surface conditions. Supporting evidence of this conclusion includes:

● Nitrogen species are mostly dominated by ammonia rather than nitrate, typical of highly reducing systems not exposed to the atmosphere.

● Iron is elevated at neutral pH, a condition which is unlikely to occur in an oxygenated, natural system.

● Groundwater brought to the surface at some boreholes is initially black, changes to orange over a time period ranging from hours to days, then eventually turns clear while forming an orange precipitate. This is characteristic of water initially containing reduced ferrous iron that eventually oxidizes after contact with atmospheric oxygen to ferric iron.

Further investigation is needed to determine the origin of the elevated concentrations in the groundwater, as well as refinement of the overall pumping requirements for the underground mining operation. Because of difficulties in handling these waters once they have been pumped to the surface, the additional testing remains a recommendation and must wait until surface management structures (ponds) and permitting have been completed.

17.1.9 Hydrology (Surface Water)

Surface water samples have been collected as part of baseline sampling on a periodic basis since early 2014. Surface water sampling locations were selected to establish a baseline monitoring perimeter both upstream and downstream from all proposed facilities in the Project area. All samples were analyzed by Midwest Laboratory in Omaha for a comprehensive suite of metals and other inorganic analytes plus a panel of pesticides and herbicides. Surface water quality data was collected by NioCorp using sampling procedures developed by SRK and NioCorp. These procedures included prescriptive methods for sample collection, preservation, chain of custody and transport to Midwest Labs. The laboratory maintains ISO/IEC 17025:2017 certification and incorporates appropriate QA/QC procedures in their analysis. The preliminary results of the baseline program are as follows:

● Surface water in and around the Project area exhibits minor water quality impairment, as indicated by concentrations outside the limits of several secondary drinking water standards and several aquatic life criteria (i.e., aluminum, iron, and manganese).

● Average stream TDS concentrations fluctuate appreciably; however, this variability is most likely the result of post-harvest runoff containing excess sediments.

● Stream pH is consistently circum-neutral, ranging from about 6.6 to 8.2 standard units.

● Gross alpha, beta, Ra-226 and Ra-228 have been detected in several surface water samples, but at concentrations below their respective EPA MCL.

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17.1.10 Wetlands/Riparian Zones

A wetland delineation was completed by Olsson for the Project in 2015. The wetland delineation identified wetland and drainage features within the proposed Project boundary. The study area consisted of agricultural fields, pastures, farmsteads and unnamed tributaries to Todd and Elk Creeks. All unnamed tributaries within the Project boundaries consisted of riparian areas and ponds that drained to Elk Creek. Many of the wooded areas not situated along drainages were located along fence lines as windbreaks. Most of the study area had been impacted by grazing livestock. The 2015 wetland delineation identified a total of 16 wetlands encompassing a total area of approximately 0.755 acres. One intermittent and two ephemeral channels were found during the field investigation for a total length of 8,887 feet. All three channels are unnamed tributaries to Elk Creek. Additionally, eight open water features were identified within the Project boundary, totaling 0.745 acre. Following the delineation, a USACE jurisdictional determination (File Number: 2015-00226-WEH) was approved for the project on September 6, 2016 and indicated many of the features located on the Project site as isolated and non-jurisdictional. The site layout was designed to avoid the wetlands and streams determined to be jurisdictional by the USACE.

17.1.11 Geochemistry

A geochemical characterization program for the mineralized material, waste rock, and tailings has been initiated by SRK for the Project. Preliminary results are provided in the following sections.

Niobium Mineralized Material

Preliminary results suggest that the mineralized material has the potential to leach various constituents due to exposure to meteoric precipitation. Laboratory leach tests of a composite sample of this material from drill hole NEC11-001 indicate that, at a minimum, fluoride and nitrate are likely to be mobilized during surface stockpiling. Note: fluoride is the only analyte that exceeds the EPA MCL in the leach testing. Nitrate and several metals are detectable, but not at concentrations exceeding their respective MCL for drinking water).

Contained within the mineralized material are naturally occurring uranium and thorium. Based on existing drilling data, the average thorium and uranium content in the Mineral Resource is 0.034% and 0.0045%, respectively (0.0395% in total). Leach testing of potential waste materials has not produced concentrations of these radionuclides above regulatory limits. However, the concentrations in the rock are relatively elevated (approximate relative concentrations):

● Uranium = 33 ppm;

● Thorium = 303 ppm;

● Gross alpha = 200 pCi/g;

● Gross beta = 160 pCi/g;

● Radium 226 = 56 pCi/g; and

● Radium 228 = 18 pCi/g.

The current assay database for the Elk Creek Project contains 6,288 samples for which uranium and thorium were analyzed and detected. Of this dataset, 1,122 samples (~18%) had a combined uranium+thorium concentration of greater than (“>”) 500 ppm. The mean and median concentration of uranium+thorium was 336 ppm and 273 ppm, respectively. The mineralized

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material suitable for mill feed will require proper management during the periods it is exposed on the surface, prior to processing.

Waste Rock and Overburden

There are two basic types of waste rock associated with the Deposit. These include:

● Pennsylvanian limestones and mudstones – The upper 30 m (100 ft) of lithology consists of unconsolidated glacial till, underlain by a 170 to 180 m (560 to 590 ft) of low-permeability, Pennsylvanian-aged mudstone and limestone, otherwise known as the “Pennsylvanian strata” (PENN). The PENN is reportedly continuous across the state of Nebraska, and locally it behaves as a very effective aquitard. This material is neutralizing due to its high carbonate content. In terms of metal leaching characteristics, Meteoric Water Mobility Procedure (MWMP) testing suggests that the PENN has the potential to leach antimony and selenium at concentrations above general surface water standards. Additionally, the PENN exhibits a propensity to leach gross alpha and radium above regulatory limits. This lithology is the primary source for construction aggregate in Nebraska.

● Non-ore grade carbonatite – Preliminary assessment of the host rock identified visual sulphide content of up to 1% based on observations by core loggers. Laboratory analyses confirmed the sulphide content at around 1.34%. This sulphide consists mainly of pyrite, chalcopyrite, bornite, galena, sphalerite, and possibly pyrrhotite. However, even with detectable sulphide content, the carbonatite waste rock is still net neutralizing given the high carbonate content.

● Of the 94 rock samples collected over a 255 m (837 ft) vertical length of the waste rock and mineralized zone, eight samples (8.5%) registered a reading of >25 µRads/hour. These levels are not considered to be hazardous but may be used as a diagnostic tool to identify elevated concentrations of uranium and thorium.

Temporary surface disposal of waste rock will be predicated on minimizing meteoric infiltration and leaching of this material. NioCorp has conservatively elected to line the waste rock and low-grade mineralized material stockpiles, and actively manage any runoff derived from these materials until such time as that, and residual ore and low-grade mineralized materials can be processed, and the surface waste rock transferred to the TSF for final disposal.

Tailings

Representative quantities of post-process tailings from the metallurgical testing program have been limited. Geochemical testing and characterization (including radiological testing) of the tailings was completed in Q3 of 2017 when the testing of the beneficiation process was finalized, and the need for, and usability of, tailings as underground backfill was evaluated. Characterization of the various tailings materials has included both the TCLP and the SPLP, which are designed to determine the mobility of both organic and inorganic analytes present in the liquid, solid, and multiphasic wastes, and assist in the proper classification of waste materials. The most recent tailings material testing showed negligible mobility of regulated constituents (indicating a non-hazardous classification), although the pH of the TCLP/SPLP extracts remained high. While the calcined tailings are likely to produce heat when exposed to atmospheric moisture and precipitation (i.e., exothermic hydration), this reaction is not “violent” as defined under 40 CFR § 261.23(2) Characteristic of reactivity [for hazardous wastes] (adopted by the State of Nebraska under Title 128 - Nebraska Hazardous Waste Regulations). Given the limited quantities of ore

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available for this testing, further characterization of these materials is recommended in order to establish representativeness with the mineral deposit as a whole.

17.1.12 Known Environmental Issues

There are currently no known environmental issues that are likely to materially impact NioCorp’s ability to extract the Mineral Resources or Mineral Reserves at the Elk Creek Project. However, there are several key permitting challenges and uncertainties associated with the dewatering and ground freezing program that may affect the Project financing and overall schedule. Risks are summarized in Section 22.12.2.

17.1.13 Tailings

NioCorp has chosen to design the solids portion of the TSF to include 0.61 m (2 ft) of compacted soil liner with a permeability of 1×10^-7 cm/s or less, overlain by an 80-mil HDPE liner, overlain by an overliner drain system. The water retaining portion of the facility will be lined with a double lined system consisting of a 60-mil HDPE secondary liner and 80-mil HDPE primary liner with an active leak detection system between. This conservative approach will likely ensure adequate protection of local groundwater resources. Additional details regarding the TSF are provided in Section 15.5. Closure of the TSF is discussed in Section 17.4.3.

17.1.14 Project Waste Disposal

Solid Waste

The solid waste generated by the Project, as defined by 40 CFR § 261.2, will be collected and transported to the Douglas County/Pheasant Point Landfill, located near Elk City in northwest Douglas County, 140 km (87 miles) from the Project site. Information from the waste management contractor indicates the landfill has at least 100 years of lifespan left.

Reject brines from the proposed RO water treatment plant are currently anticipated to be evaporated (crystallized) and the “solid” residue disposed of in the engineered and lined Salt Management Cells. Alternatively, these RO brines may be piped away from the mine site and re-injected into the deep underground aquifer, though this option still requires considerable evaluation before being considered viable.

Hazardous Waste

Any hazardous waste generated by the Project will be transported by licensed operators to the Clean Harbors Environmental Services facility in Deer Trail, Colorado, 756 km (470 miles) away, in accordance with hazardous waste manifest and pre-transport requirements.

17.1.15 Site Monitoring

Surface water and groundwater monitoring will continue throughout the LOM, as initiated during the baseline study program. Additional monitoring locations may be added during the regulatory review process. This will include, but not necessarily be limited to groundwater monitoring downgradient of the tailings storage facility and mine water collection pond.

Geotechnical monitoring of the TSF facility will also occur on a regular basis as per state regulatory requirements.

Ambient air quality monitoring will likely continue and may include emissions control monitoring once operations commence. This will be conducted in accordance with all applicable state regulatory requirements, which will be defined once NioCorp obtains an air operating permit.

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The presence of NORMs in the mineralized ore and several of the process waste streams will necessitate the need for comprehensive site-wide monitoring. At a minimum, the Broad Scope License will require the development and implementation of a formal Radiation Safety program for the facility, including environmental and personnel monitoring programs, which are discussed further in the following sections.

17.1.16 Water Management

Operational Water Management

For the first several years of construction, the advancement of the shaft and underground workings will require limited dewatering, anticipated to be through lower-level sumping and pumping for surface collection and disposal. Initially, water will be stored in the lined Salt Management Cell #1 during construction or will be trucked off-site for treatment at a local publicly owned treatment works. Excess water in the Salt Management Cells will be spray evaporated within the footprint of the Cell, to avoid the reintroduction of soluble salts into the water treatment system. Temporary on-site storage or off-site shipment and disposal of the crystallizer solid waste may be necessary until construction of the Salt Management Cells is completed.

Once full operations commence, NioCorp anticipates a shortfall of approximately 233 L/s (3,700 gpm) of operational and processing water, as the underground mine dewatering is expected to produce an average of 63 L/s (1,000 gpm) of total discharge, made up of a LOM average of 31.5 L/s (500 gpm). To make up this shortfall, NioCorp proposes the following sources for additional water:

1. Tecumseh Board of Public Works water supply line (~2,000 gpm) – Tecumseh Board of Public Works, which maintains the infrastructure and supplies residential and commercial users in the City of Tecumseh, might run a line to the project site to supply all of the necessary shortfalls.

2. Local Landowner Well #1 (~500 gpm) – A new well on a local landowner’s property has the potential to supply up to 500 gpm of the project’s needs. Because there will be a transfer of water from one property to another, a Groundwater Transfer Permit will need to be issued by the Nemaha Natural Resources District pursuant to Chapter 11 of the Management Area Rules and Regulations for Groundwater Quantity Management Areas.

3. Local Landowner Well #2 – NioCorp has the option to connect to an existing well as well as install a new well to supply an additional 1,500 gpm.

NioCorp is pursuing approval of all three sources as insurance that there are no disruptions in the water supply during operations. None of the permitting for these alternative water sources is considered particularly onerous or time-consuming

Once tailings begin being deposited in the TSF, internal contact water (from residual moisture in the tailings and precipitation falling within the impoundment footprint) will need to be actively managed. This water will be collected and treated using lime softening to precipitate hydroxide and carbonate solid forms for many of the inorganic constituents. The treated water will be filtered to remove the solids (which will be returned to the TSF for disposal), and the clean water will be pumped to the process plant RO system for further treatment. The clean water from the process plant RO unit will be used in the process plant, and the reject concentrate will be crystallized and deposited back into the Salt Management Cells.

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Post-Closure Water Management

Upon cessation of mining, the limited subsurface mine water pumping operations will be halted, and the workings will be allowed to flood. Until such time that the TSF closure cover can be constructed, and any residual water or seepage eliminated, the TSF contact water will require active management. Whether the singular TSF brine stream from the RO plant can continue to be crystallized and deposited in the Salt Management Cells or if another disposal method needs to be considered (i.e., disposal in the deep mine workings or in an off-site disposal facility), it will be evaluated during the final years of operation.

17.1.17 Chemical and Reagents Handling

Process reagents and chemicals, including but not limited to: sulphur (molten), sodium hydroxide (NaOH), magnesium hydroxide (MgOH), lime (CaO), aluminum, iron (scrap and powder), boiler feed chemicals, water treatment plant chemicals, cooling tower chemicals, and solvent extraction circuit chemicals, will be trucked to the site and stored in specially designed and constructed containers located within concreted and concrete-bermed areas. For liquid chemicals and reagents, these bermed areas will be designed to contain at least 110% of the capacity of the largest storage tank or tanks in series within the berm. Solid chemicals and reagents will be stored in flow bins or silos specifically designed for these materials. Reagents will be stored in a manner that inhibits any inter-mixing and subsequent reactions.

Fuel (i.e., gasoline, diesel fuel, and propane), antifreeze, petroleum oils, and solvents will be delivered to the mine in tanker trucks, totes and barrels for transfer to authorized storage tanks. Storage tanks or tanks in series will be enclosed by berms sized to contain at least 110% of the capacity of the largest tank in the event of a spill or tank rupture. NioCorp will develop a comprehensive Spill Prevention Control and Countermeasures Plan (SPCC) to be implemented in the event of a spill or release of petroleum products.

Explosive materials transported to the site will include blasting agents and initiation devices. Blasting agents are comprised primarily of ammonium nitrate and fuel oil. The ammonium nitrate and fuel oil will be stored in appropriate storage bins separate from the explosives magazine. Blasting initiation devices will be stored in prefabricated magazines in conformance with U.S. Bureau of Alcohol, Tobacco and Firearms (BATF), MSHA, and applicable state and local regulations.

17.2 Project Permitting Requirements

Engagement of local, state and federal regulators has commenced. Initiation of the balance of permitting for the Project is dependent upon the completion of the mine plan and surface facilities being developed as part of this technical document. Typically, larger mining operations such as this have the benefit of a pre-feasibility stage of analysis and development from which permitting is generally initiated. With the completion and publication of this Feasibility Study, the balance of permitting for the Project can commence.

The Project has considered and will likely be held to permitting requirements that are determined to be necessary by Johnson and Pawnee counties, the State of Nebraska, and the U.S. Environmental Protection Agency and USACE national policies, such as the National Environmental Policy Act (42 U.S.C. 4321) and the Clean Water Act (33 U.S.C. 1251 et seq.). The list of potentially applicable permits and authorizations for the Project are presented in Table 17-1.

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Table 17-1: Project Permits

Permit/Approval	Issuing Authority	Permit Purpose	Status
Federal Permits Approvals and Registrations
Explosives Permit	U.S. Bureau of Alcohol, Tobacco and Firearms (BATF)	Storage and use of explosives	Mine Safety and Health Administration (MSHA) and the Department of Homeland Security (DHS) will also regulate explosives at a mining operation.
EPA Hazardous Waste ID No.	U.S. Environmental Protection Agency (EPA)	Registration as a Conditionally Exempt Small Quantity Generator (CESQG) or a Small Quantity Generator (SQG) of waste	NioCorp laboratory facilities are likely to generate small quantities of hazardous waste.
Spill Prevention, Control, and Countermeasure (SPCC) Plan	U.S. Environmental Protection Agency (EPA)	Regulation of facilities having an aggregate aboveground oil storage capacity greater than 1,320 gallons or a completely buried storage capacity greater than 42,000 gallons with a nexus to jurisdictional waters	REQUIRED. Adjacent jurisdictional drainages.
Notification of Commencement of Operations	Mine Safety and Health Administration (MSHA)	Mine safety inspections, safety training plan, mine registration	REQUIRED. All mining operations in Nebraska.
Obstruction Evaluation / Airport Airspace Analysis (OE/AAA)	Federal Aviation Administration (FAA)	Notification of the Administrator of the FAA for any construction or alteration exceeding 200 ft above ground level.	REQUIRED: If any project components exceed 200 feet in height.
Federal Communications Commission Permit	Federal Communications Commission (FCC)	Frequency registrations for radio/microwave communication facilities	REQUIRED. If NioCorp intends to use business radios to transmit on their own frequency.
State Permits, Authorizations and Registrations
Permit to Appropriate Water	State of Nebraska Department of Natural Resources (DNR)	Regulates the use and storage of surface and ground waters	REQUIRED to appropriate water.
Explosives Permit	Nebraska State Patrol	Regulates the use, storage, or manufacture of explosive materials.	REQUIRED. Also regulated by BATF, MSHA, and DHS.
Permit to Discharge under the National Pollutant Discharge Elimination System (NPDES)	State of Nebraska Department of Environmental and Energy (NDEE)	Multiple permits are applicable to the discharge of industrial wastewater and stormwater.	REQUIRED. The project will require construction and industrial stormwater discharge permit. The project will not discharge wastewater. A permit will be required to operate the facility’s water treatment system.
Mineral Exploration Permit	NDEE	Regulates the exploration for minerals by boring, drilling, driving, or digging.	REQUIRED. Already obtained for the exploration drilling program.
Air Construction Permit	NDEE	Regulates emissions during construction activities to protect ambient air quality.	REQUIRED. Under Nebraska Administrative Code (NAC) Title 129. Permit was applied for and was issued by the NDEE on June 2, 2020.
Air Operating Permit	NDEE	Regulates emissions during operation to protect ambient air quality. Will be based on a Feasibility Study mine plan.	REQUIRED. Class I (Title V) federal major source PSD operating permit will likely be required as per NAC 129. Application required no sooner than 1 year after operations commence
Water Well Installation Declaratory Ruling Request	Nebraska Department of Health and Human Services, Division of Public Health	Water well installation requirements; well must be registered with the Department of Natural Resources.	REQUIRED. Already obtained for the hydrogeological portion of the exploration drilling program.
Authorization for Class V Well Underground Injection	NDEE	All activities conducted pursuant to Title 122 - Rules and Regulations for Underground Injection and Mineral Production Wells.	REQUIRED. Already obtained for the hydrogeological portion of the exploration drilling program. Will also be required for future disposal of tailings and/or crystalized RO brine gels in underground workings.
Septic Systems – Permit for Onsite Wastewater Treatment System Construction/Operations	NDEE	Protects surface water and groundwater as well as public health and welfare through the use of standardized design requirements.	REQUIRED. Needed if the septic system does not meet the “Authorization by Rule” requirements due to the quantity or quality of the wastewater, as per NAC 124.
Boiler Inspection Certificate	Nebraska Department of Labor	Protects public safety through an inspection and approval process of boilers.	REQUIRED. For installation of the boiler(s) is installed in any of the facility buildings.
Section 401 Water Quality Certification	NDEE	The program evaluates applications for federal permits and licenses that involve discharge to waters of the state and determine whether the proposed activity complies with NAC Title 117- Nebraska Surface Water Quality Standards. Isolated wetlands are included in NAC Title 117.	NOT REQUIRED. Only required as part of Section 404 authorization. Not currently anticipated.
Development Permit	NDEE/Johnson County Floodplain Administrator	The program regulates building requirements for any structures that are constructed on a floodplain.	REQUIRED. Will be needed if NioCorp constructs any building on a designated floodplain.
Fire and Life Safety Permit	Nebraska State Fire Marshall	Review of non-structural features of fire and life safety.	REQUIRED. Project proponent to submit operating and building plans. State Fire Marshall will then determine required inspections as per NFPA 101.

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Permit/Approval	Issuing Authority	Permit Purpose	Status
State Business License	Nebraska Secretary of State	License to operate in the state of Nebraska.	REQUIRED. All business entities in Nebraska.
Retail Sales Permit or Exemption Certificate	Nebraska State Tax Commissioner	Permit to buy wholesale or sell retail.	MAY BE REQUIRED. Will be required if NioCorp is direct selling niobium product.
Solid Waste Management Permit	NDEE	Regulates the construction and operation of solid waste management facilities.	REQUIRED. Will be needed if NioCorp intends to create an on-site solid waste management facility. This may include the TSF and salt impoundments.
Drinking Water Construction Permit	Nebraska Department of Health and Safety	The Drinking Water Construction Permit regulates the design and construction of a public water system.	MAY BE REQUIRED. All drinking water systems that serve more than 25 individuals and are considered to be “non-transient and non-community” are required to obtain a Drinking Water Construction Permit. This will include the use of RO permeate produced at the plant site.
Drinking Water Permit to Operate	Nebraska Department of Health and Safety	Defines testing and water quality criteria for public drinking water systems.	MAY BE REQUIRED. All drinking water systems that serve more than 25 individuals and are considered to be “non-transient and non-community” are required to obtain a Drinking Water Permit to Operate.
Radioactive Materials Program and Licensing	Nebraska Department of Health and Human Safety	Regulates and inspects users of radioactive materials.	REQUIRED. If the plant uses sealed sources for process measurements or if naturally occurring, radioactive materials are possessed as a result of beneficiation activities.
Hazardous Waste Management	NDEE	Management and recycling of hazardous wastes.	REQUIRED. As per Title 128 of the Nebraska Hazardous Waste Regulations NioCorp must notify the NDEE of hazardous wastes generated or transported from the facility.
Dam Safety Approval	State of Nebraska DNR	Regulates the design and construction of any dam (i.e., any artificial barrier with the ability to impound water or liquid-borne materials).	REQUIRED. May be required for TSF (dam) and may be required for the Mine Water Pond depending on the final design capacity.
Water Storage Permit	State of Nebraska DNR	Regulates any water impoundment that has a normal operating water volume of at least 15 AF below the spillway.	MAY BE REQUIRED. May be required for the Mine Water Pond, if it will impound greater than 15 AF below the spillway.
Local Permits for Johnson and Pawnee Counties
Water Well Permit	Nemaha Natural Resources District	Regulates installation of groundwater wells	REQUIRED. This permit will be required to install a new water supply well.
Water Well Transfer Permit	Nemaha Natural Resources District	Regulates transfer of groundwater off overlying land	REQUIRED. This permit will be required to transfer water from wells located on a separate property to be used for water supply.
Building and Construction Permits	Johnson County Zoning Administrator	Ensure compliance with local building standards/requirements.	REQUIRED. This permit will most likely be included with the Permitted Use Zoning Permit
County Road Use and Maintenance Permit/Agreement	Johnson County Zoning Administrator	Use and maintenance of county roads.	MAY BE REQUIRED. Will be needed if NioCorp intends to maintain any of the area county roads.
County Road Use and Maintenance Permit/Agreement	Pawnee County Commission	Use and maintenance of county roads.	MAY BE REQUIRED. Will be needed if NioCorp intends to maintain any of the area county roads.
Special Use Permit	Johnson County Zoning Administrator	Regulates and authorizes permitted uses.	REQUIRED. Issuance of this permit will require completion on an application form, and at least one meeting with the county zoning regulators and at least one public comment meeting. Permit was issued to the Company on December 24, 2019
Special Use Permit	Pawnee County Assessor	Regulates and authorizes permitted uses	REQUIRED. TSF land currently zoned for agriculture. Zoning regulations allow for mineral extraction.
Permitted Use Zoning Permit	Johnson County Zoning Administrator	Regulates the construction of new buildings	REQUIRED. Application must be submitted 5 days in advance of the start of construction

Source: SRK, 2017

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The following is a brief discussion of some of the more material permits which are likely to be required for the project (Note: with respect to the Underground Injection Control (UIC) permit, the discussion is included only as an alternative to the planned treatment and disposal of excess water).

17.2.1  Nebraska Underground Injection Control (UIC)

In the event that crystallization of the RO water treatment brines becomes impractical, NioCorp may alternatively opt to reinject the reject waters back underground. This activity will, necessarily, require a permit. The UIC Program of the NDEE Water Division issues and reviews permits, conducts inspections and performs compliance reviews for wells used to inject fluids into the subsurface. The program must ensure that injection activities are in compliance with state and federal regulations, and that groundwater is protected from potential contamination. Injection wells are classified by activity, with most activity concentrating on Class I, II, III, and V wells. Class II wells are associated with oil and gas production and are regulated by the Nebraska Oil and Gas Conservation Commission. NDEE has authority over and manages, Class I, III and V wells. A water treatment system brine re-injection well is likely to be a Class V well.

The EPA delegates the UIC program to the NDEE and provides authority for the program through the Safe Drinking Water Act. The Natural Resource Districts across the state have also developed sets of rules and regulations (NDNR) regarding permitting requirements and the installation of wells based on specific Groundwater Management Plans, and the NDNR requires that all wells installed in the state must be registered. Additionally, the NDNR is charged with issuing permits for industrial use of groundwater.

17.2.2  DHHS Radioactive Materials Program and Licensing

The Elk Creek Mineral Resource, and thus the residual post-processing tailings, will contain trace amounts of uranium and thorium, which are Naturally Occurring Radioactive Materials (NORM). At issue will be the ultimate classification of the tailings because of these constituents, and the occurrence of these constituents in the processing circuit. Preliminary discussions with the State of Nebraska have indicated that either a Specific or Broad Scope Radioactive Materials License, issued under 180 NAC 3-013 by the Nebraska (DHHS), will likely be necessary, as confirmed with the DHHS on December 6, 2018.

As defined by the Nebraska Radiation Control Act, radioactive material means any material, whether solid, liquid, or gas, which emits ionizing radiation spontaneously. Radioactive material includes but is not limited to, accelerator-produced material, by-product material, naturally occurring material, source material, and special nuclear material. The classification of radioactive material appears to be irrespective of any concentration – it merely has to emit ionizing radiation. The material for processing, waste rock, and tailings are likely to be seen as naturally occurring material, and therefore, classified as a radioactive material.

The DHHS retains the right to require registration or licensing of [any] radioactive material in order to maintain compatibility and equivalency with the standards and regulatory programs of the federal government or to protect the occupational and public health and safety and the environment [NRS 71-3507(2)]. At the same time, the DHHS can exempt certain sources of radiation or kinds of uses or users from licensing or registration requirements when the department finds that the exemption will not constitute a significant risk to occupational and public health and safety and the environment [NRS 71-3507(4)].

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At a minimum, the Broad Scope License will require the development and implementation of a formal Radiation Safety program for the facility, including environmental and personnel monitoring programs, appropriate warning signage be displayed around the site, and a final permanent closure cover for the TSF be engineered and constructed. DHHS oversight and the Broad Scope License will necessarily cover all points of potential worker exposure, including but not limited to underground mining, crushing, transportation and stockpiling, conveying, and processing, especially in areas were airborne dust containing uranium and thorium (as well as radon gas) can occur. Worker protection from ionizing radiation and radon will also be regulated by the U.S. Department of Labor, Mine Safety and Health Administration (MSHA) under 30 CFR PART 57 – Safety and Health Standards – Underground Metal and Nonmetal Mines, Subpart D – Air Quality, Radiation, Physical Agents, and Diesel Particulate Matter. Both programs will examine potential exposure limits, engineering and administrative control requirements, the use of appropriate Personal Protective Equipment (PPE), and monitoring/reporting programs to ensure worker protection.

In the likely event that the Elk Creek facility is regulated in this way, some land restrictions may be invoked at the time of mine closure. While these requirements appear to be directed at uranium mills and commercial radioactive waste disposal facilities, and not necessarily mine tailings for operations containing NORM, the law makes no clear distinction between the facility types; the State of Nebraska may apply them under either scenario, which might even include the possibility of deeding the land to the State of Nebraska following closure.

Irrespective of ultimate classification, the tailings (and their disposal facility) will require financial assurance for reclamation and closure. Again, these rules appear to be directed at uranium mill tailings and low-level radioactive waste facilities but are non-specific enough that they may be applied to other situations where NORMs are being actively managed. In addition to a direct reclamation financial assurance, it is probable that the state will require a funding mechanism (i.e., trust fund, escrow, etc.) for monitoring and maintenance of the facility in the longer term as part of a Broad Scope License.

DHHS License Timing

NioCorp estimates that a Broad Scope License for the Project will take approximately 16 months to acquire once the formal application has been submitted and will involve several months of discussions and negotiations related to engineering, design, monitoring, and terms and conditions. At this time, the federal Nuclear Regulatory Commission shall play a purely advisory role in these negotiations.

17.2.3  Nebraska Air Quality Permitting

The Nebraska air regulations are primarily based on regulations developed by the U.S. EPA to address the Clean Air Act requirements. Air quality permits are the primary tool used by the NDEE to implement the Clean Air Act. For businesses that intend to operate unit sources that emit regulated pollutants that will exceed Nebraska air quality thresholds, a construction permit will be required.

There are two types of construction permits: state construction permits and federal construction permits, known as New Source Review or Prevention of Significant Deterioration (PSD) permits. The type of construction permit that is needed will depend on the quantity of air pollutants that potentially may be released from the new plant or expansion project. Given the emissions profile

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of the project, a state construction permit was required and was obtained by the Company on June 2, 2019.

In addition to the construction permit, the NDEE also issues operating permits based on a source’s level of emissions. There are two types of operating permits: major source (federal program) and minor source (state program). As before, the potential to emit associated with the sulphuric acid plant will necessitate the issuance of a major source permit for the operation. The federal major source program (a.k.a., Class I or Title V) regulates larger sources of air pollution. A Class I source has the potential-to-emit quantities greater than:

● 100 t/y of any criteria air pollutant, excluding lead;

● 10 t/y of any single hazardous air pollutant (HAP) or 25 t/y of a combination of HAPs; or

● 5 t/y of lead.

The operating permit incorporates all of a source’s requirements into one permit, including all construction permit limitations and federal regulations. Operating permits usually require additional monitoring, stack testing, reporting, and recordkeeping. However, the application for the operating permit need only be submitted within 12 months after the emissions unit(s) begin operation, or within 12 months of becoming subject to the operating permit requirements, whichever is earlier.

Earthworks associated with digging holes, grading soil, stockpiling of topsoil, and land clearing where the new source will be located, which will not result in a change in actual emissions, and are not of a permanent nature, do not require a construction permit or prior approval of the NDEE under Title 129, Chapter 17 (Acceptable Pre-Construction Dirt Work dated August 2016).

17.2.4  Nebraska Dam Permitting

The Department of Natural Resources (DNR) regulates the construction, operation, and maintenance of dams in Nebraska to protect life and property from dam failures. The DNR regulates all dams in the state that:

● Have a total height of 7.62 m (25 ft) or more and an impounding capacity at the top of the dam that is greater than 1.85 hectare-15 (15 acre-ft);

● Have an impounding capacity at the top of dam of 6.17 hectare-metres (50 acre-ft) or more and a total height that is greater than 1.8 m (6 ft); or

● Are located in a high hazard potential location.

As promulgated in Chapter 46, Article 16 - Safety of Dams and Reservoirs, approval of applications shall be issued within 90 days after receipt of the “completed” application plus any extensions of time required to resolve matters diligently pursued by the applicant. At the discretion of the DNR, one or more public hearings may be held on an application (46-1654). This will, of course, add additional time to the overall permitting process for the TSF and Mine Water Pond.

17.2.5  Greenhouse Gas Permitting

The NDEE defines Greenhouse Gases (GHG) as chemical compounds that, when emitted into the atmosphere, have the potential to cause climate change. There are currently 73 GHG chemicals identified in 40 CFR § 98 Table A-1 to Subpart A, which include, but are not limited to CO₂, CH₄, N₂O, and Fluorinated GHGs (SF6, PFCs, HFCs). Recent rulemaking by the EPA incorporates changes impacting the regulation of GHGs and establishes emission thresholds for GHG emissions, while

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provides the State of Nebraska (among others) the authority to issue PSD permits governing GHGs.

To date, the EPA has not implemented a minor source program for GHGs, and Nebraska has not chosen to implement a minor source program either. At this time, no fees will be collected, but all sources will be required to report GHG emissions.

17.2.6  Permitting Status

Initial permitting activities commenced in January 2015 with the submission of a Jurisdictional Delineation report to the USACE for the mine site. In addition, several high-level meetings with federal, state and local agencies have been held in order to introduce the Project to the local regulatory communities.

NDEE Construction Air Permit

A pre-application meeting took place with the NDEE on September 8, 2016. A formal application for a Construction Air Permit was submitted on July 24, 2019, and the permit was issued by the NDEE on June 2, 2020.

Special Use Permit

Johnson County, Nebraska issued a Special Use Permit for the project on December 24, 2019. The permit authorizes the land use for the construction and operation of the Project.

Temporary Limestone Processing

The Project may require temporary limestone processing during the construction of the mine shaft. Third-party portable limestone processing equipment may be used on site to crush and handle limestone removed from the mine shaft, so long as that material meets construction specifications and does not leach potentially deleterious constituents (i.e., heavy metals or NORMs). The NDEE has confirmed that third-party operators will be required to have an air quality permit to operate equipment on site.

17.2.7  Post-Performance and Reclamation Bonding

In addition to lacking hardrock mining regulations for reclamation and closure, there are also limited requirements for the provision of financial sureties with respect to hardrock mining operations in Nebraska. One possible exception may include the scenario in which the facility falls under a broad scope radiological license, which has financial assurance requirements for reclamation and closure (“decommissioning funding plan”). As noted before, however, these rules appear to be directed at uranium mill tailings and low-level radioactive waste facilities, but are vague enough that they may be applied to other situations where NORMs are being managed, though NioCorp has conservatively assumed that the licensure program and financial surety requirements will apply to the Project. These surety requirements extend to long-term site monitoring, maintenance, and care, and include the following mechanisms:

● Pre-payment (Trust Fund)

● Surety Bond

● Insurance

● Letters of Credit

● Corporate Guarantee (provided parent company passes the financial test)

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In addition, financial assurances will also be required for the TSF, for which jurisdiction will fall under the NDEE Title 132 - Integrated Solid Waste Management Regulations, and includes the requirement for a detailed, third-party closure cost estimate, proper disposal of all materials or wastes left at the site, and post-closure care for the solid waste disposal area in compliance with the post-closure plan. Allowable mechanisms for financial assurance under the solid waste regulations include:

● Trust Funds

● Surety Bonds Guaranteeing Payment or Performance

● Letters of Credit

● Insurance

● Corporate Financial Tests

● Local Government Financial Tests

● Corporate Guarantees

● Local Government Guarantee

At this time, the type and phased amount of financial surety for the Project has not yet been established, though the amount of bond will only reflect the liability on the ground at any given time (i.e., NioCorp is not likely to be required to bond for reclamation of all of the TSF cells when only one will be active and un-reclaimed at any time). The specific requirements will be refined through meetings and negotiations with the two agencies and the submission of formal permit applications. The company met with the NDEE and presented the reclamation cost estimates in detail on February 7, 2018, and subsequently provided the NDEE with detailed reclamation cost estimate calculations for their review.

17.3  Community Relations and Social Responsibilities

Community relations and stakeholder engagement have been undertaken in parallel with field operations in Nebraska and have included town hall and individual meetings with local landowners. Some early communications have occurred between NioCorp and Johnson, Pawnee, Nemaha and Richardson County representatives (including the county commissioners) as well as the Southeast Nebraska Development District. Given the schedule proposed by NioCorp for the Project, all of the relevant regulatory agencies will need to be formally engaged as soon as possible using the designs presented herein as the basis for permitting. Any significant deviations from this design may have an impact on overall Project timing.

NioCorp is committed to ensuring that a proper Social License is garnered from the community and stakeholders. Thus far, support for the Project has been positive from those who have been engaged and notified of the pending Project. However, as with any major mining project, there remain vocal opponents and non-governmental organizations (NGOs) who will oppose the Project on principal alone. These groups are likely to include organizations such as Bold Nebraska, a citizen group focused on “taking actions critical to protecting the Good Life.” NioCorp has already engaged with Bold Nebraska in early discussions about the Project on May 23, 2016 and has kept the group informed of major developments.

17.3.1  Safety and Health

Occupational Safety and Health at the Project will be strictly regulated by the U.S. Department of Labor, Mine Safety & Health Administration (MSHA), under Title 30 of the Code of Federal

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Regulations, Mineral Resources, Parts 1 through 199 (30 CFR Parts 1 through 199). This includes all of the training requirements specified in 30 CFR Parts 46 through 49. Given the radiological nature of the mineralized material, MSHA will likely institute radon exposure and monitoring requirements on all underground workers in accordance with 30 CFR § 57.5039 thru § 57.5047.

Because Nebraska has not enacted any workplace safety and health rules, the federal Occupational Safety and Health Act (OSH Act) governs workplace health and safety requirements in private (private businesses and non-profit organizations) sector workplaces. In addition, the Nebraska Occupational Safety and Health Surveillance Program (NOSHP), established in 2010 under the Nebraska Department of Health & Human Services, provides state-based occupational health surveillance, while the Nebraska Department of Labor (DOL) Office of Safety is charged with the protection of people and property through enforcement of the Nebraska Amusement Ride, Boiler Inspection, and Conveyance Safety Acts. With respect to the Project, DOL safety staff will inspect boilers and pressure vessels to ensure that they are properly installed and maintained.

To facilitate safe working environments within the proposed operation, MSHA safety standards are incorporated in the mine design and include dual secondary means of mechanical egress, backup power for both auxiliary hoists, partial ventilation system, and one air compressor which feeds compressed air to the underground. Twelve-person mobile refuge chambers are included and will be in active working areas over the LOM. In addition, construction of the mine will include two permanent 30-person refuge chambers. The mine will have a communications system that has both mine phones and wireless communication. A mine rescue team will be required to support the mine’s underground operation. The mine safety program will integrate with local providers in case of any mine emergency. Additionally, a stench gas emergency warning system will be installed in the mine’s intake ventilation system. This system can be activated to warn underground employees of a fire situation or other emergency whereupon emergency procedures will be followed. The shop areas and underground fueling station will be equipped with automatic closure doors that will operate in case of fire (Nordmin 2019). NioCorp will follow all applicable laws and regulations regarding site specific safety measures, including things such as a Fire and Life Safety Permit, Boiler Inspection Certificate, and MSHA safety inspections.

17.4  Reclamation and Closure

Without specific hardrock mining regulations, there are limited obligatory requirements for reclamation and closure of mining properties in Nebraska. There are provisions, however, within the applicable regulatory framework which are likely to be applied to the Project during the permit and licensing processes, specifically those associated with the TSF. The following sections provide a summary of the key elements to the approaches proposed for closure and reclamation of the Project and form the basis for the closure cost estimate.

17.4.1  Surface Disturbance

The principal objective of the surface reclamation plan will be to return disturbed lands to productive post-mining land use. Soils, vegetation, wildlife and radiological baseline data will be used as guidelines for the design, completion, and evaluation of surface reclamation. Final surface reclamation will blend affected areas with adjacent undisturbed lands so as to re-establish original slope and topography and present a natural appearance. Surface reclamation efforts will strive to limit soil erosion by wind and water, sedimentation, and re-establish natural drainage patterns.

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17.4.2  Buildings and Equipment

All surface structures and equipment will be evaluated for appropriate post-closure re-use or disposal. Buildings and equipment will be decommissioned, decontaminated (as necessary), dismantled, and either salvaged or disposed of in an appropriate on-site or off-site disposal facility.

All wells, including dewatering and production wells, monitoring wells, and any other wells within the Project Area used for the collection of hydrologic or water quality data or incidental monitoring purposes, will be properly abandoned in accordance with NDEE and DNR requirements.

17.4.3  Tailings Disposal Facility

Since the definition of Solid Waste in Chapter 1 of Title 132 – Integrated Solid Waste Management Regulations includes material generated from mining operations, the Tailings Storage Facility (TSF) and the Salt Management Cells at the Project will likely be subject to all or part of the Title 132 regulations, including the closure requirements. The design of the TSF cells allows for concurrent reclamation in order to reduce the amount of precipitation contact water that will require active management. Once a cell of the TSF has reached design capacity, it will be closed. For purposes of closure cost estimating and potential future bonding requirements, this approach will assume that only one cell will be active at any given time for which reclamation (and bonding) may be required. In addition, the approach to TSF construction and material placement will allow the operator to concurrently close portions of each cell as they reach capacity.

The initial closure cover will consist of surface grading and placement of a geomembrane liner over the graded tailings. This liner requires an over-liner drainage system that discharges to the outer slope of the embankment of each TSF cell, and placement of adequate thickness of cover to allow for vegetation; though a root barrier may be necessary to prevent rooting into the tailings. With respect to post-closure requirements, operators of solid waste disposal areas shall provide for post-closure care for a period of at least 30 years. At this time, there is no anticipated post-closure solution/draindown management consideration for the TSF cells given the nature of the tailings materials and the conceptual closure approach. This approach to the closure of the TSF cells is considered conservative and was selected to demonstrate the feasibility and permit ability with respect to the NDEE landfill regulations and on the advice of the agency. Given the current LOM expectation, additional technologies and/or approaches to equally effective closure options may likely be developed prior to actual reclamation of the site.

The Salt Management Cells will be closed in a manner similar to the TSF.

17.4.4  Closure Cost Estimate

Direct reclamation and closure costs for the Project, including estimates for post-closure monitoring and maintenance, were estimated at approximately US$ 44.7 million in 2019. Including financial assurance premiums for the first five years of operations brings the total to US$ 50.2 million. This conservative approach and estimate consider the fact that 1) none of the facilities are constructed (i.e., final actual configurations are unknown), 2) costs for materials and services are difficult to predict 30 years in advance, and 3) no trade-off studies or final risk assessments have been performed on the closure approach (normally done later in the LOM).

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17.5  International Standards and Guidelines

The United States is a Designated Country with respect to the Equator Principles. Designated Countries are those countries deemed to have robust environmental and social governance, legislation systems, and institutional capacity designed to protect their people and the natural environment (Equator Principles Association, 2020).

The current version of the Equator Principles (EP IV) was launched in July 2020 and became effective as of October 1, 2020. This version of the Equator Principles requires the same assessment and management structures for projects whether they are in Designated or Non-Designated (i.e., developing) countries. The Project has developed an Environmental and Social Management System that conforms to the 2020 EP IV requirements.

17.6  Qualified Person’s Opinion

It is Olsson’s opinion that NioCorp has adequately addressed environmental compliance, permitting and the concerns of local individuals to this point in the project’s development.

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18.  CAPITAL AND OPERATING COSTS

Capital and operating cost estimates were prepared by Tetra Tech, Adrian Brown Consultants, L3 Process Developments, Optimize Group and Metallurgy Concept Solutions with contributions from NioCorp. These estimates have been reviewed by Dahrouge.

18.1  Capital Cost Estimate

18.1.1  Basis of Estimate

The estimate meets the classification standard for a Class 3 estimate as defined by AACE international and has an intended accuracy of ± 15%. The estimate is reported in Q1 2019 U.S. constant dollars. The primary purpose of this report is to address changes to the resource estimate to include contained rare earth elements (REE’s). A subsequent addition of the REEs to the mineral reserve and economics will require additional metallurgical work. Costs have for the most part been retained from the previous 2019 Feasibility Study.

The capital cost estimate reflects a detailed bottom-up approach that is based on key engineering deliverables that define the Project scope. This scope was described and quantified within material take-offs (MTO’s) in a series of line items. Capital costs are divided among the areas of underground mining, processing, infrastructure, water management, tailings management, mining indirects (indirect costs), and contingency. Sustaining capital costs are related to underground mining fixed equipment and development, process plant, infrastructure maintenance, tailings management, mine closure and contingency.

18.1.1.1    Mining, Process, and Infrastructure Capital Costs

The mining capital costs were developed, including a combination of vendor and contractor quotations, first principles buildup, allowances, and historical database costs. The estimates include labour, materials, fixed equipment purchase and operation cost, rental equipment, supplies, freight, and energy. The costs developed include direct and indirect costs and included separate contingencies on both. Fixed equipment-purchase costs include freight, an allowance for transporting underground, initial training and commissioning.

18.1.1.2    Tailings and Tailings Water Management Capital Costs

The capital cost for tailings facility construction was based on contractor estimates for earthworks and liner installation. A local equipment supplier quoted equipment for loading, hauling and placement of the tailings. SRK developed some costs internally for items where no quotes were obtained. The SRK estimates were developed from recent and relevant costs on other projects or developed from first principles. Approximately 10% of the tailings facilities costs were from SRK estimates.

18.2  Capital Cost Summary

Table 18-1 shows the breakout in LOM initial and sustaining capital estimates, which total US$ 1,607 million. An overall 9.79 % contingency factor has been applied to the initial capital estimate, while a smaller 2.06 % contingency was applied to the sustaining capital estimate. The pre-production period is defined from April 2022 to the end of construction in June 2025 plus a six-month ramp-up period through the end of December 2025. Commercial production is then to be declared on January 1, 2026. The initial capital estimate of US$ 1,141 million will be partially offset

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by a Gross Pre-production Revenue Credit of US$ 257 million, (generated by pre-production product sales) which equates to a net cost of US$ 884 million.

Table 18-1: Capital Costs Summary (US$ 000’s)

Description	Initial	Sustaining	Total
Capitalized Preproduction Expenses	$77,053		$77,053
Site Preparation and Infrastructure	$40,569	$15,007	$55,576
Processing Plant	$367,439	$96,448	$463,886
Water Management & Treatment	$73,756	$23,613	$97,369
Mining Infrastructure	$256,981	$198,482	$455,463
Tailings Management	$21,423	$78,855	$100,277
Site Wide Indirects	$7,368		$7,368
Processing Indirects	$96,028		$96,028
Mining Indirects	$41,130		$41,130
Process Commissioning	$13,350		$13,350
Mining Commissioning	$1,578		$1,578
Owner’s Costs	$33,619		$33,619
Mine Water Management Indirects	$8,520		$8,520
Closure and Reclamation		$44,267	$44,267
Contingency	$101,730	$9,385	$111,116
Total Capital Costs	$1,140,544	$466,058	$1,606,601
Preproduction Revenue Credit	($256,910)		($256,910)
Net Project Total	$883,634	$466,058	$1,349,692

Source: NioCorp, 2022

18.2.1  Capitalized Pre-production Costs

Pre-production costs are defined as production operating expenses that are incurred in the pre-production period before the declaration of Commercial Production phase.

For this study, costs were categorized as capital for taxation purposes, by which, per US federal tax rules, 70% of the annual cost can be expensed in the year incurred with the remaining 30% of cost amortized over next five years.

Table 18-2 shows the breakout between the different operating production costs incurred from April 2022 with the start of mine development activities and throughout the commissioning and ramp-up period from March 1, 2025, and December 31, 2025.

Consequently, all production operating costs incurred after the planned declaration of Commercial Production on January 1, 2026, are 100% expensed in the year incurred and fully deductible for taxation purposes.

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Table 18-2: Capitalized Pre-production Cost Summary (US$ 000’s)

Mining	$23,926
Processing	$42,897
G&A	$0
Other Infrastructure	$2,167
Water Management	$6,741
Tailings Management	$1,323
Total	$77,053

Source: NioCorp, 2022

18.2.2  Mining Capital Costs

Mining capital costs primarily comprise the following areas: shaft sinking, lateral mine development, and stationary/fixed mine infrastructure. It has been assumed that a shaft sinking and mine development/production contractor would be operating at the site from the beginning of the project to the end of mine life. The mine contractor would be responsible for sinking both shafts concurrently, developing the underground drifts, including the internal ramp, footwall and hanging wall access drifts, other underground mine infrastructure, the ventilation system and full production activities. The contractor would also develop all internal vertical development (ventilation raises, ore and waste passes). The Mining Capital costs were divided between direct and indirect costs.

The direct mining capital cost contribution is summarized in Table 18-3.

Table 18-3: Initial Direct Mine Capital Cost Estimate (US$ 000’s)

Surface Infrastructure	$114,404
Shaft and Structure	$53,817
Underground Development	$52,911
Underground Other	$31,902
Spares	$3,947
Subtotal	$256,981
Contingency	$25,885
Total	$282,866

Source: NioCorp, 2022

A further breakdown of each category is summarized as follows:

● The Surface Infrastructure includes:

○ Production shaft permanent hoist house

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○ Production shaft permanent headframe/collarhouse/bins

○ Ventilation shaft permanent hoist house

○ Ventilation shaft permanent headframe/collarhouse

○ Permanent surface mine ventilation systems

○ Surface material handling

○ Temporary generator farm

○ Shaft sinking Freeze Plant

○ Backfill plant

○ Mine dry/offices/warehouse

○ Mine electrical substations

○ Surface services

○ Surface site work

● The Shafts and Structures include:

○ Shaft geotechnical drilling

○ Shaft sinking setup at the production shaft and ventilation shaft

○ Temporary shaft sinking facilities (both shafts)

○ Shaft sinking for both the production shaft and ventilation shaft

● Underground Development includes vertical and lateral development during the pre-production phase of the mine.

● Underground Other includes underground mine ventilation, dewatering, material handling, garage/shops, and services.

● Spares includes all capital spare parts.

The contingency is based on a line-item analysis by category and averages 10.1% for the mine infrastructure capital. No contingency is included on sustaining capital. Table 18-4 shows the contingency by category.

Table 18-4: Initial Direct Mine Capital Cost Contingency Estimate

Category	Percent
Surface Infrastructure	10.66%
Underground Development/Other	9.93%
Shafts and Structures	9.78%

Source: NioCorp, 2022

Indirect Cost

The indirect mining cost is summarized in Table 18-5. A contingency of 10.6% was applied to the indirect cost. The indirect costs include detailed engineering, testing programs, EPCM, per diem, temporary power generation and distribution and energy costs.

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Table 18-5: Mining Indirect Cost

Category	US$ 000’s
Contractor Indirects	16,047
Owner Indirects	19,750
Diamond Drill Program	5,333
Pre-production Opex	17,209
Commissioning	1,578
Subtotal Mining	59,917
Contingency	6,369
Total Mining Indirect	66,286

Source: NioCorp, 2022

18.2.3  Processing Plant Capital Costs

The surface processing plant capital summarized in Table 18-1 is further broken down in Table 18-6.

Table 18-6: Process Plant Costs Summary

Item	US$ 000’s
Mineral Processing	24,871
Hydromet	243,700
Pyromet	22,341
Acid Plant	76,526
Total	$367,439

Source: Tetra Tech, 2019

Each category includes the following:

● Building, including foundation, structural steel, roofing, envelope, louvres, doors, elevated floors, control room, offices & electrical rooms, overhead cranes, etc.

● Building services, including ventilation, heating, plumbing, natural gas distribution, compressed air, etc.

● Mechanical equipment

● Chutes

● Dust collection equipment, including ducting

● Process piping

● Utility piping

● Protective coating on equipment & piping where applicable

● Electrical work

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● Instrumentation & control

18.2.3.1   Processing Indirects

The processing indirect capital cost summarized in Table 18-1 is further broken down in Table 18-7.

Table 18-7: Processing Indirect Costs Summary

Item	US$ 000’s
Detailed Engineering, Procurement & Construction Management (EPCM)	62,592
Other Professional Services Temporary Services	6,471
Construction Management Facilities other than EPC	439
Worker’s Lodging, Meals & Incidentals (Per Diem)	19,016
Early Operations & Construction Energy	269
Inventory and First Fills	6,233
Capital Spares	1,007
Total	96,028

Source: Tetra Tech, 2019

Each category includes:

● Detailed Engineering, Procurement & Construction Management (EPCM):

○ Detailed process engineering

○ EPCM contractor fee and expenses

○ EPC contractors’ fees and expenses

● Other Professional Services:

○ Hydromet process testing program

○ Software programming

○ Supplementary Geotechnical study

○ Surveying and quantity control

○ Quality control of fabrications

● Construction Management Facilities other than EPC:

○ Rental and installation of modular trailer offices

○ Office consumables

● Worker’s Lodging, Meals & Incidentals (Per Diem):

○ Construction workers: 90% of workers will come from outside the region and receive the Per Diem

○ Construction management and supervision personnel: all personnel will receive the Per Diem

● Early Operations & Construction Energy:

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○ Maintenance of temporary power distribution system

○ Operation and maintenance of the Water Treatment Plant, including chemical products

● Inventory and First Fills:

○ Two weeks consumption of reagents

○ Diesel and Fuel Gas tanks filled

○ Wear and tear, consumables store items, one set of each

○ Maintenance supplies and repair parts estimated at 50% of annual cost

○ Lubricant estimated at 25% of annual cost

● Capital Spares: High-Pressure Grinding Rolls

18.2.3.2   Process Commissioning

Process commissioning totals US$ 13.35 million and include the following:

● Pre-commissioning includes pre-operational verifications by contractors, vendors, and specialists, beginning three months before commissioning.

● Commissioning:

○ Commissioning by Operations personnel

○ Assistance from vendors

○ Commissioning spares and consumables

○ Assistance from engineering

○ Assistance from contractors.

18.2.4  Tailings Water Management and Salt Management Cells

Basis

Capital cost for tailings and salt management facility construction was based on contractor estimates for earthworks and liner installation. A local equipment supplier quoted equipment for loading, hauling and placement of the tailings. SRK developed some costs internally for items where no quotes were obtained. The SRK estimates were developed from recent and relevant costs on other projects or developed from first principles. Approximately 10% of the tailing’s facilities costs were from SRK estimates.

Initial Capital

Tailings Plant Site Cell 1 is located directly east of the process plant site. Salt Management Cell 1 is located west of the process plant site. Construction for this cell will occur during the summer and fall before the plant goes into production. Plant Site Cell 1 construction will consist of approximately 465,000 m³ of cut to fill earthworks and 111,000 m² of geomembrane installation for the tailings facility and associated stormwater pond. The estimated costs, including earthworks, project management, piping and geosynthetic costs, and stormwater management, are summarized in Table 18-10.

Salt Management Cell 1 will be constructed early in the construction schedule in order to be ready to receive salt from Water Treatment operations once they commence and to support

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hydrogeological testing. This cell will consist of approximately 446,400 m³ of cut to fill earthworks and 120,800 m² of geomembrane installation. A small salt holding facility will be constructed adjacent to the Water Treatment Plant to temporarily store salt in advance of hauling the salt to the Salt Management Cell. The estimated costs, including earthworks, project management, piping and geosynthetic costs, are summarized in Table 18-8.

Table 18-8: Pre-production Tailings and Salt Management Facility Construction Cost

Item	US$000’s
Earthworks	$10,952
Equipment	$3,036
Crystallized Brine Management	$3,636
Salt Storage Facility	$416
Salt Haulage	$313
Rock Storage Facility	$3,070
Subtotal	$21,423
Contingency	1,964
Total	$23,386

Source: SRK, 2019

In addition, the Project will haul and deposit both tailings and salt. Tailings and salt will be loaded from their respective storage areas with a front-end loader and hauled by articulating trucks to the tailings and salt management cells. Both the tailings and salt will be spread in thin lifts with a mid-size dozer and compacted with a soil compactor. The cost of this equipment, based on budgetary pricing from a local equipment supplier, is shown in Table 18-9.

Table 18-9: Tailings and Salt Placement Equipment Pre-Production Capital

Item	Unit Cost US$ 000	Number of Units	Total Cost US$ 000’s
Cat 980M Loader	558	2	1,116
Cat D6TXL dozer w/ ripper	418	1	418
Cat 815K compactor	582	1	582
Ledwell LW2000 gal water truck	108	1	108
Magnum MTL4060K Light Plant	10	1	10
Cat 730C2 Articulating Truck	453	2	907
Subtotal			3,140
Contingency			157
Total			3,297

Source: SRK, 2019

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Capital Contingency

Capital contingency was assigned, with some exceptions, on the following assumptions

● 5% for equipment with quotes from supplier;

● 10% for earthwork and liner with contractor quotes; and

● 15% for items estimated from similar projects.

18.2.4.1   Temporary Waste Rock Storage Facility

The Temporary Waste Rock Storage Facility construction will be very similar to the construction of the tailings cell. Costs for similar activities from the tailings cell were applied to the Temporary Waste Rock Storage Facility. Contingencies were also estimated for this construction as they were for the tailings cell.

The storage facility construction will consist of approximately 16,600 m³ of cut to fill earthworks and 69,900 m² of geomembrane installation. The estimated costs for this facility, broken out by major components, are shown in Table 18-10.

Table 18-10: Pre-production Temporary Waste Rock Storage Facility Construction Cost

Item	US$ 000’s
Project Management	150
Access Road/Pipeline Corridor	161
Site Preparation	557
Earthworks	258
Geosynthetics	1,376
Overliner & Drains	567
Subtotal	3,070
Contingency	290
Total	3,360

Source: SRK, 2019

18.2.5  Water Management and Infrastructure

Water Management and Infrastructure include the costs to construct a Water Treatment Plant which will treat mine water, process wastewater, cooling water blowdown and fresh water to supply the facility with its operational water needs as well as produce a solid salt that will be impounded on site. Veolia provided the capital cost for the Water Treatment Plant on a Design-Build-Operate (DBO) basis, inclusive of commissioning, indirect and contingency costs.

Water Management and Infrastructure also include the costs for a series of hydrogeologic investigations and the costs for supplying additional water to the facility from two local landowners and the Tecumseh Board of Public Works. These costs are detailed in Table 18-11.

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Table 18-11: Water Management and Infrastructure Cost

Item	US$ 000’s
Hydrogeology Investigation	6,050
Water Treatment Plant	64,730
Water Supply	2,976
Subtotal	73,756
Contingency	446
Total	74,202

Source: NioCorp, 2019

18.2.6  Site Preparation and Infrastructure Capital Costs

The site preparation and infrastructure capital summarized in Table 18-1 is further broken down in Table 18-12.

Table 18-12: Site Preparation and Infrastructure Costs Summary

Item	US$ 000’s
Site Preparation	18,495
On-site Infrastructure	15,265
Auxiliary buildings	6,146
Surface Mobile Equipment Fleet	664
Total	40,569

Source: Tetra Tech, 2019

Summary of the general items of each category:

Site Preparation

● Site clearing and grubbing

● Topsoil removal and berm construction

● Site grading, pad preparation & access way

● Site roads and parking infrastructure

● Site fencing and access gates

● Construction silt fencing/control, stormwater sediment retention pond

● Architectural landscaping at the main entrance

On-site Infrastructure

● Electrical main substation

● Electrical main power distribution

● Natural gas distribution to site loads

● Surface fuel storage and delivery system

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● Firewater distribution system

● Wastewater network

● Potable water distribution

● Water treatment plant feeder line from the water pipeline

● Stormwater drainage

● Tailings conveyors

● Tailings impoundment facility water recovery system

● Process control, telecommunications, IT, CCTV

● Truck scale at the main gate

● Site lighting on poles

● Waste storage

Auxiliary Buildings

● Gatehouse (leased estimate includes furniture and equipment only)

● Administration and service building (leased estimate includes furniture and equipment only)

● Process analysis laboratory

● Maintenance shop and warehouse building

● Processing plant modular office trailers (leased estimate includes furniture and equipment only)

● Maintenance shop modular office trailers (leased estimate includes furniture and equipment only)

● Mine change house

Surface Mobile Equipment Fleet

● Carry Deck Crane (5T)

● Weld Truck Ford (1T) 4WD

● Ambulance and fire services will be supplied from local municipalities

● Mine Rescue Vehicle

● Mine Rescue Trailer

● Snow Removal Plow blade for Dump Trucks

● Pick-up trucks and service cars will be leased

18.2.6.1   Site Wide Indirects

The site-wide indirect capital summarized in Table 18-1 is further broken down in Table 18-13.

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Table 18-13: Site Wide Indirect Costs Summary

Item	US$ 000’s
Temporary Works	3,192
Temporary Services	4,176
Total	7,368

Source: Tetra Tech, 2019

Each category includes the following:

Temporary Works

● Construction and silt fencing

● Environmental protection

● Main construction parking for processing plants

● Secondary construction parking for processing plants

● Mining area construction parking

● Contractors’ trailer park,

● Laydown area

● Temporary access gate

● Temporary gravel roads

● Temporary electrical power & lighting

● Sanitary installations

● Communications

● Removal of all temporary facilities

Temporary Services

● Site security

● Snow removal

● Grading of parking, roads and lay down areas

● Dust abatement

● Solid Waste management during the pre-production period

● Janitorial services

● Potable water and first aid

● Medical and first aid

18.2.7  Owner’s Costs

Table 18-14 shows the Owner’s cost breakout totalling US$ 33.6 M. No formal contingency is applied to the Owner’s Costs. The Land Acquisition estimate is high, as the key parcel of land hosting the majority of the mineral resource and mineral reserve was acquired by the Company in April 2021 for US$ 6.2 M. The original 2019 budget for Land Acquisition of US$ 11.7M has been

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split into a budget for the remaining land of US$ 5.5 M and an implied contingency for Owner’s Cost of US $6.2M.

Table 18-14: Owner’s Costs Summary

Item	US$ 000’s
Permitting	655
Land Acquisition	5,445
Electrical Utility Company Cash Down Payment	726
Owner’s Team during Project	4,250
Operations Readiness	12,880
Construction Umbrella Insurance	3,253
Other Costs	180
Implied Contingency for Owner’s Costs	$6,230
Total	33,619

Source: Tetra Tech, 2019 and NioCorp, 2022

The general items that make up each category are summarized as follows:

● Permitting

● Land acquisition

● Electrical Utility Company prepayment for the construction of the main substation and establishing electric service to the site.

● Owner’s team during Project execution includes the cost of salary and expenses of the Owner’s personnel dedicated to Project execution

● Operations readiness includes:

○ Specialized assistance for preparing and monitoring the Operations Readiness plan.

○ External assistance for the hiring of personnel.

○ Relocation of personnel.

○ Training program.

○ Preparation of the operations procedures.

○ Preparation of the operating and maintenance schedules, including programming and data input.

○ Procurement activities to fill stores.

○ ERP software, including system configuration, Block development, HMI graphics, Programming, System architecture drawings, Network drawings and training.

○ Other software, including purchase, licenses, and maintenance.

○ Construction umbrella insurance including General Liability, Employer’s and Excess Liability, Worker’s Comp, Builder’s Risk, Contractor’s Pollution Liability, Owner’s Protective Professional and Professional Liability.

○ Other Costs includes the environmental monitoring program.

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18.2.8  Closure and Reclamation

Closure Cost Basis

The closure cost estimate for the Project was developed using the Standardized Reclamation Cost Estimator (SRCE) (available at www.nvbond.org) and a user-defined cost data file (CDF). The inputs to the CDF were obtained from the following sources:

● Equipment costs have been obtained from Gana Trucking, a local Nebraska contractor. These include all-in operator rates, fuel consumption, consumables, and preventive maintenance.

● The operator rates are included in the equipment hire costs. The labour rates are input separately for non-operator rates only.

● Material costs have been obtained from current quotes, where available.

Plant and Mine Facilities

Facilities and equipment associated with the underground mine and processing plant will be reclaimed as follows:

● Plant site buildings will be decontaminated, the buildings will be demolished, and the debris hauled off-site.

● Stockpile underliners will be removed and hauled to the underground for disposal.

● Ponds no longer in use will have sediment and liners removed and hauled to the underground mine for disposal.

● Residual wastes (solid and/or hazardous), will be hauled to appropriate off-site disposal facilities.

● Groundwater wells will be no longer required at the end of operations and will be plugged and abandoned.

● Underground workings (production shaft and ventilation shaft) will be capped to prevent public access post-closure.

● On-site water pipelines will be removed.

● On-site powerlines within the Project boundary will be removed. The utility company will own the substation and would be responsible for its continued use or demolition, once site operations are complete. The natural gas metering station on site that would be owned by a utility would also be managed in this fashion.

● General disturbances will be covered with growth media if necessary and revegetated.

Tailings and Salt Storage Facilities

The Tailing and Salt Management Facilities consist of a series of separate impoundments (cells) for which the exposed tailings and salt surface will be reclaimed as follows:

● Subgrade preparation (i.e., tailings regrading, assumed to be part of the operational costs).

● Synthetic liner installation.

● Above-liner drainage layer construction (i.e., gravel drain layer).

● Above-liner growth media layer placement.

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● Amendment of the growth media layer (scarification and nutrient addition) and placement of sod as an anti-erosion measure.

Reclamation will be carried out concurrently as each of the cells reaches its design capacity.

Post-Closure Monitoring

Monitoring is assumed to continue for 30 years after the end of operations and includes baseline and radiochemical profiles. Monitoring around the tailings and salt management cells will be conducted at three points. Long-term management costs will be limited to the maintenance of a fence around the cells, which is on private property, in perpetuity.

18.2.9  Sustaining Capital Costs

A contingency is included on sustaining capital only on tailings to address construction unknowns.

Mining

The sustaining capital is in the categories of lateral and vertical waste development, mine fixed equipment, and definition drilling/exploration drilling. The sustaining capital captures all costs related to supporting mining activities and applies a percentage of cost towards fixed equipment purchase prices over the life of mine. Table 18-15 presents the sustaining capital for mining.

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Table 18-15: Sustaining Capital for Mining (US$ 000’s)

	Lateral Waste Development	Vertical Waste Development	Fixed Equipment	Definition/ Exploration Drilling	Subtotal	Contingency	Total
Year 4*	10,471	3,962	1,179	6,667	22,279	2,228	24,506
Year 5*	13,454	4,906	1,155	6,667	26,182	2,618	28,800
Year 6*	9,642	1,315	1,204	4,667	16,827	1,683	18,510
Year 7*	5,664	930	1,180	2,667	10,442	1,044	11,486
Year 8*	1,766	623	1,169	2,667	6,224	622	6,846
Year 9*	9,609	930	1,206	2,667	14,412	1,441	15,854
Year 10 to 19*	20,665	1,539	11,788	13,333	47,325	4,733	52,058
Year 20 to 29*	0	0	12,110	10,000	22,110	2,211	24,321
Year 30 to 39*	0	0	11,638	3,000	14,638	1,464	16,102
Total	71,271	14,205	42,629	52,333	180,438	18,044	198,482

Source: NioCorp, 2022 

* All years are from project start with Year 0 being the initial project year.

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Process and Infrastructure

The sustaining capital for process plants, buildings and infrastructure was estimated using a ratio to the direct cost Capital Costs. This ratio has been set to zero from year one to year five and then ramped up to its maximum value between years six to 30 of the Project life. The ratio is decreasing year 29 and 30 and set at zero for years 31 to 38. Table 18-16 presents the sustaining capital for process and infrastructure.

Table 18-16: Sustaining Capital for Process and Infrastructure (US$)

Sustaining Capital	Applicable Capital Costs	Yr 0-5	Yr 6	Yr 7	Yr 8-28	Yr 29	Yr 30	Yr 30-38
Infrastructure	29,504,000	-	49,000	98,000	147,520	98,000	49,000	-
Buildings	53,476,000	-	178,000	357,000	534,760	357,000	178,000	-
Process Plants	292,274,000	-	1,461,000	2,923,000	4,384,110	2,923,000	1,461,000	-
Capital Costs per year		-	1,688,000	3,378,000	5,066,390	3,378,000	1,688,000	-

Source: Tetra Tech, 2019

Tailings, Salt, and Tailings Water Management

A new tailings facility (Plant Site Cell 2) will be needed in 2027 and that facility will have to be expanded (Cell 3) in 2034. When Plant Site Cell 3 in nearly full, a separate facility will be constructed in “Area 7” in 2043. The Area 7 facility is designed to handle all the remaining tails for life of the Project. Based on preliminary designs, the construction cost of these facilities is shown in Table 18-17. A second Salt Management Facility will be constructed in 2041 to replace the initial Salt Management facility which will have reached capacity.

Table 18-17: Tailings and Salt Facility Sustaining Capital Cost

Item	Sustaining Capital US$ 000’s
Tailings Cell 2	14,860
Tailings Cell 3	13,455
Salt Management Cell 2	5,365
Subtotal	62,455
Contingency	5,987
Total	68,441

Source: SRK, 2019

Equipment to load, haul and place tailings will be replaced over the life of the Project. It was assumed that the mobile equipment would be replaced at 40,000 machine hours, with the articulating trucks and the soil compactor replaced at 30,000 machine hours. Light plants are replaced at 10,000 hours. Area 7 Tailings and Salt Management Facility #2 will require hauling

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tailings on Nebraska highway 50. When the Area 7 tailings facility goes into service, a contractor will be used to haul the tailings and salt using over the road trucks. The Project will load the contractor’s trucks and be responsible for spreading and compacting the tailings and the salt. Table 18-18 shows the estimated LOM replacement cost of the tailings placement of mobile equipment.

Table 18-18: Tailings Placement Equipment LOM Replacement Capital

Item	Unit Cost US$ 000’s	Number of Units	Total Cost US$ 000’s
Cat 980M Loader	558	2	1,116
Cat D6TXL Dozer w/ Ripper	418	1	418
Cat 815K Compactor	582	1	582
Ledwell LW2000 Gal Water Truck	108	1	108
Magnum MTL4060K Light Plant	10	1	10
Cat 730C2 Articulating Truck	453	2	907
Subtotal			3,140
Contingency			157
Total			3,297

Source: SRK, 2019

18.2.10   Contingency

The contingency is based on a line-item analysis by category and averages 10% for the capital. Table 18-19 shows the contingency by category for the initial capital.

Table 18-19: Initial Capital Contingency Summary

Capitalized Pre-production Expenses	Percent (%)	Total US$ 000’s
Site Preparation and Infrastructure	10.6%	4,298
Processing Plant	13.9%	50,912
Mine Water Management & Infra	0.6%	446
Mining Infrastructure	10.1%	25,885
Tailings Management	9.2%	1,964
Site Wide Indirects	14.6%	1,075
Processing Indirects	8.8%	8,481
Mining Indirects	10.0%	6,369
Process Commissioning	16.2%	2,157
Mining Commissioning	10.0%	144
Owner’s Costs	0%	0
Total Contingency	10%	101,730

Source: NioCorp, 2022

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18.3 Operating Cost Estimate

18.3.1 Basis of Estimate

The operating cost estimate has an intended accuracy of ± 15%. Operating cost estimates were developed to show monthly and annual costs for production. All unit costs are expressed as US$/tonne processed and are based on Q1 2019 US$. The primary purpose of this report is to address changes to the resource estimate to include contained rare earth elements (REE’s). A subsequent addition of the REEs to the mineral reserve and economics will require additional metallurgical work. Costs have for the most part been retained from the previous 2019 Feasibility Study.

18.3.1.1 Mining Operating Costs

The development and operation of the underground mine will be carried out by mining contractors. Mine operating costs were developed from first principles. Input from mining contractors, blasting suppliers and equipment vendors, was considered for the key parameters and contractor unit rates. The rates developed from first principles were adjusted based on benchmarking and the experience and judgment of the mine design team in collaboration with mining contractors. Vendor quotations for high use materials were obtained that included freight. Productivity information was developed based on first principles for the mining tasks. The required labour was developed based on mine plan requirements for equipment and material movements. The mine plan quantities also dictated the material quantities required and unit pricing based on vendor quotes were applied to determine material costs. Maintenance supplies and labour for fixed equipment, as well as management and technical personnel, were included in the mining cost. A unit cost for backfill was developed and included in the mine operating cost. The costs vary by year based on production requirements. Haulage distance was taken into consideration on haulage costs. A contingency was applied to mine operating costs.

18.3.1.2 Process Plants Operating Costs

The operation of the surface plants will be carried out by the mine owner. The annual process operating costs were determined by estimating the required quantities of workforce, natural gas, electrical power, reagents, and consumables required for one year and applying current unit cost rates to develop an annual operating cost for each area.

18.3.1.3 Tailings and Tailings Water Management Operating Costs

Basis of the tailings operating cost includes the following cost items supplied by the client:

● The hourly wage rate for the equipment operator and truck driver: US$ 23.29/hour

● Labor burden: 35.91%

● Dyed diesel fuel: US$ 2.09/gallon

The equipment operating cost was developed from Infomine Costmine (2016)

The tailings and salt operating cost for haulage from the plant site to Area 7 TSF and Salt Cell #2 are based on quotes provided by Gana Construction, a construction contracting firm based in Southeast Nebraska.

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18.3.1.4 Site G&A Operating Costs

The annual Site G&A operating costs were determined by estimating the required quantities of workforce and using allowances for fixed costs such as consumables, supplies, etc., based on SRK’s experience with analogous projects and discussions with NioCorp Project team members.

18.3.1.5 Owner’s Costs Capital Costs

The Owner’s Capital Costs were estimated mainly from allowances based on Tetra Tech’s experience with analogous projects and discussions with NioCorp Project team members.

18.3.1.6 Water Supply Operating Costs

The operating cost is based on discussions between NioCorp and area landowners as well as a cost estimate provided by the Tecumseh Board of Public Works.

18.3.1.7 Closure and Reclamation

The closure cost estimate for the Project was developed using the Standardized Reclamation Cost Estimator (SRCE) (available at www.nvbond.org) and a user-defined cost data file (CDF). The inputs to the CDF were obtained from the following sources:

Equipment costs have been obtained from Gana Trucking. These include all-in operator rates, fuel consumption, consumables, and preventive maintenance.

Operator rates are included in the equipment hire costs. The labour rates are input separately for non-operator rates, only. Material costs have been obtained from quotes, where available.

18.3.2 Operating Cost Summary

Table 18-20 summarizes the operating costs estimate by area, which equals US$ 195.94/t ore. These unit rates are stated on a LOM basis where the costs are estimated from the beginning of construction to the end of mine life. LOM operating costs include the pre-production and first/last years of production.

Table 18-20: LOM Operating Cost Unit Rate Summary 

Description	LOM US$/t ore
Mining Cost	42.38
Process Cost	106.70
Water Management Cost	16.62
Tailings Management Cost	2.01
Other Infrastructure	5.47
Site G&A Cost	8.20
Other Expenses	6.22
Subtotal	187.59
Royalties/Annual Bond Premium	8.35
Total LOM Operating Costs	195.94

Source: NioCorp, 2022

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18.3.2.1 Mining Operating Costs

Mine operating costs for the LOM, (pre-production and steady state) are US$ 42.38 /t of ore produced. Table 18-21 summarizes the breakdown by mine production activities and shows the general services costs and labour that are allocated over all the tonnage produced. These costs include the cost to drill, blast, install ground support, shotcrete, grouting, load and haul, crush and handle materials to the surface, ventilation, pumping, general maintenance, technical services, backfill, and mine management. The operating cost varies by year, by mine location and production. The annual operating cost varies by year but averages approximately US$ 44 million per year over the LOM. The mining operating cost is based on a Q1 2019 cost basis.

Table 18-21: Steady State Mining Operating Unit Cost (after pre-production) 

Description	Steady State (US$ 000’s)	Cost per Tonne Ore
Production Drill, Blast, Backfill	433,337	11.82
Development	167,297	4.56
Trucking and Hauling	287,692	7.85
Power	180,634	4.93
Underground Services and G&A	340,180	9.28
Subtotal Operating Cost	1,409,140	38.44
Contingency	144,185	3.93
Total Operating Cost	1,553,325	42.38

Source: NioCorp, 2022

18.3.2.2 Process Plant Operating Costs

The annual LOM operating costs for the Process and Infrastructure portion of the plant is estimated at US$ 106.70/t of mineralized material processed (Table 18-22). This estimate includes four primary areas of the surface plant: Mineral Processing, Hydrometallurgical Plant, Pyrometallurgical Plant, and Infrastructure. The estimate for each of these four areas was developed by determining the required quantities of workforce, energy (natural gas, electrical power, and fuel) reagents, consumables and other general costs required for one year of operation and then applying current unit cost and feed rates to develop an annual operating cost for each area. These costs were then used to calculate other valuable metrics, such as dollars-per-ton-milled. Table 18-22 summarizes the costs for each area.

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Table 18-22: ROM Processing Operating Cost Unit Rate Breakdown

Cost Items	Annual Cost (2,764 t/d) (US$/y)	Annual Cost Per Tonne Milled (2,764 t/d) (US$/y)
Mineral Processing	$4,046,169	4.38
Workforce	$1,356,869	1.47
Energy	$1,281,071	1.39
Reagents	0	0.00
Consumables	$1,324,176	1.43
Other Processing	$84,053	0.09
Hydromet	$78,327,091	84.73
Workforce	$4,867,874	5.27
Energy	$36,272,182	39.24
Reagents	$31,670,766	34.26
Consumables	$5,262,275	5.69
Other Processing	$253,994	0.27
Pyromet	$16,258,784	17.59
Workforce	$1,701,265	0.99
Energy	$1,684,128	1.82
Reagents	$11,827,660	12.80
Consumables	$918,735	0.99
Other Processing	$126,997	0.14
Infrastructure	$1,654,671	1.79
Water Management	$15,331,749	16.38
Product Packaging	$801,716	0.87
Other	$2,597,556	2.81
Total Process Cost	$119,017,735	128.55

Source: NioCorp, 2022

The operating costs for this Project are based on processing 2,764 t of ore per day to produce an average of 7,220 t/y of ferroniobium. These operating costs are based on Q1-2019 pricing data.

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18.3.2.3 Tailings, Salt and Tailings Water Management Operating Costs

Tailings Operating Costs

During tailings and salt disposal operations, tailings and salt will be loaded from a storage building near the backfill plant and water treatment plant respectively and hauled to the tailings storage or salt impoundment facilities in 30-ton articulating trucks. In order to maximize the density, both the tailings and salt will be dozed into thin lifts and compacted using a soil compactor.

Tailings storage at the backfill plant is limited, so costs for tailings placement were estimated by assuming that the work will be completed on a 10-hour shift, seven days per week. Two crews of operators, consisting of four persons per crew, will alternate on a four days on, four days off schedule. Each crew will consist of two equipment operators and two truck drivers. Half of the year, it was assumed that there would be a full-time water truck driver. One equipment operator will run the front-end loader to load the trucks. The second operator will alternate between the dozer and soil compactor at the tailings cell. Alignment of the work schedules for tailings personnel with the balance of mine and plant operational schedules will be evaluated at the detailed design stage. This same crew would also transport the salt from the water treatment storage building to the salt impoundment facilities.

Truck productivity was calculated using Caterpillar Inc.’s Fleet Production and Cast Analysis (FPC) software to determine the haul times required the trucks to place the tailings. Parameters used in determining the haulage requirements to Plant Site Cell 1 are shown in Table 18-23. This analysis shows that two trucks can handle the amount of tailings produced with sufficient capacity in case of equipment downtime.

Table 18-23: Tailings Haulage Calculations

Description	Value	Unit of Measure
Average tailings stacking	1100	t/d
Tailings bulk density	1.6	t/m3
Haul trucks	Cat 730C2
Haul truck capacity	28.00	t
Haul truck capacity	17.50	m3
Estimated load	13.50	m3
Estimated load	21.60	tonne
Loads/day	51	loads
Trucks operating	2.00	each
Hours per shift	10.00	hr
Loads per truck-shift-hr	2.55	loads/hr
Plant Site Cell 1 haul - one way	1200	m
Potential cycle time (FPC)	9.4	min
Utilization	80%
Potential 2 truck production	221	t/h

Source: SRK, 2019

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Plant Site Cells 2 and 3 have longer haul distance, but the equipment fleet will be able to handle the additional haulage cycle time.

Equipment operating costs were obtained using Costmine, modified for fuel pricing. Operating cost includes the costs associated with major component rebuild. Adjusted equipment hourly costs are shown in Table 18-24.

Table 18-24: Tailings and Salt Mobile Equipment Hourly Operating Cost

Equipment Type	Model	Fuel US$/hr	Lube US$/hr	Tires (US$/hr)	Overhaul (US$/hr)	Maint (US$/hr)	Wear  Items (US$/hr)	Total  Cost  US$/hr
Loader	Cat 980M Loader	19.98	6.64	17.56	7.75	14.40	0.66	66.99
Dozer	Cat D6TXL dozer w/ ripper	11.41	3.58	-	6.15	9.21	9.00	39.35
Compactor	Cat 815K compactor	13.15	4.47	0.34	19.68	16.11	-	53.75
30 Ton Articulating Truck	Cat 730C2 Articulating Truck	10.40	5.32	3.39	4.50	8.34	-	31.94
Skid Steer		4.56	0.78	0.32	0.88	1.66	0.14	8.33
Water Truck 2000 Gallon	Ledwell LW2000 gal water truck	10.68	3.48	2.55	1.10	2.65	-	20.46
Light Plant	Magnum MTL4060K Light Plant	0.70	0.19	0.02	0.27	0.50	-	1.69

Source: Infomine, 2016

Costs were calculated on a period basis. It was assumed that all tailings haulage operators would be paid based on working a full 10-hour shift every day of the period. Equipment utilization factors were assumed for each equipment type. Table 18-25 shows the utilization factors for equipment usage.

Table 18-25: Tailings and Salt Mobile Equipment Utilization

Equipment Type	Utilization
Loader	80%
Dozer	45%
Soil Compactor	40%
Water truck	60%
Articulating trucks	85%
Light plant	15%

Source: SRK, 2019

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In Year 19 when tailings and salt disposal move to the Area 7 TSF, highway trucks will be required when hauling tailings and salt to Area 7 and Salt Cell #2 (instead of the articulating trucks used for disposal in the Plant Site TSFs). A quote was received from a contractor to tailings haulage only for US$ 2.55/t tailings. NioCorp will be responsible for loading the contractor’s trucks and for spreading and compacting the tailings.

Estimated costs for tailings loading, haulage, and placement are shown in Table 18-26.

Table 18-26: Cost for Tailings and Salt Placement

Haulage Type	Cost, US$ /t
Owner hauled tailings and salt	3.00
Contractor hauled tailings and salt	4.38

Source: SRK, 2019

18.3.2.4 Site G&A Operating Costs

Estimates of LOM Site General and Administrative (Site G&A) operating costs for the Project were calculated from first principles. The results are presented in Table 18-27, which shows a LOM unit rate of US$ 8.20/t ore on an annual basis, Site G&A costs average US$ 8.4 million, as shown in Table 18-28.

Table 18-27: LOM Site G&A Operating Costs 

Description	US$ 000’s	US$ /t
Site Management	$42,991	$1.17
Processing Overhead	$73,660	$2.01
Technical Services	$62,414	$1.70
Health & Safety	$19,600	$0.53
Human Resources	$14,896	$0.41
Supply Chain Management	$21,727	$0.59
Information Services	$13,946	$0.38
Finance	$15,471	$0.42
Community and Social Responsibility	$3,203	$0.09
Environmental and Permitting	$22,615	$0.62
Site Services	$9,878	$0.27
Total	$300,400	$8.20

Source: NioCorp, 2022

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Table 18-28: Site G&A Annual Operating Costs

Description	US$ 000’s
Site Management
Salaries and Wages	268
Office Supplies	60
Rents/Premiums/Travel	880
Subtotal Site Management	1,208
Processing Overhead
Salaries and Wages	2,070
Materials, Supplies, Consumables, Training	-
Subtotal Processing Overhead	2,070
Technical Services
Salaries and Wages	1,754
Materials, Supplies, Consumables, Training	-
Subtotal Technical Services	1,754
Health and Safety
Salaries and Wages	465
Materials, Supplies, Consumables, Training	86
Subtotal Health & Safety	551
Human Resources
Salaries and Wages	259
Recruitment/Relocation	100
L&D Training Programs	60
Subtotal Human Resources	419
Supply Chain Management
Salaries and Wages	561
Materials, Supplies, Consumables	10
Contract Services/Head Office Support	40
Subtotal SCM	611
Information Services
Salaries and Wages	235
IT Equipment/Licenses	157
Subtotal Information Services	392
Finance
Salaries and Wages	292
Contract Services/Head Office Support	143
Subtotal Finance	435
Community and Social Responsibility
Salaries and Wages	-
Marketing Supplies	15
Community Funding	75
Subtotal CSR	90
Environmental and Permitting
Salaries and Wages	536

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Operating Costs	50
Contract Services	50
Subtotal E&P	636
Site Services
Salaries and Wages	258
Site Facilities Maintenance	20
Subtotal Site Services	278
Grand Total	8,442
Total Salaries and Wages	6,696
Total Other Fixed Costs	1,747

Source: NioCorp, 2019

The Site G&A labour costs, as shown in Table 18-28 of US$ 6.7 million per year are derived by the 80.5 staff headcount, as shown in Table 18-29. The overall organizational chart for the entire operation during LOM is shown in Figure 18-1. The Site G&A numbers do not include any mining or any direct processing operations or maintenance staff. However, both Processing Overhead and Technical Services that are part of the Site G&A area are grouped within mining and processing areas in the organizational chart for simplicity.

Table 18-29: Average Annual G&A Headcount during Operations 

Description	Headcount
Site Management	2.5
Processing Overhead	21
Technical Services	23
Health & Safety	6
Human Resources	3
Supply Chain Management	9
Information Services	2
Finance	3
Environmental and Permitting	6
Site Services	5
Total G&A Personal	80.5

Source: NioCorp, 2019

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Source: NioCorp, 2019 

Figure 18-1: Elk Creek Project LOM organizational chart

In terms of fixed costs, allowances were developed for each category based on experience from similar US mining projects (see Table 18-30). The average annual Site G&A fixed cost (non-labour) estimate is US$ 1.7 million per year. The Project is somewhat unique in that although it is located in a rural area in southeastern Nebraska, it is only one hour drive from the city of Omaha, an eight-minute drive to a 40-bed county hospital and has a high-capacity fiber optic trunk passing the property on the east side adjacent to state highway 50.

Table 18-30: Average Annual G&A Fixed Costs During Operations

Description	Whole US$
Site Management
Materials, Supplies, Consumables
Office Supplies	30,000
Postage, Courier and Light Freight	10,000
Copying and Printing	20,000
Subtotal Materials, Supplies, Consumables	60,000
Insurance
Property, Business Interruption, Bldgs, Equip, Liability	750,000
Insurance	750,000
Government Surcharges/Fees(1)
County	10,000
Other	5,000
Subtotal Property Taxes Government Surcharges/Fees	15,000
Travel/Professional
Conferences and Meetings	25,000
Outside Accounts (Small vehicle repair, dining, catering)	20,000
Dues and Subscriptions	20,000
Business Travel & Accommodation	50,000

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Description	Whole US$
Subtotal Travel/Professional	115,000
Total Site Management Fixed Costs	940,000
Total Processing Overhead Fixed Costs	-
Total Technical Services Fixed Costs	-
Health and Safety
Emergency Supplies
First Aid Stations	10,000
Fire Extinguishers	50,000
Basket/Stretcher	1,250
Subtotal Emergency Supplies	61,250
Training
Staff Training Contractor	-
Training Supplies/ Classes	25,000
Subtotal Training	25,000
Total H&S Fixed Costs	86,250
Human Resources
Recruitment/Relocation
Recruitment Allowance	15,000
Recruitment Fees	15,000
Relocation and Assignment Cost	70,000
Subtotal Recruitment/Relocation	100,000
L&D Training Programs
Staff Training Contractor	25,000
Training Supplies/ Classes	25,000
NG Corporate Head Office Support	10,000
Subtotal Human Resources	60,000
Total HR Fixed Costs	160,000
Supply Chain Management
Subtotal Materials, Supplies, Consumables	10,000
Contract Services - Trucking	10,000
Corporate Head Office Support	30,000
Total SCM Fixed Costs	50,000
Information Services
IT Equipment/Licenses
IT Equipment	7,355
IT Software (On-site IT Support FTE)	90,000
Private Mobile Radio (PMR)	10,000
Carrier Services	50,000
Total IS Fixed Costs	157,355
Finance
Contract Services/Head Office Support
Contract Services - Legal	50,000
Contract Services - Consultants (tax, acctg, mgmt)	50,000
Contract Services - External Audits	13,000
NG Corporate Head Office Support	30,000

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Description	Whole US$
Total Finance Fixed Costs	143,000
Community and Social Responsibility
Subtotal Materials, Supplies, Consumables	15,000
Community Funding
Charitable Contributions	50,000
Sponsorships	25,000
Subtotal Community Funding	75,000
Total CSR Fixed Costs	90,000
Environmental and Permitting
Subtotal Materials, Supplies, Consumables	50,000
Contract Services - Annual ESR Studies	50,000
Total E&P Fixed Costs	100,000
Site Services
Subtotal Materials, Supplies, Consumables	20,000
Total Site Services Fixed Costs	20,000
Grand Total	1,746,605

Source: NioCorp, 2019 

(1) Does not include annual county property taxes which are included in operating expenses in the technical economic model as a direct cash cost.

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19. ECONOMIC ANALYSIS

19.1 Methodology Used

The Project has been evaluated using discounted cash flow analysis. Cash inflows consist of annual revenue projections. Cash outflows consist of initial capital expenditures, sustaining capital costs, operating costs, taxes, royalties, and commitments to other stakeholders. These are subtracted from revenues to arrive at the annual cash flow projections. Cash flows are taken to occur at the end of each period. To reflect the time value of money, annual cash flow projections are discounted back to the Project valuation date using the yearly discount rate. The discount rate appropriate to a specific project can depend on many factors, including the type of product, the cost of capital to the Project, and the level of Project risks (i.e., market risk, environmental risk, technical risk and political risk) in comparison to the expected return from the equity and money markets. The base case discount rate for this Feasibility Study is 8%. The discounted present values of the cash flows are summed to arrive at the Project’s NPV (Net Present Value). In addition to the NPV, the IRR (Internal Rate of Return) and the payback period are also calculated. The IRR is defined as the discount rate that results in an NPV equal to zero. The payback period is calculated as the time required to achieve positive cumulative cash flow for the Project from the start of production.

19.2 Financial Model Parameters and Assumptions

The indicative economic results summarized in this section are based upon work performed by NioCorp in 2019 and 2022. They have been prepared on both a periodic monthly/quarterly format and an annual format. The metrics reported in this section are based on the annual cash flow model results. The metrics are on both a pre-tax and after-tax basis; a 100% equity basis with no Project financing inputs; and are in Q1 2019 U.S. constant dollars.

Key criteria used in the analysis are discussed in detail throughout this section. Principal Project assumptions used are shown summarized in Table 19-1.

Table 19-1: General Assumptions

Description	Value
Pre-Production Period	4 years
Process Plant Life	38 years
Mine Operating Days per Year	365
Mill Operating Days per Year	365
Discount Rate	EOP @ 8%
Commercial Production Year	2025

Source: NioCorp, 2022

All costs incurred prior to April 2022 are considered sunk with respect to this analysis.

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The selected Project discount rate is 8% as directed by NioCorp, and the valuation uses standard end of-period discounting. A sensitivity analysis of the discount rate is discussed later in this section.

Foreign exchange impacts were deemed negligible as most, if not all costs and revenues are denominated in US dollars.

The major criteria adopted to define when the Project enters into Commercial Production include the following: (1) all major capital expenditures to bring the mine to nameplate capacity have been completed; (2) the process plant, and other facilities have been transferred to the control of the Operations team from the Commissioning team; (3) the plant has reached at least 80% of initial design capacity following an adequate ramp-up period; (4) product recoveries are at or near expected levels; (5) the mine has the ability to sustain ongoing production of ore at the required CoG (Cut-off Grade); and (6) costs are under control or within expectations.

Mineral Resource, Mineral Reserve and Mine Life

The Mineral Resource discussed in Section 11 was converted to the Mineral Reserve outlined in Section 12. The estimated Mineral Reserve will support a 38-year production life, using the mine plan as provided in Section 13.

Metallurgical Recoveries

The basis for the process recoveries is included in Section 10, and the process design is outlined in Section 14.

Product Prices

The product price basis is discussed in Section 16.

Capital and Operating Costs

The capital and operating cost estimates are detailed in Section 18.

Closure Costs and Salvage Value

Reclamation costs were included with the capital cost estimate.

Financing

The economic analysis assumes 100% equity financing and is reported on a 100% project ownership basis.

Inflation

The economic analysis assumes constant prices with no inflationary adjustments.

19.2.1 Physicals

Mining

Table 19-2 is a summary of the estimated mine production over the 36-year LOM. Ore mined refers to Probable Mineral Reserves.

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Table 19-2: Mining Physicals

Description	Value
Ore Mined (kt)	36,655
Ore Mining Rate (t/d)	2,764
Niobium Grade	0.81%
Scandium Grade (ppm)	70.2
TiO2 Grade	2.92%
Contained Nb2O5 (kt)	297
Contained Sc (t)	2,573
Contained TiO2 (kt)	1,071

Source: NioCorp, 2022

Processing

A summary of the estimated process plant production for the Project is contained in Table 19-3 for a 38-year operating life at an average capacity of 1.01 Mt/y. Table 19-4 shows more detail of process recovery rates for each product in the three plants. Ore processed refers to Probable Mineral Reserves.

Table 19-3: Processing Physicals

Description	Value
Total Ore Processed (kt)	36,655
Processing Rate (kt/y)	1,009
Average Recovery, Nb	82.4%
Average Recovery Sc	93.1%
Average Recovery TiO2	40.3%
Recovered Nb2O5 (kt)	245
Recovered Sc (t)	2,395
Recovered TiO2 (kt)	432

Source: NioCorp, 2022

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Table 19-4: Processing Recovery Summary

Description	Nb	Ti	Sc
Mineral Processing Plant	100.0%	100.0%	100.0%
Hydrometallurgical Plant	85.8%	40.31%	93.1%
Pyrometallurgical Plant	96.0%
Overall Recovery	82.4%	40.3%	93.1%

Source: Tetra Tech Memo, 5/15/2017

19.2.2 Revenue

Based on data discussed in Section 16, Table 19-5 and Table 19-6 show benchmark product pricing assumptions used in the economic analysis. The following criteria apply to the calculation of revenue:

● Niobium measured in the resource and reserve as Nb₂O₅ but is produced as commercial ferroniobium, which is a mixture typically containing 65% Nb and 35% Fe. Ferroniobium pricing is based solely on its Nb content.

● TiO₂ is measured as TiO₂ in the resource and reserve and is produced as that same compound.

● Scandium is measured as Sc in the resource and reserve and is produced and sold as the compound Sc₂O₃.

● REEs will be recovered in processing but have not been factored into projected revenue.

Table 19-5: Pricing Assumptions 

Description	Tonnes Saleable Product	LOM Benchmark Price US$/kg product
Payable Nb	171,140	46.55
Payable Sc2O3	3,676	$3,675
Payable TiO2	431,793	0.99

Source: NioCorp, 2022

Table 19-6: Scandium Trioxide Pricing Assumptions 

Year	US$/kg
2022	3,600
2023	3,700
2024	3,800
2025	3,900
2026	4,000
2027	3,500
2028	3,000
2029	3,100
2030	3,200

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2031	3,400
2032	3,600
2033+	3,750

Source: NioCorp, 2022, OnG 2019

The following is a breakdown of netback pricing assumptions for each product:

Niobium

● Ferroniobium (65% Nb) product (FeNb product) with constant LOM Benchmark/Provisional Price of US$ 47/kg Nb.

● All settlement Nb prices have a 3.75% discount to the netback price of benchmark price minus Buyers Logistics Costs (BLC) except with customers buying on spot pricing.

● It is assumed that all FeNb product purchases have a 10 Net Days Outstanding (NDO) A/R term. At the time of this report, the Project had two committed offtake customers signed up for 10-year terms with all remaining annual FeNb production sold on a spot basis:

○ Buyer #1 – US-based metals trader with mill operations located in the southern half of the US:

■ 10-year commitment to purchase 25% of annual offtake production to a maximum of 1,875 t/y.

○ Buyer #2 – European-based manufacturer with global mill operations:

■ 10-year commitment to purchase 50% of annual offtake production to a maximum of 3,750 t/y.

○ Spot Buyer - It is assumed that all annual FeNb production not sold under an offtake agreement is sold at spot (or benchmark) pricing of constant US$ 47/kg Nb on an ex-mine gate basis with a 10-day NDO A/R term.

Based on these pricing assumptions, the average realized LOM Nb price is US$ 46.55/kg.

Titanium Dioxide

● No offtake agreements have been negotiated at the time of writing this report.

● It is assumed that all annual TiO₂ production is sold at spot (or benchmark) pricing of constant US$ 0.99/kg on an ex-mine gate basis with a 10-day NDO A/R term.

Scandium Trioxide

● Scandium Trioxide (Sc₂O₃) product with an average realized LOM price of US$ 3,675/kg.

● It is assumed that all Sc₂O₃ product purchases have a 10-day NDO A/R term.

At the time of this report, the Project has one committed offtake customer signed up for a 10-year term with all remaining Scandium Trioxide sold on a spot basis.

● 10-year commitment to purchase a minimum of 12 tonnes per year.

● Under the agreement, the buyer has exclusive rights to the aerospace and sporting goods sectors.

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19.2.3 Operating Costs

Operating cost metrics in the technical, economic model are reported on a LOM basis meaning that all of these unit rates are stated on a LOM basis where the costs are estimated from the beginning of construction to the end of mine life. LOM operating costs include the pre-production and first/last years of production.

The total LOM operating cost unit rate of US$ 195.94/t processed is summarized in Table 19-7.

Table 19-7: Operating Cost Summary

Description	US$/t ore
Mining	42.38
Processing	106.70
G&A	8.20
Water Management	16.62
Tailings Management	2.01
Other Infrastructure	5.47
Other Expenses	6.22
Subtotal Operating Costs	187.59
Royalties/Bond Premium	8.35
Total All-in Operating Costs	195.94

Source: NioCorp, 2022

19.2.4 Capital Costs

Total LOM capital costs totalling US$ 1,562 million, not including US$ 44 million of final closure/reclamation costs are summarized in Table 19-8. Total initial capital costs of US$ 1,141 million, including a 10% contingency, are part of this total.

Table 19-8: Capital Cost Summary (US$ 000’s)

Description	Initial	Sustaining	Total
Capitalized Pre-production Costs	77	-	77
Site Preparation and Infrastructure	41	15	56
Processing Plant	367	96	464
Water Management & Treatment	74	24	97
Mining Infrastructure	257	198	455
Tailings Management	21	79	100
Site Wide Indirects	7	-	7
Processing Indirects	96	-	96
Mining Indirects	41	-	41
Commissioning	15		15
Owner’s Costs	34	-	34
Mine Water Management Indirects	9	-	9
Contingency	102	9	111
Total Capital Costs	US$ 1,141	US$ 422	US$ 1,562

Source: NioCorp 2022

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Further detail of the initial capital estimate is presented in Table 19-9 which shows it is partially offset by a Gross Pre-production Revenue Credit of US$ 257 million generated from product sales made in the preproduction period, which is before the start of commercial production. 

Table 19-9: Initial Capital Cost Summary

Description	US$ Million	% of Total
Capitalized Pre-Production Costs	77	7%
Process Commissioning	15	1%
Subtotal Pre-production Costs	92	9%
Site Preparation and Infrastructure	41	4%
Processing Plant	367	32%
Mine Water Management	74	6%
Mining Infrastructure	257	22%
Tailings Management	21	2%
Subtotal Direct Costs	US$ 852	75%
Site Wide	7	1%
Processing	96	8%
Mining	40	3%
Owner’s Costs	34	3%
Mine Water Management	9	1%
Subtotal Indirect Costs	US$ 186	16%
Project Total Before Contingency	US$ 1,039	91%
Contingency	102	9%
Project Total Before PP Revenue Credit	US$ 1,141	100%
Gross Pre-production Revenue Credit*	(257)
Project Total*	US$ 884

Source: NioCorp, 2022 

*Revenue from sales occurring during commissioning and ramp-up phases

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An estimate of US$ 30 million of working capital is estimated for the last year before and the first year of commercial production. The assumptions used for this estimate are as follows:

● Accounts Receivable (A/R) – product /offtake agreement specific (see revenue section)

● Accounts Payable (A/P) – 30-day delay

● Consumable Inventory – 60-day supply

Annual adjustments to working capital levels are made in the technical economic model with all working capital recaptured by the end of LOM resulting in a LOM net free cash flow (FCF) impact of 0.00 US$.

19.3 Cashflow Forecasts and Annual Production Forecasts

Cashflow forecasts are summarized on a LOM basis in this section.

The technical, economic model metrics are prepared on an annual pre-tax and after-tax basis, the results of which are summarized in Table 19-10, Table 19-11 and Table 19-12.

Table 19-10: Indicative Economic Results (US$ 000’s unless otherwise indicated)

Description	Value
Realized Market Prices
Nb	US$ 46.55
TiO2	US$ 0.99
Sc2O3	US$ 3,674
Payable Metal
Nb (t)	171,140
TiO2 (t)	431,793
Sc2O3 (t)	3,676
Total Gross Revenue	US$ 21,899,726
Operating Costs
Mining Cost	(1,553,325)
Process Cost	(3,911,116)
Site G&A Cost	(300,400)
Concentrate Freight Cost	(10,472)
Other Infrastructure Costs	(200,407)
Water Management Cost	(609,195)
Tailings Management Cost	(73,822)
Property Tax	(217,540)
Royalties	(300,503)
Annual Bond Premium	(5,500)
Total Operating Costs	(US$ 7,182,280)
Operating Margin (EBITDA)	US$ 14,717,445
Effective Tax Rate	16.42%
Total Taxes	(US$ 2,246,186)
Working Capital	0
Operating Cash Flow	US$ 12,471,258
Capital
Initial Capital	(1,140,544)

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Sustaining Capital	(412,405)
Reclamation/Salvage Capital	(44,267)
Total Capital	(US$ 1,606,601)
Metrics
Pre-tax Free Cash Flow	US$ 13,121,263
Pre-tax NPV @ 8%	US$ $2,819,000
Pre-tax IRR	29.2%
After-tax Free Cash Flow	US$ $10,875,077
After-tax NPV @ 8%	US$ 2,350,000
After-tax IRR	27.6%
After-tax Undiscounted PB from Start of CP (Years)	2.69

Source: NioCorp, 2022

Table 19-11: Elk Creek Economic Results 2022

Description	(US $millions)
Pre-Tax Net Present Value (“NPV”) (8% discount)	$2,819
Pre-Tax Internal Rate of Return (“IRR”)	29.2%
After-Tax NPV	$2,350
After-Tax IRR	27.6%
After-Tax Payback Period from Production Onset (years)	2.69
Total Upfront Capex	$1,141
Mine Life (Years)	38
LoM Gross Revenue ($M)	$21,900
Niobium	$7,968
Scandium	$13,504
Titanium	$427
Averaged Annual EBITDA over LoM	$397.5
Averaged EBITDA Margin (EBITDA as % of total revenue)	69%
Averaged Annual Operating Cash Flow over LoM	$337
Average Annual Operating Cost, LoM (“OPEX”) (US$/t)	($195.9)
Averaged Annual EBITDA over RoM	$403
Averaged EBITDA Margin (EBITDA as % of total revenue)	68%
Averaged Annual Operating Cash Flow over RoM	$340
Effective Tax Rate	16.4%

Source: NioCorp, 2022

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Table 19-12: Indicative Economic Results

Operating Year		1	2	3	4	5	6	7	8	9	10	20	30
Production
Niobium	t-Nb	4,960	4,742	4,949	4,903	4,949	4,716	4,715	4,733	4,799	4,672	4,772	4,773
Scandium	kg-Sc2O3	116	114	113	109	112	109	105	102	101	101	102	107
Titanium	t-TiO2	13,063	12,120	12,747	12,605	12,606	12,114	11,846	12,167	11,926	11,544	12,365	12,527
Realized Pricing
Niobium	$/kg	$45.46	$45.46	$45.46	$45.46	$45.46	$45.46	$45.46	$45.46	$45.46	$45.84	$47.00	$47.00
Scandium	$/kg	$3,986	$3,487	$2,989	$3,088	$3,188	$3,387	$3,586	$3,735	$3,734	$3,750	$3,750	$3,750
Titanium	$/kg	$0.99	$0.99	$0.99	$0.99	$0.99	$0.99	$0.99	$0.99	$0.99	$0.99	$0.99	$0.99
Gross Revenues ($M)		$701	$626	$575	$573	$596	$594	$602	$608	$608	$606	$617	$637
Total Opex ($M)		($205)	($200)	($201)	($207)	($210)	($196)	($201)	($202)	($210)	($211)	($207)	($200)
EBITDA ($M)		$496	$426	$374	$366	$386	$398	$401	$406	$398	$395	$411	$436
EBITDA Margin		71%	68%	65%	64%	65%	67%	67%	67%	65%	65%	67%	69%
Operating CF ($M)		$496	$426	$353	$328	$341	$346	$342	$345	$339	$339	$339	$356
EBT ($M)		$227	$202	$181	$188	$222	$259	$284	$295	$287	$283	$293	$326
Net Income ($M)		$227	$202	$161	$150	$176	$207	$225	$234	$228	$226	$221	$245
Income Margin		32%	32%	28%	26%	30%	35%	37%	39%	38%	37%	36%	39%

Source: NioCorp, 2022

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Based on current assumptions and design listed in this report, the Project returns a pre-tax NPV 8% of US$ 2,819 million and an IRR of 29.2% along with an after-tax NPV 8% of US$ 2,350 million and IRR of 27.6%.

Figure 19-1 and Figure 19-2 present annual pre-tax and after-tax free cash flow versus payable Nb production and shows that the Project is expected to generate a stable positive cash flow through the LOM.

Source: NioCorp, 2022

Figure 19-1: Annual Project Metrics Summary (Pre-Tax)

Source: NioCorp, 2022

Figure 19-2: Annual Project Metrics Summary (After-Tax)

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19.4 Sensitivity Analysis

The cash flow model was tested for sensitivity to variances in milled tonnes, head grades (Nb, Sc, and Ti), process recoveries (Nb, Sc, Ti), metal prices, initial/sustaining capital expenditure and operating costs (mining, processing, water management, tailings management, site G&A and royalties).

Figure 19-3 and Figure 19-4 illustrate the results of pre/post tax basis with respect to four of the operational parameters and product prices along with recovery and head grades. The anticipated project cash flow is sensitive to the price of scandium and niobium compared to capital and operating costs, which were both quite similar.

Source: NioCorp, 2022

Figure 19-3: Pre-Tax NPV 8% Sensitivity Graph

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Source: NioCorp, 2022

Figure 19-4: After-Tax NPV 8% Sensitivity Graph

Sensitivity graphs in Figure 19-5 and Figure 19-6 demonstrate the Project IRR is sensitive to changes in Sc₂O₃ and Nb prices on both a pre-tax and after-tax basis, but capital costs clearly have a greater effect than operating costs.

Source: NioCorp, 2022

Figure 19-5: Pre-Tax IRR Sensitivity Graph

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Source: NioCorp, 2022

Figure 19-6: After-Tax IRR Sensitivity Graph

Figure 19-7 and Figure 19-8 illustrate the results of pre/post tax basis with respect to head grades and process recoveries of the three products. Not surprisingly, the impact of a head grade reduction is exactly equivalent to a process recovery reduction for each of the products.

Source: NioCorp, 2022

Figure 19-7: Pre-Tax NPV 8% Sensitivity Graph 

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Source: NioCorp, 2022

Figure 19-8: After-Tax NPV 8% Sensitivity Graph

Sensitivity graphs in Figure 19-9 and Figure 19-10 demonstrate the Project IRR is sensitive to changes in Sc₂O₃ and Nb head grade and recovery on both a pre-tax and after-tax basis, but with limited to no impact from TiO₂.

Source: NioCorp, 2022

Figure 19-9: Pre-Tax IRR Sensitivity Graph

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Source: NioCorp, 2022

Figure 19-10: After-Tax IRR Sensitivity Graph

Given the relative sensitivity and impact of product pricing on project returns, Table 19-13 and Table 19-14 further summarize the financial results at different niobium and scandium price points.

For each table, the prices for the other products are held constant to their base case values. For example, when the price of niobium is raised to US$ 55.87/kg (120% of the base case value), scandium is held at an average of US$ 3,675/kg and titanium to US$ 0.99/kg.

Table 19-13 and Table 19-14 demonstrate that at a US$ 0/kg price for niobium, the project retains a US$ 706 million NPV (pre-tax) and a US$ 482 million NPV (after-tax). For scandium, the project’s break-even pricing is US$ 1,028/kg (pre-tax) and US$ 1,130 (after-tax). These scandium prices represent, respectively, 28% and 31% of the base pricing.

Table 19-13: Niobium Price Sensitivity (Sc and Ti Prices Remain Constant)

Niobium Pricing (US$/kg)	% of Base Model	Pre-Tax NPV (US$ million)	Pre-Tax IRR	After-Tax NPV (US$ million)	After-Tax IRR
$60.55	130%	$3,447	33.2%	$2,866	31.3%
$55.88	120%	$3,238	31.9%	$2,696	30.1%
$51.22	110%	$3,029	30.6%	$2,523	28.9%
$46.56	100%	$2,819	29.2%	$2,350	27.6%
$41.90	90%	$2,610	27.8%	$2,177	26.3%
$37.24	80%	$2,401	26.4%	$2,003	25.0%
$32.57	70%	$2,192	25.0%	$1,829	23.7%
$27.91	60%	$1,982	23.6%	$1,655	22.3%
$23.25	50%	$1,773	22.1%	$1,481	21.0%
$11.59	25%	$1,250	18.3%	$1,044	17.4%
0.00	0%	$727	14.2%	$598	13.5%

Source: NioCorp, 2022

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Table 19-14: Scandium Price Sensitivity (Nb and Ti Prices Remain Constant)

Scandium Pricing (US$/kg)	% of Base Model	Pre-Tax NPV (US$ million)	Pre-Tax IRR	After-Tax NPV (US$ million)	After-Tax IRR
$4,776	130%	$3,896	36.2%	$3,210	33.9%
$4,408	120%	$3,537	33.9%	$2,926	31.9%
$4,041	110%	$3,178	31.6%	$2,641	29.8%
$3,674	100%	$2,819	29.2%	$2,350	27.6%
$3,306	90%	$2,461	26.8%	$2,059	25.4%
$2,939	80%	$2,102	24.3%	$1,766	23.1%
$2,572	70%	$1,743	21.7%	$1,473	20.7%
$2,204	60%	$1,384	19.1%	$1,176	18.3%
$1,837	50%	$1,025	16.4%	$875	15.7%
$1,469	40%	$667	13.6%	$570	13.1%
$1,102	30%	$308	10.7%	$261	10.4%
$918	25%	$128	9.1%	$99	8.9%
$814	22.15%	$26	8.2%	$5	8.0%

Source: NioCorp, 2022

Discount rate sensitivity is always important in a project valuation, and with respect to this Project, there is a complex process plant flow sheet and market uncertainty to account for. NPV profile charts are presented in Figure 19-13 and Figure 19-14, which shows pre- and after-tax NPV results for 100 basis point increments between 0% and 20%. It should be noted that with current assumptions, the Project breaks even at a ~20% hurdle rate on an after-tax basis.

Source: NioCorp, 2022

Figure 19-11: Before-Tax NPV Profile

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Source: NioCorp, 2022

Figure 19-12: After-Tax NPV Profile

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20. ADJACENT PROPERTIES

There are no significant properties adjacent to the Elk Creek Project.

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21. OTHER RELEVANT DATA AND INFORMATION

21.1 Project Implementation Plan

The key project objectives are to:

● Deliver the Elk Creek Mine Project on time and on budget.

● Ensure to meet environmental compliance.

● Ensure the safety of all Project stakeholders.

● Ensure compliance with all applicable laws and regulations, at the local, state and federal levels.

● Ensure positive economic impacts for SE Nebraska, including the use of local businesses wherever feasible, the employment of local residents and tax benefits for local governments.

● Maintain a high level of engagement and communication with all stakeholders.

● Ensure the Project meets design parameter objectives; throughput, quality, and operating budget objectives.

The Project Implementation Plan (PIP) execution is based on the use of two main EPCM (Engineering, Procurement and Construction Management) contractors. One contractor with responsibilities for all mining related work and a second contractor responsible for all other site-wide related work including the process related facilities. Certain portions of the site-wide work will be performed with EPC sub-contracts awarded to companies that specialize in process and technology related packages, such as the acid plant. The approach is reflected in the capital cost estimate for the Project.

21.1.1 Project Cost Objectives

Table 19-8 and Table 19-9 present the cost of the Project by the main category of work. The cost objective of the Project is to reach 100% of production capacity within the total cost of US$ 884 million (includes gross pre-production revenue credit). Numbers are rounded to the nearest thousand.

21.1.2 Project Schedule Objectives

The scheduling objective is to deliver a fully constructed and commissioned mining facility as per the following timeline.

The project timeline is based on achieving the First Metal milestone at 39 months after Authorization to Proceed, plus an additional six months of the ramp-up to 100% of production capacity for a total Project schedule lasting 45 months.

The schedule highlights are as follows:

● The total duration of the project is 45 months from Authorization to Proceed to the end of the ramp-up period.

● A six-month ramp-up period (included in the overall schedule) is allotted to increase the site throughput to 100% of nameplate rating.

● The Project timeline is linked to both the mining-related activities and the surface operations in both sequencing and duration. The construction of the main surface plant buildings and supporting infrastructure is not on the critical path.

● The critical path activities include the following:

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○ Completion of drilling, sampling and final hydrogeological investigation.

○ Engineering and procurement for shaft sinking and mining components.

○ Construction of temporary power plant for shaft sinking and construction activities along with the temporary Freeze Plant for shaft sinking activities

○ Establishing commercial natural gas and electricity service to the Project site.

○ Sinking both the production and ventilation shafts, including freezing.

○ Underground mine infrastructure.

○ Completion of commissioning of the processing plant up to First Metal.

○ Ramp up of processing plant to full production capabilities.

Table 21-1 provides a summary of key activities leading up to First Ore (deemed “Advance of”), while all activities completed after First Ore are deemed “Post.”

The Project Pre-production Schedule, makes use of a monthly timescale, utilizing a declining monthly countdown (i.e., Authorization to Proceed is 39 months in advance of First Ore (-39)).

Table 21-1: Key Project Milestones

Activity	Completion Month (With Respect to First Ore)
Full Project Authorization	38 (Advance of)
Shaft Freezing to Limestone/Carbonatite Interface	29 (Advance of)
Commence Production Shaft Sinking	24 (Advance of)
Commence Ventilation Shaft Sinking	24 (Advance of)
Production Shaft Sinking Complete	20 (Advance of)
Natural Gas Available	11 (Advance of)
Production Shaft Sinking Complete	20(Advance of)
Ventilation Shaft Sinking Complete	16 (Advance of)
Permanent Power Available	11 (Advance of)
Water Treatment Plant Construction Completion	6 (Advance of)
Mineral Processing Construction Completion	4 (Advance of)
Underground Pre-Production Development Complete	0
Underground Major Infrastructure Complete	0
First Ore	0
Water Treatment Plant Commissioning Completion	3 (Post)
Hydromet Commissioning Completion	3 (Post)
Acid Plant Commissioning Completion	4 (Post)
HCl Regen Commissioning Completion	4 (Post)
Pyromet Commissioning Completion	4 (Post)
First Metal	4 (Post)

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Acid Plant Commissioning Completion	4 (Post)
Full Mill Production Begins	8 (Post)

Source: NioCorp, 2022

21.1.3 Early Works

Project Execution requires key early work that includes the following:

● Finalize contracting approach and contract key contractors (EPCM).

● Optimization metallurgical testing to confirm details for plant detailed design.

● Permitting activities required for early works construction activities.

● Complete the drill program to finalize the hydrogeological and geotechnical reviews in order to properly locate both shafts, and determine the final approach to freezing and shaft sinking.

● Perform detailed engineering and procurement of long lead time items as available.

● Commence construction on the third-party natural gas pipeline and electric power supply by Omaha Public Power District.

21.1.4 Project Team

As previously stated, the primary execution of the Project will be performed with an EPCM approach utilizing one specialized EPCM contractor to manage and execute the construction of the mining related items and a second EPCM contractor to manage and execute the remainder of the Project construction. The two EPCM contractors will report to the NioCorp Project Sponsor.

The Owner’s Project team organization will mirror the EPCM management structure. As presented in Figure 21-1, the Owner’s team will have environmental, safety, and permitting staff, personnel for engineering, finance, controls, procurement oversight, construction management including scheduling and reporting personnel. The commissioning personnel will include key vendors and NioCorp operations personnel that will continue after the construction effort is completed, in order to operate the plant.

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Source: NioCorp, 2019

Figure 21-1: Summary Level – Owner’s Project Team

The NioCorp corporate team will remain in Denver, CO with a Project Team located both on the Elk Creek Mine site and in the Company’s offices in nearby Tecumseh, NE.

21.1.5 Project and Document Control

NioCorp will utilize a project controls system for monitoring, reporting, and controlling the Project schedule, the cost, and the scope of work (change management).

● The NioCorp Project Team will be responsible for establishing the project controls procedures and assuring its consistent application throughout the Project timeline.

● The Project team will also develop a control budget to aid in managing the overall effort and will develop an appropriate Project accounting system to meet the Project needs.

● The accounting system will be used to baseline the Project cost and aid in forecasting cash flow needs. The system will aid in the creation of Earned Value Reporting (EVR) for the Project.

● The Project team will maintain the Project schedule with the use of scheduling software such as PrimaveraTM P6 or equivalent. The schedule will be updated on a regular basis to track Project progress, note any deviations.

● Change management will also be a function of the Project controls system and will be used to identify and track changes in the scope of work throughout the course of the Project.

● The Project controls system will provide Project KPIs (Key Performance Indicators) through dashboards, monthly reports, and management reports. KPIs will be determined by management in conjunction with the EPCM contractors to measure Project success.

21.1.6 Engineering

Following the completion of the Feasibility Study, design and engineering activities will be undertaken by engineering consultants. The design and engineering activities will be managed by the EPCM Contractor Engineering Managers and will be divided as follows:

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● Mine water management

● Mining and mine infrastructure

● Tailings facility

● Paste backfill

● Processing plants

● Above-ground infrastructure

● Water Treatment Plant

● HCI Regeneration Plant

● Acid Plant

21.1.7 Supply Chain and Procurement

The supply chain management responsibilities will reside with the EPCM Contractors. These duties include procurement, contracting, site material management, and development and management of work packages. The contracting strategy will include the use of a “pre-qualified bidders list” and contracts that are fixed-price, lump sum or time and materials (T&M), as the work package dictates. As mentioned previously, certain packages will be turnkey EPC contracts. The EPCM contractors will perform procurement work consisting of:

● Development of the Long Lead Equipment list.

● Development of site-wide procurement needs and packages.

● Development of Equipment Procurement Packages.

● Procurement of goods and services as required.

● Administration of purchase orders.

● Expediting of deliveries.

● Quality Control of Fabrications.

● Logistics.

The key long lead-time equipment currently identified are:

● Mine hoists, and conveyances.

● Mine substation transformer.

21.1.8 Construction Management

The EPCM contractors will perform construction management functions, including planning, organizing, and resolving issues involving the site contractors. Ensuring that contractors’ work is performed according to the Project’s safety, quality, schedule, and cost requirements. Additionally, the EPCM contractors are required to provide the facilities and services, including security, to support the sub-contractors. This practice will ensure that quality standards are maintained and will improve the use of shared resources and equipment. The primary functions include planning and coordination, contractor management, quality assurance, resolving design engineering issues, quantity measurement, and materials management.

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21.1.9 Commissioning, Operational Readiness, and Early Operations

Commissioning

The Owner’s team, in conjunction with the EPCM Contractors, will be responsible for commissioning activities. The team will develop a detailed commissioning plan during the course of the work that will address the following:

● Lists of Handover Packages & Commissioning Systems.

● Transition process.

● Commissioning Sequence.

● Alignment of Boundaries between Handover Packages and Construction Work Packages.

● Commissioning Schedule.

● Roles and Responsibilities.

● Scope of Work Alignment.

● HES Management for Commissioning.

● Handover Documentation.

● Vendor Management.

● Monitoring of Inspection and Testing performed by work Contractors.

● Commissioning Deficiencies.

● Acceptance process.

● Reporting.

The team will also partner with other key stakeholders (vendors, and suppliers) to complete the commissioning effort to hand over the Project to operating personnel for early operations and ramp up.

Operational Readiness and Ramp-up

Two Operations Readiness Plans will be prepared: the first Plan will be specific to the operation of the mine; the second Plan will be specific to the surface plant.

Training on equipment (both factory-based and on-site) will be provided by vendors. Request for quotations will require all vendors to supply Operation and Maintenance manuals, lists of spare parts for the first year of operation, list of commissioning spare parts, and training manuals. Vendors may be requested to perform on-site training based on the complexity of the equipment and/or its controls.

Ramp-up consists of bringing the plant production from First Metal, achieved by commissioning of the plant, up to 100% of commercial capacity. For the purpose of ramp-up, commercial capacity involves the production of Superalloy materials at 80% of the facility nameplate capacity, in salable quality.

NioCorp internal resources will execute the ramp-up under the responsibility of the Owner’s Team.

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Early Operations

The Project has a number of early operational tasks that are required at the onset of the Project. These activities include:

● Main Plant and mine electrical substations.

● Natural gas distribution.

● Temporary Freeze Plant.

● Temporary and permanent electrical distribution.

● Specific areas may be operated and maintained by internal Project staff or by third-party work contractors to be decided on a case by case basis. Early operations activities are included through the completion of the Project.

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22. INTERPRETATIONS AND CONCLUSIONS

22.1 Introduction

The QPs note the following interpretations and conclusions in their respective areas of expertise, based on the review of data available for this TRS.

22.2 Geology & Mineral Resource

In Understood’s opinion, the geological setting, mineralization style, and structural and stratigraphic controls are sufficiently well understood to provide useful guides to exploration and Mineral Resource estimation.

The Elk Creek Carbonatite intruded older Precambrian granitic and low to medium grade metamorphic basement rocks. Elevated niobium and titanium concentrations are directly related to magnetic mineralization in the Carbonatite, and anomalous scandium grades are spatially associated with the magnetic mineralization. The magnetic mineralization is observed to be continuous along a northwest to southeast trend with an average thickness of 200 metres. Rare earth concentrations are noted to increase from southwest to the northeast, across the trend of magnetic carbonatite domain. Three wireframes were constructed for the deposit to reflect the geologic and grade observations using the available drilling data. Block model estimation was completed in Vulcan using 5 m by 5 m by 5 m blocks that encompass the wireframes, as summarized in Section 11.

Understood classified the Mineral Resource into Indicated and Inferred Resources categories based on geological and grade continuity as well as drill hole spacing. The Mineral Resource Estimate has been reported based on NSR cut-off grade to reflect processing methodology and assumed revenue streams from Nb₂O₅, TiO₂, and Sc for the deposit. The Mineral Resource features the addition of REO to the estimate. Furthermore, the Mineral Resource also represents an increase in contained scandium and titanium metal and a decrease in contained niobium metal. Additional material exists in the geological model, which has not been classified as Indicated or Inferred resource.

The deposit remains open along strike in both directions and at depth, and there exists significant resource expansion potential based both on these factors as well as areas of the block model that require improved definition through diamond drilling.

Understood is not aware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the Mineral Resource Estimate that is not discussed in this TRS.

A variety of factors may affect the 2022 Elk Creek Mineral Resource Estimate, including but not limited to: changes to product pricing assumptions, re-interpretation of geology geometry and continuity of mineralization zones, mining and metallurgical recovery assumptions, and additional infill or step out drilling. In Understood’s opinion, the estimation methodology is consistent with standard industry practice and the Indicated and Inferred Mineral Resource Estimates for Elk Creek are reasonable and acceptable.

22.3 Exploration, Drilling and Analytical Data Collection in Support of Mineral Resource Estimation

Exploration completed to date has resulted in the delineation of the Elk Creek Deposit and a number of exploration targets.

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Dahrouge is not aware of any drilling, sampling, or recovery factors that could materially impact the accuracy and reliability of the results. In Dahrouge’s opinion, the drilling, core handling, logging, and sampling procedures meet or exceed industry standards and are adequate for the purpose of Mineral Resource Estimation.

The QA/QC protocols in place for the Project are considered acceptable and in line with standard industry practice. Based on the data validation and the results of the standard, blank, and duplicate analyses, Dahrouge is of the opinion that the assay and density databases are of sufficient quality for Mineral Resource Estimation at the Elk Creek Deposit.

No limitations were placed on Dahrouge’s data verification process, and it considers the resource database reliable and appropriate to support a Mineral Resource Estimate.

22.4 Processing and Metallurgical Testing

Mineral Processing

The Feasibility-level comminution test work was completed in two stages at SGS along with pilot scale HPGR testing at NRRI. The primary stage was conducted on six composite samples and 13 variability samples and included the determination of standard comminution parameters (SGS 2016a). The second stage of comminution test work was conducted on a single composite sample, using a LABWAL HPGR semi-pilot scale test work program (SGS 2016b). The test work results indicate that the Project ore is categorized as soft to moderately hard in terms of ore hardness, and amenable to standard grinding as well as an HPGR operation. The pilot HPGR testing indicates that the ore is amenable to processing via the HPGR. Autogenous layer buildup and flake generation were both acceptable, and there was, on average, 40% < 1 mm product generated from the HPGR when in steady state.

Hydrometallurgical Plant

Pilot test programs showed that high recovery rates of the niobium, scandium and titanium could be achieved, and that recycling and regeneration of reagents was also possible; thus, minimizing fresh reagent input and waste generation. Recoveries of 85.8% Nb₂O₅ and 93.1% Sc₂O₃ have been demonstrated while achieving 40.3% recovery of TiO₂.

Further understanding of the process was achieved with respect to the kinetics of each unit operation, which suggested that the design be adjusted. Among the changes, the following are of interest:

● The temperature of the HCI Leach was adjusted to control leaching of the iron. The Fe to Nb ratio in the Leach residue has an impact on the precipitation of Nb and the co-precipitation of titanium.

● Acid Bake total mixing and reaction time was reduced to 2.5 hours.

● Iron Reduction step was optimized based on actual reduction of Fe3+, which resulted in an improved iron consumption.

● Dilution ratio in the Niobium Precipitation was reduced from 5:1 to 0.6:1, thus reducing reagent consumption and equipment size. This, however, comes at the expense of a slight reduction in Nb recovery and an increase in Ti co-precipitation.

Secondary scandium recovery from the barren sulphate solution was developed. Selective precipitation of the scandium over impurities was achieved. Scandium precipitated in this section is combined and recovered in the Sc Solvent Extraction.

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● A scandium purification step was added that provided a 99.9% scandium product (as Sc₂O₃).

● HCI Acid Regeneration development proved that recovery of chlorides in excess 99% is achievable.

● Further test work and development provided the basis to greatly reduce the need for neutralizing reagents while increasing the recovery of sulphur; therefore, greatly reducing the need for sulphur import.

● Mixed sulphur oxide gas is treated and cleaned prior to being sent to the Acid Plant, therefore, reducing the size and cost of the Acid Plant.

Pyromet

Lab testing has confirmed most of the anticipated findings from the mathematical model that was developed by applying thermodynamic principles:

● The aluminothermic reduction of niobium oxide precipitate and iron oxide has been demonstrated. Ferroniobium particles have been formed, and the chemical proportion of iron and niobium is just what was expected.

● The change to produce a higher TiO₂ content product from the Hydromet did not change what was anticipated: Ti content in the FeNb alloy did not increase, and the reduction of Nb₂O₅ did not seem to be affected by this higher amount of titanium oxide.

● Different temperatures have been used in various tests which have provided good reference points on the slag behaviour. The Electric Arc Furnace (EAF) temperature during the operation is expected to correspond to a temperature between 1850°C and 1900°C.

22.5 Mining & Mineral Reserve

Geotechnical

A geotechnical field characterization program has been undertaken to assess the expected rock quality. This program included logging core, laboratory strength testing incl. strength of frozen soils, in situ stress measurements and oriented core logging of jointing. The results of this program have provided adequate quantity and quality data for the feasibility-level design of the underground workings.

A geotechnical assessment of the orebody shape and ground conditions has determined that long-hole open stoping mining is an appropriate mining method. Stopes have been sized to maintain stability once mucked empty. A primary/secondary extraction sequence with tight backfilling allows optimization of ore recovery while maintaining ground stability. Primary stopes will be backfilled with cemented paste backfill, while secondary stopes will be backfilled with either light-cement paste backfill or uncemented waste rock from development.

The design has been laid out using empirical design methods based on similar case histories. The stability of the mine design has been checked with 3D numerical stress-strain models of the working, which included consideration for mine-scale faulting. The modelling results confirm that stopes and access drifts are predicted to remain stable during active mining, including areas adjacent to paste backfilled primary stopes. The revised stope dimensions have been reverified using empirical design methods. The current design has not been reverified using numerical analyses, but this reverification is recommended as the mine design is advanced to the final design.

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Ground support requirements have been based on empirical ground support methods and have considered variable levels of required ground support.

The location of underground infrastructure (i.e., shafts, ventilation raises, shops, etc.) have been situated to minimize the adverse impact of encountering geologic structures (i.e., weaker faults and shear zones).

Hydrological

A geohydrological field characterization program has been undertaken to assess the expected mine water inflow conditions. This program included drilling, logging, permeability testing, injection testing, and water level measurement in hydrological boreholes, and large-scale long-term testing of the carbonatite aquifer. The results of this program have provided adequate quantity and quality data for the feasibility-level design of mine inflow control for the underground mine.

Geohydrological assessment of a range of mine inflow control methodologies (including dewatering with discharge to the Missouri River, dewatering with desalination of extracted brine, installation of a freeze-wall around the orebody, installation of a grout curtain around the mine, and grouting of inflow conduits during mining) has determined that grouting of inflow conduits from the mine during mining is the appropriate inflow control strategy for this project. Groundwater inflow from the carbonatite to the shafts, development drifts, and stopes will be controlled with injection grouting to plug water conduits, and primary stopes will be filled with low permeability cemented backfill to further limit inflow to mined-out areas (as well as to allow total extraction). Grout and cemented backfill will be prepared and piped to injection points underground. This strategy is designed to limit maximum groundwater inflow to the mine to 66 L/s (1,000 gpm), and life-of-mine average inflow to 32 L/s (500 gpm).

The general use of grouting in underground mine inflow control in fractured and karstic rock environments provides assurance that grouting will be effective. However, the mine inflow control design feasibility has not yet been verified in this carbonatite rockmass. Verification by grout injection into one or more test holes in the Elk Creek orebody is recommended as the mine design is advanced to the final design.

Mine Design

Longhole stoping is seen as the appropriate mining method for the deposit geometry. The large stope sizes minimize the mining cost. The increased dilution due to large stopes sizes is not particularly material to the mine plan as dilution has some grade.

An NSR approach was used focused on targeted amounts of Nb₂O₅ and takes into account revenue for three elements (Nb₂O₅, TiO₂, and Sc) and generates three separate products (TiO₂, FeNb, and Sc₂O₃). Stope optimization was completed to identify economic mining areas. The 3D mine design was completed on an elevated CoG, which achieved over three times the actual calculated cut-off. Two mining blocks were designed, giving a 38-year LOM, although additional material, classified as indicated, exists below the mine plan presented here.

The underground mine is accessed through a 6.0 m diameter production shaft system. A 6.0 m diameter ventilation/exhaust shaft serves as the mine exhaust, the second means of access, and second mechanical emergency egress. Both shafts are excavated using conventional shaft sinking methods in conjunction with freezing technology to an elevation 200 m below surface.

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If upon review, it is found that the overall air volume requirement increases or decreases, it is currently assumed that current shaft sizing will not change. However, an increase in air volume may require additional considerations with respect to shaft infrastructure aerodynamics and conveyance stability.

Tonnage and grades presented in the reserve include dilution and recovery and are benchmarked to other similar operations. Productivities were generated from first principles with inputs from mining contractors, blasting suppliers, and equipment vendors where appropriate. The productivities were also benchmarked to similar operations. Equipment used in this study is standard equipment used worldwide with only standard package/automation features.

A monthly and yearly production schedules were generated using Deswik scheduling software. The steady-state mine production schedule of 2,764 t/d ore was based on the processing throughput. The mine design targeted higher annual ferroniobium production during the first five years of ore delivery at full production, which resulted in an average annual production rate of 7,500 tonnes per year over this period. The steady-state average annual ferroniobium production was 7,450 tonnes during full production years

22.6 Recovery Methods

Based upon the ore body samples retained, all bench testing performed, and process analyses completed to date, L3 (hydrometallurgy), MCS (pyrometallurgy) and Megami Mining (surface plant comminution) are confident that the current design will yield the FeNb, TiO₂ and Sc₂O₃ in the quantities expected. While this level of design is feasibility, it is expected that additional design and optimization effort during the detail phase will likely yield better results and further improve the efficiency and yields of the processes.

22.7 Infrastructure

Onsite and Offsite Infrastructure

Based upon the most current operating and process design information and expectations, the on-site and off-site infrastructure and services will meet each of the required needs of this entire facility.

Infrastructure buildings, office space, locker facilities and showers were sized and designed based upon current workforce projections for the site, as well as a tentative work schedule of 12-hr shifts for shift personnel, and standard 8-hr shifts for non-shift staff. A change in the number of shifts and/or shift durations may have an impact on the requirements of these facilities.

Likewise, both potable water and wastewater distribution systems were sized based upon the above shift criteria. Changes in the number of personnel, and/or changes in numbers of shifts and shift durations may have an impact on the potable and wastewater demands which must be addressed during the detail phase of this design.

Off-site infrastructure in the form of natural gas and electrical power services provided by others are readily available, and well within the current demand requirements of the facility. Potable water sources yielding approximately 4,000 gpm are available from the local municipality (City of Tecumseh), as well as from two private landowners. The public water source would require a service extension from the existing system, while the private sources would require pipelines from the respective owner’s wells.

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Changes to the process during the detail design phase, in particular, changes to the Hydromet process could also have an impact on both potable and process water and an impact on the WWTP (Wastewater Treatment Plant). Additionally, changes to the process could have an impact on the quantities and type of reagents required, which could, in turn, change the size of storage tanks and facilities, as well as the types and materials of construction of these facilities (tanks, totes, bunkers, etc.).

Foundation designs for large loads and structures, as well as roadway designs, were based upon the most current geotechnical report and best engineering practices for the local site conditions.

The most current geotechnical report partially addressed the recommended designs for deep foundations or foundations for large loads, building columns, columns with bridge crane loads, large process equipment or structures. It will be important that the final geotechnical report address these types of loads and provide specific recommendations, but that the final geotechnical site evaluation includes test borings in the final locations of buildings, process equipment and major structures. Recommendations should further include expected settlements, as well as pavement designs with material and compaction recommendations.

22.7.1 Tailings

The tailings storage facilities (TSF) are designed for storage of dry tailings solids in lined facilities permitted under State of Nebraska Industrial Solid Waste regulations. Separate lined “leachate collection ponds” (LCPs) will be used for management of precipitation contacting the tailings solids. Based on the parameters and assumptions outlined in Section 15.5, the Plant Site and Area 7 TSFs have been designed with adequate containment and capacity to manage the planned filtered water leach residue, calcined excess oxide, and slag deposition for a 38-year LOM.

22.8 Environmental, Permitting and Social or Community Considerations

NioCorp has developed information and conducted a number of environmental studies related to baseline characterization for the Project, the most important of which are the studies related to hydrogeology and geochemistry. The production rate and geochemistry of dewatering water will dictate what is critical to the onsite water balance and any additional management (active or passive) that may be required.

The geochemistry and characterization/classification of the ore and waste materials (including the final process waste streams making up the bulk of the tailings mass and the crystallized RO water treatment salts), directly influences the management of these materials given the presence of naturally occurring radioactive materials (NORMs) (i.e., uranium and thorium) and the potential for limited reaction to contact with water. These materials currently classify as non-hazardous based on regulatory testing. Site-wide management of non-contact and contact stormwater will be essential to Project compliance.

Engagement of local, state, and federal regulators has commenced, and initial permitting to facilitate the start of project construction has been completed. Initiation of the operational permitting program for the Project is dependent upon the completion of the mine plan and surface facilities being developed as part of this technical document, as well as additional characterization of the waste materials and potential worker exposures under the jurisdiction of the Nebraska Department of Health and Human Services (DHHS) and U.S. Department of Labor — Mine Safety and Health Administration (MSHA), both of whom will have primary oversight of worker safety and monitoring programs with respect to the presence of NORMs in the ore and waste rock.

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Without specific hardrock mining regulations, there are limited obligatory requirements for reclamation and closure of mining properties in Nebraska. There are provisions, however, within the applicable regulatory framework that are likely to be applied to the Project during the permit and licensing processes, specifically those associated with the TSF and mineral processing facilities. This will include the provision of financial surety for proper closure and reclamation of the site. The 2019 estimate costs for closure and reclamation of the Project are US$ 45 million (excluding costs to secure financial assurance).

Overall, the Project appears to be sufficiently advanced to initiate the submission of formal operationsal permit applications which will govern additional aspects of construction, operation, and closure of the mine. However, given the complexity of the mine design, process operations, accelerated schedule currently envisioned by NioCorp, and the inexperience of the state regulators with this type of mining, one must recognize that risks remain within the permitting process that could slow Project development, even with the overwhelming support that the Project appears to have from the communities and stakeholders.

22.9 Market Studies and Contracts

Market studies for niobium, titanium dioxide and scandium trioxide are an important part of the proposed Elk Creek Mine. These products, especially niobium and scandium trioxide (scandium), are thinly traded without an established publicly available price discovery mechanism.

Marketing studies and product price assumptions are based on research and forecasts for the following products:

● Ferroniobium (FeNb): Roskill’s Global Industry, Markets and Outlook 2018 (Roskill, 2018)

● Scandium Trioxide (Sc₂O₃): OnG Commodities LLC (OnG, 2017, 2019) – specializes in the scandium alloys and scandium markets.

● Titanium Dioxide (TiO₂): USGS Commodity Market Summaries (Bedinger, 2019) and Adroit Market Research (Johnson, 2019).

NioCorp is considering selling ferroniobium, scandium trioxide and titanium dioxide products from the Project through all avenues, which include entering into long-term offtake contracts and Letters of Intent with buyers.

Niobium, titanium, and scandium comprises the mineral reserve supporting this Feasibility Study, as well as the mineral resource. However, the mineral resource also includes rare earth elements (REEs), which are not included in the mineral reserve for the Project. The rare earth elements (lanthanides plus yttrium), comprise a wide variety of markets, some more thinly traded and opaque than others. However, the magnetic rare earths (neodymium, praseodymium, terbium, and dysprosium) are more widely traded and are the primary REEs of interest for the Project. The Company has utilised market studies and forecasts from Adamas Intelligence (Adamas Intelligence, 2019, 2020, and 2022) to support inclusion of the REEs into the mineral resource.

At the time of this report, NioCorp had entered into three off-take agreements covering 75% of the ferroniobium and 10-15% of the scandium trioxide (minimum 12 t/y) production over the first 10 operational years from the Project.

No off-take agreements have been executed at the time of the report for the titanium dioxide product from the Project. It is assumed this product and all other material not covered by an off-take agreement will be sold on a spot price, ex-mine gate basis.

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22.10 Capital and Operating Costs

The estimate has an intended accuracy of ± 15% and an overall contingency of 10%. The estimate is reported in Q1 2019 U.S. constant dollars. The primary purpose of this report is to address changes to the resource estimate to include contained rare earth elements (REE’s). A subsequent addition of the REEs to the mineral reserve and economics will require additional metallurgical work.

Total LOM capital costs, including initial, sustaining and reclamation costs, are US$ 1,609 million. The initial capital estimate is US$ 1,141 million

The operating cost estimates were developed to show annual costs for production. Unit costs are expressed as US$ 195.94/tonne processed. LOM operating costs are estimated to be 1,553 million. The operating cost varies by year, by mine location and production. The annual operating cost varies by year but averages approximately US$ 44 million per year over the LOM. The operating cost is based on a Q1 2019 cost basis.

22.11 Economic Analysis

This TRS is based on an assumption of processing 36,656 (kt) over a 38-year life of mine (LOM) to produce 171,140 tonnes of Nb in the form of ferroniobium, 3,676 tonnes of Sc₂O₃ and 431,793 tonnes of TiO₂.

On a pre-tax basis, the NPV (8% discount) is US$ 2,819 million, the IRR is 29.2%, and the assumed payback period is within 2.67 years.

On a post-tax basis, the NPV (8% discount) is US$ 2,350 million, the IRR is 27.6%, and the assumed payback period is within 2.69 years.

22.12 Opportunities and Risk Assessment

The Project’s Opportunity and Risk Analysis (Nordmin, 2019) was reviewed by NioCorp in conjunction with, Optimize Group, Dahrouge, SRK, Tetra Tech, Adrian Brown, Zachry, MCS, Cementation and L3, looking at both opportunities and risks that were identified.

The process used in the Opportunity and Risk Analysis was as follows: Each participant was provided with a semi-quantitative risk matrix where the likelihoods and consequences were assigned numbered levels that were multiplied to generate a numerical description of risk ratings. The values that were assigned to the likelihoods and consequences were not related to their actual magnitude, but to the numerical value that was derived for risk (Figure 22-1). This approach provided for a standardized grouping and generation of indicated risk ratings. Each participant worked independently and reported their findings which were then compiled and summarized below.

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Source: NioCorp, 2022

Figure 22-1: Likelihood and Consequence Matrix

22.12.1  Opportunities

Opportunities recognized during the analysis included:

Mine Operations

● Optimizing the mine plan based upon market conditions. At present, the production stopes are dictated by their niobium content. There are existing areas within the footwall zone that have high concentrations of scandium, but they have been dismissed as ore due to their lower content of niobium. If the scandium market demand remains intact and the processing plant can increase scandium throughput possibly through a separate circuit, then there would be additional ore within the existing vertical extent of the present mine design.

○ The current resource model has many resource blocks that have an NSR greater than $500/tonne that are currently not in the mine plan for they do not meet the niobium head grade requirements but do consist of high grade Scandium. As such, if market conditions change, there is an opportunity for the operation to adjust to meet the market needs. The location of the Footwall drift and associated infrastructure may need to be adjusted to maximize the available ore to be mined.

● After completion of additional diamond drilling underground and development within the ore zone, there could be a reason to increase the width of the stopes from 15 m wide to 20 m wide, if geotechnical factors allow. This would decrease ore drive development by 25%, which is the predominant development activity.

● There could be an opportunity to replace the mining contractor after approximately three years of steady-state production. After this period of time, the full requirements to obtain sustainable production levels would be understood, and the owner could replace the contractor with their own workforce. The resulting operating cost should decrease; this would be partially offset with the purchase and sustainable capital for mobile equipment.

● Once more accurate geotechnical and hydrological characteristics and condition data is available from pilot drilling at the shaft locations, the shaft sinking methodology could be optimized. If freezing is able to be avoided, this could lead to considerable cost savings.

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● Given the pricing and interest in rare earths, there is the potential to add rare earths to the mineral reserve and incorporate them into the economics for the project.

Ventilation

● There is a potential to decrease ventilation requirements. With present-day equipment manufacturing capabilities, it may be unreasonable to expect a mining contractor to equip themselves with an electric powered mucking and hauling fleet. It is reasonable to transition the diesel-powered haulage fleet to electric power as the technology related to electrification of mining fleets is rapidly developing at the time of writing. The change over to an electric powered fleet would decrease the demand for ventilation underground, however this would have to be confirmed through a review of all other ventilation-related performance needs. Any savings related to a volume reduction would need to be evaluated against the higher haulage costs to cover the more expensive equipment.

Mine Paste Backfill

● Optimization of the backfill recipe. Test work on the paste backfill demonstrated that a 2% cement binder yielded sufficient backfill strength. There is potential for an additional positive impact to the OPEX by optimizing the recipe and reducing the cement requirement by adding fly ash in a blend with cement.

Resource/Reserve Expansion Potential

● The current deposit is open in the hanging wall, foot wall and at depth and along strike. Further drilling during the infill definition drill programs can be used to determine if the ore body can be expanded.

Cost Estimating

● Use the Hydromet and Mineral Plant buildings for tanks fabricated on site.

● Consider surface-based stormwater drainage.

EPCM Phase

● Mineral Processing and Pyromet buildings: Stick-built building vs prefabricated building.

● Hydromet building: Stick-built first storey and prefabricated second storey.

● Quality: Specialized contractor for installation of the liner in tailing and active dewatering pond.

● Environmental protection: Environmental barrier at ground level during construction.

22.12.2  Risks

The Opportunity and Risk Analysis defined 58 risks and their associated potential mitigation strategies which are not presented herein but which are summarised below.

● 25 risks were considered as a pre-response consequence of moderate, major or severe and a likelihood of likely or almost certain.

○ If the action plan is initiated, the post response consequence for these high-risk items reduces to 6 risks.

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● 30 risks were considered as a pre-response consequence of minor or moderate and a likelihood of unlikely or possible.

○ If the action plan is initiated, the post response consequence for these moderate risk items reduces to 12 risks.

The major group of risks identified and which have an action planned assigned are the following:

Mine Operational Risks

● Shaft Location - Drilling pilot holes for shaft locations to determine local geological, geotechnical and hydrological characteristics and conditions that would be encountered during shaft sinking.

● Resource/Reserve and Mine Design - Significant infill definition drilling is required during construction and operations phases to determine local geological, geotechnical and hydrological characteristics and conditions in conjunction.

● Grade Control - A daily grade control monitoring program is required to maximize the value of ore mined and fed to the surface plant. The grade control process involves the predictive delineation of the tonnes and grade of ore that will be recovered by the mining team. The program will involve incorporating the results from the infill drilling program in conjunction with an underground chip sampling program to define the boundaries of mineable ore blocks and determine the daily/weekly feed grades to the plant.

● UG Ground Support/Hydrogeology – an ongoing probe hole drill program/grout program needs to be established to support mining activities and not create significant production delays. The need to develop and deploy a high-pressure grout injection system is required to protect the mine from excess inflow to safeguard the project from injury, property damage and loss of life or equipment.

● In reviewing the 2011-2015 geotechnical drilling campaign, SRK noted both good and poor quality rock. There is thus a concern about the ramp-up rate given that regions of poor ground conditions might be encountered early in the development schedule. This could result in a risk that shaft sinking could be delayed due to the combination of ground conditions and seepage inflows (even though the ground should be frozen). There is also the risk that the first development rates could be slowed by the need to install more ground support than anticipated without having room for drill jumbos.

Ventilation

● Air Requirement – Further detailed review of the ventilation design and specifically the air quantity requirements, are needed to ensure that all concentrations of potential pollutants, radon daughters, and environmental conditions, including those relating to heat stress, are adequately addressed.

Hydrometallurgical Process Risks

A summary of the recommended test work is presented below to reduce further the risks associated with the Hydromet process design. It is expected that the work would proceed in parallel with detailed engineering for the project and would take an estimated 4 months to complete. At the time of writing, the Company has contracted with L3 process development to construct a small-scale demonstration plant to complete the recommended test work and to also

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investigate the potential to recovery rare earths into commercial-grade products. This demonstration plant is scheduled to become operational in 2022.

HCl Leach

● Optimize leaching of iron (Fe) to correlate with optimum niobium (Nb) precipitation and Fe/Nb ratios– aiming for the highest recovery of Nb while preventing titanium (Ti) co-precipitation.

● Validate the method used in the aging of the HCl Leach liquor prior to scandium (Sc) Solvent Extraction.

Acid Bake – Water Leach

● Perform vendor testing and optimization of Acid Bake operations and equipment.

● Validate process control and equipment capabilities – optimizing mixing time, temperature, acid to residue ratio.

● Optimize water to residue ratio in Water Leach.

Iron Reduction

● Verify reaction kinetics and the use of briquettes.

Nb Precipitation

● Optimize FeNb ratio.

● Optimize Precipitant (dilution water) acidity to maximize Nb precipitation and Ti selectivity.

● Optimize Final Free Acid (FAT) to maximize selectivity against Ti.

Ti Precipitation

● Further test work required to maximize the removal of uranium and thorium from the Titanium dioxide product to increase its value.

Sc Precipitation

● Optimize the H₃PO₄ addition.

● Optimize the Fe addition.

● Perform locked cycle tests on the Calcium loop.

Sc Refining

● Optimize and further evaluate Zr/Nb removal using mixed organics – stripping acid.

● Optimize conditions to minimize Sc losses.

Sc oxalate Precipitation

● Verify precipitation using solid oxalic acid – optimal amount for optimal recovery.

● Optimize acidity, temperature, and g/l with solid oxalic acid.

● Optimize the washing of Sc oxalate for calcining equipment integrity.

Acid Regeneration

● Optimize the filtration – evaluate equipment and filtration media.

Sulfate Calcining

● Optimize residence time.

● Vendor testing of different equipment and assembly.

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General

● Equipment selection, material of construction and vendor guarantee testing.

● Consider a fully integrated pilot testing to be operated onsite during construction of a full-size plant to make final adjustments and equipment selection.

● Further perform process engineering during the detailed design phase.

● Perform process simulation of the yearly or monthly elemental feed composition using the METSIM model and the compositions from the mine plan.

Scandium Market Risks and Sales Plan

At the time of this report, NioCorp had entered into one offtake agreement covering scandium trioxide production from the Project.

The scandium trioxide offtake agreement is structured similarly to the Niobium contracts. The agreement has a ten-year term and a minimum of 12 t/y. At that rate, approximately 10 – 15% of the projected annual production is contracted. Further, the customer may elect to take more material in any given year above the prescribed minimum quantity.

NioCorp is also working with other potential customers at the time of writing and discussions with these potential customers are proceeding under the provisions of Non-Disclosure Agreements (NDAs). These potential customers can be separated into the following categories or final end products:

● Scandium/Aluminum alloys used in aerospace, automotive, and other applications to increase strength and allow for a reduction of weight. Interested customers are situated at various points in the supply chains for aerospace manufacturing and operation; specialty alloy manufacturing; and specialty minerals and metal brokers/distributors.

● Solid Oxide Fuel Cells. Scandium is used in the electrolyte of solid oxide fuel cells to increase the conductivity at lower temperatures, allowing for higher efficiency and longer life. Discussions with interested customers in this industry and its supply chains are continuing.

NioCorp has produced a small quantity of 99.9% pure scandium trioxide during lab pilot testing, which meets or exceeds the purity needed for virtually all mainstream commercial applications. This material has been sent and will continue to be sent to interested customers for their analysis.

Dahrouge recommends a full update to the 2017 market assessment report for scandium be completed by OnG (OnG 2017) as a next step in assessment of the market and its potential impacts to the Elk Creek Project. As the last forecasts of the market (OnG, 2019) are now three (3) years old, an update is prudent. Moreover, new entrants into the supply side of the scandium market since the last market update (OnG 2019), in addition to recent and major global events – most notably the Russian invasion of Ukraine and COVID pandemic – further support the need for a revised market assessment for what is a very opaque market.

Rare Earth Market Risks

At the time of this report, a steady increase in demand magnet feed REEs (Nd, Pr, Tb, and Dy) is forecast. NioCorp does not have any off-take agreements at present but is investigating potential customers.

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23.  RECOMMENDATIONS

23.1  Recommended Work Programs

23.1.1  Geology and Resources

Mineral resources are uncertain because of variability at all scales and sparse sampling. Geostatistical techniques can be used to quantify the uncertainty and the expected reduction of uncertainty in resources as a function of data spacing. Understood recommends that a drill hole spacing study be completed on the deposit to better inform drill hole spacing for mineral resource classification.

After completion of the drill hole study, definition drilling should be planned and executed accordingly. Metreage and allocation of drilling resources will depend on the outcome of the drill hole spacing study.

An additional five of the 48 Drillholes within the Resource Area, including EC-025, EC-033, EC-035, EC-036, and EC-051, could not be included in the current Resource Estimate, because they lack Sc, TiO₂, and REE analytical results, preventing their incorporation into the multi-element Resource. It has been recorded that original sample material for these holes could not be located for reanalysis and because they fell at the boundaries of the deposit, it was not considered priority.  It is recommended that a follow-up sample search is completed given the potential for future Resource expansion and a recently noted improved organization of the historical material Mead storage facility.

Gaps in the REE assay record are present within the upper intervals of six 2011 and 2014 drillholes (Figure 23-1 and Table 23-1). These drillhole intervals were not originally sampled, so they do not contribute to the Resource Estimate or the By-product REE Resource. Sampling of these holes would provide additional information and potential resource expansion.

Table 23-1: 2011 and 2014 Drillhole Intervals Not Sampled and their Priority (Low to High)

Drillholes	Assays Missing	From (m)	To (m)	Length (m)	Priority
NEC11-002	Nb2O5, TiO2, Sc, REE	220.07	600.01	379.94	low
NEC11-003	Nb2O5, TiO2, Sc, REE	197.82	290.27	92.45	Moderate-Low
NEC14-020	Nb2O5, TiO2, Sc, REE	210.52	237.13	26.61	High
NEC14-021	Nb2O5, TiO2, Sc, REE	203.79	432	228.21	Moderate-Low
NEC14-022	Nb2O5, TiO2, Sc, REE	194.47	388.48	194.01	Moderate-Low
NEC14-023	Nb2O5, TiO2, Sc, REE	190.69	294	103.31	Moderate-low
Total (m)				1024.53

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Figure 23-1: Resource Area Drillhole Intervals Showing Assay Coverage (Red) and Assay Gaps (Blue).

Dahrouge recommends external pulp duplicate re-samples with complete REE analysis at a third-party check laboratory to complete QA/QC validation

A low- to moderate- priority recommendation is to assay select 2011-2014 drillhole intervals that were excluded from the 2011 and 2014 programs, should the drillholes be unpacked from storage in the future.

● Drillhole intervals should be assigned priorities based on deposit location and proximity to mineralization.

● Sampling of these intervals will provide lower-cost infill information that will increase estimation confidence for REE and Sc, and ensure they meet comparable data density to Nb₂O₅ and TiO₂ results

● No work program and cost estimate has been provided for this recommendation since it is subject to access to core.

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Dahrouge recommends external pulp duplicate re-samples with complete REE analysis at a third-party check laboratory to complete QA/QC validation.

Dahrouge recommends that life-of-mine infrastructure, such as the production shaft, refuge stations, and ventilation shafts, be further investigated to characterize the local structural, geological, geotechnical, and hydrogeological/hydrological conditions. Investigations should include, but not be limited to, the use of geotechnical logging, hydrogeologic testing (i.e., packer testing), and acoustic televiewer logging. A scoping level study that includes input from a geotechnical engineer and a structural geologist should be completed to inform the underground investigations.

23.1.1.1  Quality Assurance/Quality Control

Dahrouge recommends the following quality assurance/quality control procedures be created and followed:

● At least three certified reference material samples (CRMs) to be consistently included during sampling, comprised of the low, medium, and high values for the standardized assay.

● A clear protocol to manage CRM failures.

● Regular monitoring of the high/low CRM bias on an ongoing basis.

● A clear audit trail for re-assay.

● Perform the analysis on the 2011 assay program, which did not include selected re-assays.

● To track samples through the assay process, a work order is to be included in the assay summary sheet.

● Submit to SGS an additional, comprehensive set of samples with CRMs, explicitly focusing on mining grade ranges between 0.5 and 1.5% Nb₂O_5, to determine if a bias exists and if correction factors may be required.

● Local standards should be created for Nb₂O₅, TiO₂, and Sc using material from the Project site. This would eliminate the use of standards that are not appropriate for the deposit both from a grade and chemistry perspective.

23.1.2  Hydrometallurgical Plant

Adequate test work was conducted to support a feasibility-level design for the Hydromet Plant, and all sections of the process have been successfully tested at the pilot scale required for a Feasibility Study. However, optimization was not achieved in all areas, and certain areas will certainly benefit from further “post-Feasibility Study” test work, preferably before detailed engineering activities begin. A number of factors have not been optimized in this study, and further testing will be preferable to achieve optimal results. Such optimization could also be achieved with the help of the process simulation of the yearly or monthly elemental feed composition using the METSIM model and the compositions from the mine plan.

A summary of the recommended test work is presented below.

● HCl Leach

○ Optimize leaching of Fe to correlate with optimum FeNb ratio and Nb Precipitation – aim to best recovery of Nb while preventing Ti co-precipitation.

○ Optimize the method used in the aging of the HCl Leach liquor prior to Sc Solvent Extraction.

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● Acid Bake – Water Leach

○ Perform vendor testing and optimization of Acid Bake operations and equipment.

○ Optimize process control and equipment capabilities – optimizing mixing time, temperature, acid to residue ratio.

○ Optimize water to residue ratio in Water Leach.

● Iron Reduction

○ Verify reaction kinetics and the use of briquettes.

● Nb Precipitation

○ Optimize FeNb ratio.

○ Optimize Precipitant (dilution water) acidity to maximize Nb precipitation and Ti selectivity.

○ Optimize final free acid (FAT) to maximize selectivity against Ti.

● Ti Precipitation

○ Further test work required to maximize Th/U removal from the titanium dioxide product to increase its value.

● Sc Precipitation

○ Optimize the H₃PO₄ addition.

○ Optimize the Fe addition.

○ Perform locked cycle tests on the Calcium loop.

● Sc Refining

○ Optimize and further evaluate Zr/Nb removal using mixed organics – stripping acid.

○ Optimize conditions to minimize Sc losses.

● Sc oxalate Precipitation

○ Verify precipitation using solid Oxalic Acid – optimal amount for optimal recovery.

○ Optimize acidity, temperature, and g/l with solid Oxalic Acid.

○ Optimize the washing of Sc Oxalate for calcining equipment integrity.

● Acid Regeneration

○ Optimize the filtration – evaluate equipment and filtration media.

● Sulphate calcining

○ Optimize residence time.

○ Vendor testing of different equipment and assembly.

● General

○ Equipment selection, material of construction and vendor guarantee testing.

○ Further perform process engineering during the detailed design phase.

○ Perform process simulation of the yearly or monthly elemental feed composition using the METSIM model and the compositions from the mine plan.

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● Rare Earths

○ Bench and pilot scale testing, up to and including a small scale demontration plant, to verify the metallurgical performance, product quality and metallurgical recovery for targeted rare earth commercial products

23.1.3  Geotechnical

To advance to the final mine design, additional characterization data will be required to reduce geotechnical uncertainty. A2GC recommends the following characterization and design activities:

● Drill holes at the final shaft and ventilation raise location to confirm ground conditions for the shaft ground support, including sampling and lab testing on frozen ground.

● An additional 4 to 6 geotechnical drill holes in the footwall infrastructure and planned stope mining areas to verify the range of expected ground conditions. This includes collecting:

○ RMR/Q data

○ Structural orientation data

○ Updating the structural model and geotechnical models

○ Updating mine design parameters

● Additional geotechnical drill holes to characterize ground conditions for the final alignment of the ramps and footwall drives. These holes should be drilled from underground after the shaft is constructed and the initial access drives are mined.

● The geotechnical model should be updated to reflect the additional characterization information from new drill holes.

● The numerical model of stope stability should be reanalyzed given the revised stoping sequence. This analysis would consider any new characterization information in the geotechnical model and recent adjustments to underground infrastructure and development.

● A Ground Control Management Plan (GCMP) should be developed for guiding the initial underground development activities. This plan should include plans for geotechnical monitoring Triggered Action Response Plans (TARPs) specific to ground control, including a TARP for sudden groundwater inflows and grouting plans.

23.1.4  Mining and Reserves

The addition of rare earths to the mineral reserve should be evaluated by the company and the mineral reserve updated with a rare earth component if technically and economically feasible.

Mine Design – Ventilation

Following BBE’s review, it is recommended that a more in-depth and broader review be undertaken on the ventilation design and its optimization specifically addressing:

● Thermal conditions that could be encountered underground during the summer,

● Any need to manage radiation exposure requiring consistent ventilation though open areas,

● A more detailed study of clean engine technologies and battery electric equipment to control diesel particulate matter emissions,

● The air load diversity during concurrent development and production stages,

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● The shift air load diversity and the capacity for it to be managed through ventilation on demand strategies.

● Production shaft air velocities and the influence of conveyance velocities on airflows.

23.1.5  Recovery Methods

Mineral Processing

Following the completion of HPGR Pilot testing in 2021, no additional testwork is necessary.

Hydrometallurgical Plant

Any additional work required is included in the detailed engineering scope of work and included in the feasibility cost.

Pyrometallurgical Plant

Even though the testing has shown good results and is aligned in accordance with the mathematic model developed using thermodynamic calculations, a few minor issues remain to be addressed:

● Optimize the capacity of the Hydromet to increase the proportion of Nb₂O₅ in the precipitate. A target ratio of Nb₂O₅ / TiO₂ of 1 would be suitable.

● Perform large scale testing with an EAF to ensure good separation of slag/metal liquid and ensure the homogeneity of the ferroniobium alloy.

● Develop a flux that will enhance the fluidity of the slag at 1850°C and 1900°C.

● Select a material for the refractory that will resist the aluminothermic conditions in the Electrical Furnace.

23.1.6  Infrastructure

General Infrastructure

Tetra Tech recommends additional geotechnical investigation and design recommendations based upon the detail design requirements addressed in Section 22.5; borings in the selected building and large equipment locations, high load and deep foundation recommendations, as well as pavement design recommendations based upon the type and frequency of vehicle traffic.

Any additional work required is included in the detailed engineering scope of work and included in the feasibility cost.

Tailings

The detailed design phase of the Plant Area and Area 7 TSFs will include characterization of any additional tailings materials, additional geotechnical characterization of Plant Area TSF foundation and borrow materials, and confirmation of feasibility-level containment, water balance and stability design.

Any additional work required is included in the detailed engineering scope of work and included in the feasibility cost.

Salt Management

The final salt product will be characterized for solubility, runoff chemistry, and geotechnical characteristics to aid in the detailed design of the proposed salt management cells.

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Paste Backfill

Optimize Group recommends additional testing for the optimum paste backfill mixture during the next phase of the project. Additional testing could help further reduce the cement content, maximize early strength gain and minimize the paste backfill plant operating cost. By doing so early, the design for the paste backfill plant can be modified to allow for the addition of the relatively cheap, locally available fly ash as a binder, and for the paste backfill recipe to be perfected prior to detailed design or construction.

Additionally, investigating synergies between the surface plant and underground mining development schedules could potentially allow for waste rock produced during the mine development to be used in site based concrete manufacture, and potentially for concrete being produced on site in a modified paste backfill plant design. This opportunity, if deemed feasible, could potentially result in savings when compared to the purchase of concrete aggregate as well as concrete through a third-party supplier.

23.1.7  Environmental and Social

With respect to environmental, permitting and social/community issues for the Project, Olsson provides the following recommendations to NioCorp:

● Remain engaged and transparent with Bold Nebraska and other stakeholders/non-governmental organizations throughout the permitting process and provide them with an opportunity to participate in any public meetings or town hall discussions. This tends to garner less opposition when it comes time for formal public comments on permit applications.

● Complete more detailed hydrogeological investigations of the orebody to more accurately and precisely define the quantity and long-term quality of dewatering water expectations, and assess the feasibility of RO water treatment brine reinjection.

● Continue characterization work on the mine waste rock, process tailings, and RO water treatment crystallized salt materials in order to define the extent and partitioning of radionuclides more precisely. Assess the potential effects of the exothermic reactions from the hydration of the calcined tailings materials on the overall TSF facility, worker safety, and surrounding environment, including the potential for rad-containing, fugitive dust generation.

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23.1.8  Hoisting Plants

A review of the hoisting plants was conducted in order to evaluate the opportunity to optimize the current hoist plant designs. The result show that all four hoist plants (three hoist plants serving the production shaft, one serving the ventilation shaft) may be reduced in size and capacity all while maintaining the designed hoisting rate of material, equipment, and personnel.

Some of the design optimizations that are worth noting are as follows below.

Production Shaft – Skip Hoist

● The drum sizes may be reduced from 157.5 in diameter x 60 in wide, to 144 in diameter x 54 in wide

● The installed motor power may be reduced from 2,500 HP to 2,000 HP

● The linepull rating may be reduced from 80,000 lb to 50,000 lb

● The rope size may be reduced from 1.693 in to 1.625 in

● The skip payload may be reduced from 24,500 lb to 20,000 lb

● Hoisting speed may be reduced from 2,400 fpm to 2,200 fpm

These optimizations will lead to reduced CAPEX on the hoist plant related equipment, as well as adjacent systems and structures. A 10-12% reduction in electrical power consumption can also be expected from the revised hoist plant.

Production Shaft – Service Hoist

● The drum sizes may be reduced from 157.5 in diameter x 60 in wide, to 144 in diameter x 54 in wide

● The linepull rating may be reduced from 100,000 lb to 62,000 lb

● The rope size may be reduced from 2.0 in to 1.625 in

● The cage payload may be reduced from 44,095 lb to 25,000 lb

● The hoisting capacity was previously designed at 44,095 lbs. CUSA reviewed the largest items that need to be transported through the shaft and found that a 25,000 lbs cage would be sufficient for the heaviest item or to carry 40 workers per deck.

These optimizations will lead to reduced CAPEX on the hoist plant related equipment, as well as adjacent systems and structures.

Production and Ventilation Shaft – Auxiliary Hoists

The equipment selection for both auxiliary hoist plants has been made in such a way as to share as many major components as possible. This can be accomplished since their hoisting duties are similar, thus allowing reduced design effort as well as potentially sharing common components and spare parts.

● The drum sizes may be reduced from 120 in diameter x 72 in wide, to 96 in diameter x 52 in wide

● The installed motor power may be reduced from 900 HP to 700 HP

● The linepull rating may be reduced from 35,000 lb to 24,000 lb

● The rope size may be reduced from 1.25 in to 1.125 in

These optimizations will lead to reduced CAPEX on the hoist plant related equipment, as well as adjacent systems and structures.

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23.1.9  Summary of Costs for Recommended Work

Costs for recommended work programs are summarized in Table 23-2.

Table 23-2: Summary of Costs for Recommended Work

Area	Program	Cost Estimate (US$)
Geology and Resource	No additional work or costs have been identified or recommended beyond the work outlined in the TRS.
Processing & Metallurgical Testing	Costs of the testing program for process optimization and/or vendor equipment selection have been included in the TRS cost estimate.	$1,500,000
Processing Plants mainly the Hydrometallurgical Plant	Costs of the program in Section 23.1.2 have been included in the Detailed Engineering Phase of Work included in the TRS Cost Estimate.	$1,500,000
Mining & Reserves	Addition of rare earths to the Mineral Reserve and a full update to the TRS for the project	$3,100,000
Ventilation Design	Conduct a more in-depth review of the ventilation design with trade-offs to provide the optimal system.	$75,000 to $100,000
Geotechnical	Costs of the program in Section 23.1.3 have been included in the Detailed Engineering Phase of Work included in the TRS cost estimate.	-
Recovery- Hydrometallurgical Plant	Any additional work required is included in the detailed engineering scope of work and included in the TRS cost estimate as noted in Section 23.1.5.	-
Recovery- Pyrometallurgical Plant	Further testing is recommended during the next phase of work to optimize the system (as noted in Section 23.1.5) and is included in the TRS cost estimate	-
General Infrastructure	Tetra Tech and Cementation recommend additional investigation per Section 23.1.6.	-

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Tailings	Costs of the program in Section 23.1.6 have been included in the early stages of Detailed Engineering Phase of Work included in the TRS.	$450,000
Paste Backfill	Costs associated with additional testing of the paste backfill mix design. Testwork of waste rock for use as aggregate in on-site concrete manufacture process.	$ 60,000
Environmental and Social	No additional work is identified other than that included in workplan provided for the detailed engineering phase of work and cost estimate included within the TRS capital estimate	-
Market Studies	An updated market assessment for scandium is recommended	30,000
Economic Analysis	No additional work is identified other than that included in workplan provided for the detailed engineering phase of work and cost estimate included within the TRS capital estimate	-
Total		$6,715,000, - $6,740,000

Source: NioCorp, 2022

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24. REFERENCES

24.1 References

Adamas Intelligence (2019). Rare Earth Elements: Market Issues and Outlook (Q2 2019). Adamas Intelligence. 10 pages.

Adamas Intelligence (2020). Rare Earth Magnet Market Outlook to 2030. Adamas Intelligence. 198 pages.

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Argus Media (2022). Argus Metal Pages. Retrieved Online https://www.argusmedia.com/en/metals/argus-metal-prices. Argus Media. Accessed March 29, 2019, and March 11, 2022.

Altinsel Y., Topkaya Y., Kaya Ş., Şentürk B. (2018). Extraction of Scandium from Lateritic Nickel-Cobalt Ore Leach Solution by Ion Exchange: A Special Study and Literature Review on Previous Works. In: Martin O. (eds) Light Metals 2018. TMS 2018. The Minerals, Metals & Materials Series. Springer, Cham.

Androit Market Research. (2019). Retrieved from https://adroitmarketresearch.com/industry-reports/titanium-dioxide-market.

Anzman, J.R., (1976). Interpretation of Gravity and Magnetic Data, Elk Creek Anomaly, Johnson and Pawnee Counties, Nebraska. Molycorp, Inc. June 29, 1976. 22 pages.

Agapito Associates, INC. (Nov. 2014). Determination of Secondary Principal Horizontal Stresses NioCorp Elk Creek Project Using the Downhole Overcoring Method.

Baiinfo (2022). Baiinfo Rare Earth Weekly, Issue 22-19. Bainnfo Research Centre, 35 pages.

Barton, N.R., Lien, R. and Lunde, J. (1974). Engineering classification of rock masses for the design of tunnel support. Rock Mechanics and Rock Engineering, Springer Science and Business Media, 6 (4): 189–236.

Bedinger, G.M., (2019). Titanium and Titanium Dioxide, USGS Mineral Commodity Summaries 2019, United States Geological Survey, January 2019, pp 174-176.

Belem T., Benzaazoua M., Bussie’re B., (2000) Mechanical behaviour of cemented paste backfill. In: Proceedings of 53rd Canadian geotechnical conference. Geotechnical Engineering at the dawn of the third millennium. 15–18October, Montreal, vol 1, pp 373–380.

Berendsen, P. and Weis, T., (2001). New Kimberlite Discoveries in Kansas: Magnetic Expression and Structural Setting. Transactions of the Kansas Academy of Science. 104(3-4), pp. 223-236.

Bieniawski, Z.T., (1976). Rock mass classification in rock engineering. In Exploration for Rock Engineering, (Edited by Z.T. Bieniawski), 1, 97-106.Cape Town, Balkema.

Birkett, T.C. and Simandl, G.J. (1999) Carbonatite Associated Deposits: Magmatic, Replacement and Residual; in Selected British Columbia Mineral Deposit Profiles, Volume 3, Industrial Minerals, G.J. Simandl, Z.D. Hora and D.V. Lefebure, Editors, British Columbia Ministry of Energy and Mines.

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Brookins, D. G., Treves, S. B., and Bolivar, S. L., (1975). Elk Creek, Nebraska Carbonatite: Strontium geochemistry: Earth Planet. Sci. Lett., v. 28, p. 79-82.

Brown, M.B. (2014). Endangered Species in Pawnee and Johnson Counties. Email. Nebraska: School of Natural Resources, University of Nebraska. May 27. Recipient: Scott Honan, NioCorp.

Burchett, R.R., 1982, Thickness and structure maps of the Pennsylvanian and Permian rocks across Southern Nebraska: Nebraska Geological Survey, Report of Investigations No. 7.

Burchett, R.R. and Reed, E.C., (1967). Central Guidebook to the Geology of southeastern Nebraska. The University of Nebraska-Lincoln, Conservation and Survey Division, Lincoln, Nebraska Geological Survey.

Carlson, M.P. and Treves, S.B., (2005). The Elk Creek Carbonatite, Southeast Nebraska — An Overview. Natural resources Research, Vol. 14, No. 1 March 2005. pp. 39-45.

Carlson, M.P., (1997). Tectonic implications and influence of the Midcontinent Rift System in Nebraska and adjoining areas: Geological Society of America Special Papers v. 312, pp. 231-234.

Carlson, M.P., (1992). Tectonic implications and influence of the Midcontinent. Rift System in Nebraska and adjoining areas. In: (Richard W. Ojakangas, Albert B. Dickas, John C. Green, Eds. Basement tectonics 10: proceedings of the Tenth International Conference on Basement Tectonics, Duluth, MN.

Carter, T.G., (1992). A New Approach to Surface Crown Pillar Design. Proc. 16th Can. Rock Mechanics Symposium, Sudbury, pp. 75-83.

CIM, (2014). Canadian Institute of Mining, Metallurgy and Petroleum Standards on Mineral Resources and Reserves: Definitions and Guidelines, May 10, 2014.

Clark, L and Pakalnis, R, 1997. An empirical design approach for estimating unplanned dilution from open stope hanging walls and footwalls, in Proceedings CIM AGM, Calgary (Canadian Institute of Mining, Metallurgy and Petroleum: Westmount).

Cook W.B. and Shearer, R.L., (1986). Proposed Land Retention for 1986, Elk Creek, Nebraska. February 5, 1986.

Cordell, L. and Grauch, V.J.S. (1985) Mapping Basement Magnetization Zones from Aeromagnetic Data in the San Juan Basin, New Mexico. In: Hinze, W.J., Ed., The Utility of Regional Gravity and Magnetic Anomaly Maps, Society of Exploration Geophysicists, 181-197.

Cordell, L. (1979) Gravimetric Expression of Graben Faulting in Santa Fe Country and the Espanola Basin, New Mexico. In: Ingersoll, R.V., Ed., Guidebook to Santa Fe Country, New Mexico Geological Society, Socorro, 59-64.

Drenth, B.J., (2014). Geophysical expression of a buried niobium and rare earth element deposit: The Elk Creek carbonatite, Nebraska, USA. Interpretation 2, 169-179.

Erdosh, G., (1979). The Ontario Carbonatite Province and Its Phosphate Potential. Economic Geology, Vol. 74, pp. 331-338.

FHWA, (2016). Federal Highway Administration, HY-8 v 7.5, developed in coordination with Aquaveo LLC, and Environmental Modeling Research Laboratory, FHWA Task Order DTFH61-05-T-63053, build date July 28, 2016.

Frankel, A.D., Petersen, M.D., Mueller C.S., Haller, K.M., Wheeler, R.L., Leyendecker, E.V., Wesson, R.L., Harmsen, S.C., Cramer, C.H., Perkins, D.M., and Rukstales, K.S. (2002). Documentation for the 2002 Update of the National Seismic Hazard Maps, Open-file Report 02-420.

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Fucuchima L., Frontini P., Simao P. (2022). Brazil’s CBMM to Supply Niobium Battery Cells for Horwin Electric Motorcycles. Thomson Reuters accessed through U.S. News. Available Online https://www.usnews.com/news/technology/articles/2022-03-31/brazils-cbmm-to-supply-niobium-battery-cells-for-horwin-electric-motorcycles. Accessed online May 24, 2022.

Geology (2022). REE – Rare Earth Elements and their Uses. Available online: https://geology.com/articles/rare-earth-elements/. Accessed online May 23, 2022.

Harding, B. & Deutsch, C. (2019). Change of Support and the Volume Variance Relation. Geostatistical Lessons. https://geostatisticslessons.com/lessons/changeofsupport.

King, P.B., (1969). The Tectonics of Middle North America. Hafner Publishing Company.

Li, L., and Aubertin, M., (2014). “An improved method to assess the required strength of cemented backfill in underground stopes with an open face.” Int. J. Min. Sci. Technol., 24(4), 549–558.

McBee, W., (2003). Nemaha Strike-Slip Fault Zone. Search and Discovery Article #10055.

McCallum, N.G. and Cathro, M.S., (2010). Technical Report on the Elk Creek Property, South-East Nebraska; for Quantum Rare Earth Developments Corp. August 9, 2010. 173 pages.

McPherson, Malcom J., (1993), Subsurface Ventilation and Environmental Engineering, Mine Ventilation Services, Incorporated USA; p. 9-13.

Mitchell, R.H., (2005). Carbonatites and Carbonatites and Carbonatites. The Canadian Mineralogist; December 2005; v. 43; no. 6; p. 2049-2068.

Nickson, S.D., (1992). Cable support guidelines for underground hard rock mine operations. M.A.Sc. thesis, University of British Columbia, Vancouver.

NOAA, (2013). National Oceanic Atmospheric Administration Atlas 14: Precipitation-Frequency Atlas of the United States, Volume 8 Version 2, 2013.

Nordmin Resource & Industrial Engineering, (2019). NI 43-101 Technical Report, Feasibility Study, Elk Creek Superalloy Materials Project, Nebraska, Effective Date: April 16^th, 2019, Report Date: May 29, 2019.

NRCS, (2015). United States Department of Agricultural Natural Resources Conservation Service Web Soil Survey. http://websoilsurvey.sc.egov.usda.gov/App/HomePage.htm. Accessed April 7, 2015.

OnG, (2015). Scandium: A Market Assessment, prepared for NioCorp Ltd., by OnG Commodities LLC, Belmont, MA, July 2015.

OnG, (2017). Scandium: A Market Assessment, prepared for NioCorp Ltd., by OnG Commodities LLC, Belmont, MA, April 2017.

OnG, (2019). Niocorp Inc. Scandium Feasibility Study Update, by OnG Commodities LLC, Belmont, MA, April 2017.

Palacas, J.G., Schmoker, J.W., Dawes, T.A., Pawlewicz, M.J., and Anderson, R.R., (1990), Petroleum source-rock assessment of Middle Proterozoic (Keweenawan) sedimentary rocks, Eischeid #1 well, Carroll County, Iowa, in Anderson, R.R., ed., The Amoco M.G. Eischeid #1 Deep Petroleum Test, Carroll County, Iowa, Preliminary Investigations: Iowa Department of Natural resources, Geological Survey Bureau, Special Report Series No. 2, p. 119-134.

Parker, H.M. (2012). Reconciliation principles for the mining industry. Mining Technology. v. 121. pp. 160-176.

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Potvin, Y., and Milne, D., (1992). Empirical cable bolt support design. Proceedings of International Symposium on Rock Mechanics, Sudbury, ON, Canada.

Richardson, D.G. and Birkett, T.C., (1996). Carbonatite associated Deposits; in Geology of Canadian Mineral Deposit Types, O.R. Ecstrand, W.D. Sinclair and R.I. Thorpe, Editors, Geological Survey of Canada, Geology of Canada Number 8, pp. 541-558.

Rio Tinto (2022). Rio Tinto becomes the first producer of scandium oxide in North America. Retrieved online at https://www.riotinto.com/news/releases/2022/Rio-Tinto-becomes-the-first-producer-of-scandium-oxide-in-North-America, May 23, 2022.

Roskill, (2018.). Global Industry, Markets and Outlook 2018. Market Report. Retrieved from https://roskill.com/market-report/niobium/.

Roskill, (2017). CONFIDENTIAL: Niobium: Market Outlook to 2017 - Twelfth Edition, 2013, Copyright Roskill Information Services Ltd. ISBN 978 0 86214 592 7.

Schneider, R., K. Stoner, G. Steinauer, M. Panella, and M. Humpert (Eds.), (2011). The Nebraska Natural Legacy Project: State Wildlife Action Plan. 2nd ed. The Nebraska Game and Parks Commission, Lincoln, NE

Shaw, W.J., A. Weeks, S. Khosrowshahi, and M. Godoy. (2013). Reconciliation – Delivering on Promises. Retrieved from https://www.csaglobal.com/wp-content/uploads/2015/03/2013-Shaw-Reconciliation-delivering-on-promises-APCOM-A086.pdf

Sciencebriefs (2022). Lithium versus. Titanium Batteries. Retrieved from https://sciencebriefss.com/physics/lithium-versus-titanium-batteries/ May 23, 2022.

SGS, (2017). An Investigation into Flowsheet Development for the Elk Creek Flowsheet — Acid Bake Through to Titanium Precipitation for the Elk Creek Deposit, Project 14379-015 prepared by SGS Canada Inc., July 4, 2017.

SGS, (2016). Whole Ore Pre-Leaching as Part of the Flowsheet Development for the Elk Creek Deposit, Project 14379-013 prepared by SGS Canada Inc., dated November 10, 2016

SGS, (2017). An Investigation into an Integrated Scandium Solvent Extraction Pilot Plant for the Elk Creek Project, Project 14379-014 prepared by SGS Canada Inc., May 8, 2017.

SGS, (2017). An Investigation into Acid Regeneration Pilot Plant Elk Creek Deposit Project, 14379-016 prepared by SGS Canada Inc., February 15, 2017.

SGS Canada Inc. (2016a). An Investigation into the Grinding Circuit Design Based on Bench Scale Grindability Testing for the Elk Creek Project. Project CALR-14379-008A-Final Report-Rev 1. 2nd November 2016.

SGS Canada Inc. (2016b). HPGR Characterization of a Single Sample from the Elk Creek Project. Prepared for Elk Creek Resources. Project 15953-001-Final Report. 20th December 2016.

Sisernos and Yernberg, Molycorp Internal Memo Niobium Analytical Standardization. June, (1983).

SRK, (2014b). NI 43-101 Technical Report on resources, Elk Creek Niobium Project, Nebraska, Effective Date: September 9, 2014, Report Date: November 3, 2014, Prepared by SRK Consulting (U.S.), Inc. for NioCorp Developments Ltd.

SRK, (2015b). Amended NI 43-101 Technical Report, Updated Preliminary Economic Assessment, Elk Creek Niobium Project, Nebraska, Effective Date: August 4, 2015, Original Report Date: September 4, 2015, Amended Report Date: October 16, 2015, Prepared by SRK Consulting (U.S.).

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SRK, (2017). NI 43-101 Technical Report, Feasibility Study, Elk Creek Niobium Project, Nebraska, Effective Date: June 30, 2017, Report Date: August 10, 2017.

Syberg, F.J., (1972), A Fourier Method for the Regional-Residual Problem of Potential Fields. Geophysical Prospecting, 20: 47-75. doi:10.1111/j.1365-2478.1972.tb00619.x.

Tetra Tech, (2012). Report to: Quantum Rare Earth Developments Corp, Elk Creek Nb Project, Nebraska, US, Resource Estimate Update, Document No. 1291370100-REP-R0001-02, Effective Date: April 23, 2012, prepared for Quantum Rare Earth Developments Corp. by Tetra Tech Wardrop.

Tomlinson, E., et al., (2008). Tomlinson, E., Nebraska Statewide Probable Maximum Precipitation (PMP) Study, Applied Weather Associates, LLC, Nebraska Department of Natural resources, and Metstat, Inc., 2008.

Treves, S.B., and Low, D.J., (1983). Precambrian Geology of Eastern and Central Nebraska. GSA Abstracts with Programs, north central Section, 15, No.4, pp.266-267.

USDA SCS, (1984). United States Department of Agriculture (USDA) Soil Conservation Service (SCS), Soil Survey of Johnson County, Nebraska. National Cooperative Soil Survey. 1984.

USACE, (2016). U.S. Army Corps of Engineers, Hydrologic Engineering Center, Hydrologic Modeling System (H EC-HMS) Version 4.5, 2016.

Van Gosen, B.S., Verplanck, P.L., Seal, R.R., II, Long, K.R., and Gambogi, Joseph, 2017, Rare-earth elements, chap. O of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. O1– O31, https://doi.org/10.3133/pp1802O.

Wayne, W.J. (1981). Kansan Proglacial Environment, east-central Nebraska. American Journal Science 281:375-389.

Wikipedia contributors. (2019, March 14). Gy’s sampling theory. In Wikipedia, The Free Encyclopedia. Retrieved 13:25, May 15, 2019, from https://en.wikipedia.org/w/index.php?title=Gy%27s_sampling_theory&oldid=887738023

Woolley, A.R., (1989). The Spatial and Temporal Distribution of Carbonatites. In: Carbonatites, Genesis and Evolution (K. Bell, ed.). Unwin Hyman, London, pp 15-37.

Woolley, A.R., Kempe D.R.C., (1989). Carbonatites: nomenclature, average chemical compositions, and element distribution. In: Bell K (ed) Carbonatites: genesis and evolution. Unwin Hyman, London, pp 1–13

Wyllie, P.J. and Lee, W-J., (1998). Model System Controls on Conditions for Formation of Magnesiocabonatite and Calciocarbonatite Magmas from the Mantle. Journal of Petrology, Volume 39, Number 11&12. Pp 1885-1893. 21 May 1998. 9 pages.

Xu, A., 1996, Mineralogy, Petrology, Geochemistry and Origin of the Elk Creek Carbonatite, Nebraska: Ph.D. thesis, University of Nebraska-Lincoln.

24.2  Glossary

The Mineral Resources and Mineral Reserves have been classified according to CIM (CIM, 2014). Accordingly, the resources have been classified as Measured, Indicated or Inferred, the reserves have been classified as Proven, and Probable based on the Measured and Indicated resources as defined below.

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24.2.1  Mineral Resource

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.

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.

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 the 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.

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 the 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.

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

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.

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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.

24.2.3   Definition of Terms

Table 24-1 summarizes the general mining terms potentially used in this Technical Report.

Table 24-1: Definition of Terms

Term	Definition
Assay	The chemical analysis of mineral samples to determine the metal content.
Capital Expenditure	All other expenditures not classified as operating costs.
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	The initial process of reducing the ore particle size to render it more amenable for further processing.
Cut-Off Grade (CoG)	The grade of mineralized rock, which determines as to whether or not it is economical to recover its gold content by further concentration.
Dilution	Waste, which is unavoidably mined with ore.
Dip	The 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 an orebody or stope.
Gangue	Non-valuable components of the ore.
Grade	The measure of the concentration of gold within the mineralized rock.
Hanging wall	The overlying side of an orebody or slope.
Haulage	A horizontal underground excavation which is used to transport mined ore.
Hydrocyclone	A process whereby material is graded according to size by exploiting centrifugal forces of particulate materials.
Igneous	Primary crystalline rock formed by the solidification of magma.
Kriging	An interpolation method of assigning values from samples to blocks that minimize the estimation error.
Level	A horizontal tunnel, the primary purpose is the transportation of personnel and materials.
Lithological	Geological description pertaining to different rock types.
LRP	Long Range Plan.
Material Properties	Mine properties.
Milling	A general term used to describe the process in which the ore is crushed and ground and subjected to physical or chemical treatment to extract the valuable metals to a concentrate or finished product.
Mineral/Mining Lease	A lease area for which mineral rights are held.

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Mining Assets	The Material Properties and Significant Exploration Properties.
Ongoing Capital	Capital estimates of a routine nature, which is necessary for sustaining operations.
Ore reserve	See Mineral Reserve.
Pillar	Rock left behind to help support the excavations in an underground mine.
Sedimentary	Pertaining to rocks formed by the accumulation of sediments, formed by the erosion of other rocks.
Shaft	An opening cut downwards from the surface for transporting personnel, equipment, supplies, ore and waste.
Sill	A thin, tabular, horizontal to sub-horizontal body of igneous rock formed by the injection of magma into planar zones of weakness.
Smelting	A high-temperature pyrometallurgical operation conducted in a furnace, in which the valuable metal is collected to a molten matte or dolt phase and separated from the gangue components that accumulate in a less dense molten slag phase.
Stope	The underground void created by mining.
Stratigraphy	The study of stratified rocks in terms of time and space.
Strike	The direction of the line formed by the intersection of strata surfaces with the horizontal plane, always perpendicular to the dip direction.
Sulphide	A sulphur-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 statistical representation of the characteristics (usually grade).

24.2.4   Abbreviations

The following abbreviations may be used in this Technical Report.

Abbreviation	Unit or Term
A	ampere
AA	atomic absorption
Airn2	amperes per square metre
ANFO	ammonium nitrate fuel oil
Au	gold
BATF	U.S. Bureau of Alcohol, Tobacco and Firearms
bgs	below ground surface
°C	degrees Celcius
CAA	Clean Air Act
CAPEX	capital expenditure
CIM	Canadian Institute of Mining, Metallurgy, and Petroleum
CoG	cut-off grade

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cm	centimetre
cm2	square centimetre
cm3	cubic centimetre
cfm	cubic feet per minute
ConfC	confidence code
CRec	core recovery
CRC	Cultural Resources Consulting
CRM	certified reference material
CSS	closed-side setting
CSV	comma separated values
CTW	calculated true width
°	degree (degrees)
dia.	diameter
DOL	Department of Labor
DNR	Department of Natural Resources
EIS	Environmental Impact Statement
EMP	Environmental Management Plan
EPA	U.S. Environmental Protection Agency
ft	foot (feet)
ft2	square foot (feet)
ft3	cubic foot (feet)
g	gram
g/cm3	grams per cubic centimetre
gpd	gallons per day
g/t	grams per tonne
Ga	giga-annum (1 billion years)
gal	gallon
GHG	greenhouse gases
g/L	gram per litre
g-mol	gram-mole
gpm	gallons per minute
g/t	grams per tonne
>	greater than
ha	hectare (10,000 m2)
HAP	hazardous air pollutant
HDPE	height density polyethylene

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HG	high grade
High-Ti	high titanium basalt
hp	horsepower
HTW	horizontal true width
ICP	induced couple plasma
ID2	Inverse-Distance Squared
IFC	International Finance Corporation
ILS	intermediate leach solution
IRR	internal rate of return
kA	kiloamperes
kg	kilogram
kg/m2	Kilogram per cubic metre
kg/m3	Kilogram per square metre
km	kilometre
km2	square kilometer
koz	thousand troy ounce
kt	thousand tonnes
kt/d	thousand tonnes per day
kt/y	thousand tonnes per year
kV	kilovolt
kW	kilowatt
kWh	kilowatt-hour
kWh/t	kilowatt-hour per metric tonne
<	less than
L	litre
L/s	litres per second
L/s/m	litres per second per metre
LG	low grade
lb	pound
LHD	long-haul dump truck
LLDDP	linear low-density polyethylene plastic
LOI	loss on ignition
LOM	life of mine
m	metre
m2	square metre
m3	cubic metre

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m3/h	cubic metre per hour
masl	metres above sea level
Ma	mega-annum (1 million years)
MCL	maximum contaminant levels
MDA	Mine Development Associates
µm	micrometre per micron
µRads/hour	microradian/hour
mg/L	Milligrams per litre
M	million
MJ	megajoules
mm	millimetre
mm2	square millimetre
mm3	cubic millimetre
MME	mine & mill engineering
Moz	million troy ounces
MSHA	Mine Safety and Health Administration
Mt	million tonnes
Mtpa	Million tonnes per annum
MTW	measured true width
MW	million watts
MWMP	meteoric water mobility procedure
m.y.	million years
NDEE	Nebraska Department of Environmental and Energy
NORM	naturally occurring radioactive material
NGO	non-governmental organization
NI 43-101	Canadian National Instrument 43-101
NN	Nearest Neighbour
NPDES	national pollutant discharge elimination system
NRCS	Natural Resources Conservation Service
OK	Ordinary Kriging
OP	open pit
OPEX	operating expense
opt	ounce per tonne
OSC	Ontario Securities Commission
oz	troy ounce
%	percent

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%w/w	percent mass fraction for percent mass
PENN	Pennsylvanian-aged mudstone and limestone (Pennsylvanian strata)
pCi/g	picocuries per gram
PLC	programmable logic controller
PLS	pregnant leach solution
PMF	probable maximum flood
ppb	parts per billion
ppm	parts per million
PSD	prevention of significant deterioration
QA/QC	quality assurance/quality control
RC	rotary circulation drilling
RO	reverse osmosis
ROM	run of mine
RPD	relative percentage difference
RQD	rock quality description
SEC	U.S. Securities & Exchange Commission
sec	second
SG	specific gravity
SOFC	solid oxide fuel cells
SPCC	spill prevention, control, and countermeasures
SPLP	synthetic precipitation leach procedure
SPT	standard penetration testing
st	short ton (2,000 pounds)
t	tonne (metric ton) (2,204.6 pounds)
TCLP	toxicity characteristic leaching procedure
t/m3	tonnes per cubic metre
t/h	tonnes per hour
t/d	tonnes per day
t/y	tonnes per year
TSF	tailings storage facility
TSP	total suspended particulates
UG	underground
USACE	U.S. Army Corps of Engineers
UIC	underground injection control
USGS	United States Geological Survey
UTM	Universal Transverse Mercator
V	volts

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VFD	variable frequency drive
W	watt
XRD	x-ray diffraction
y	Year

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25. RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT

The QP opinions contained herein are based on information provided to the QPs by NioCorp throughout the course of the investigations

The QPs used their experience to determine if the information from previous reports was suitable for inclusion in this TRS and adjusted information that required amending. This report includes technical information, which required subsequent calculations to derive subtotals, totals and weighted averages. Such calculations inherently involve a degree of rounding and consequently introduce a margin of error. Where these occur, the QPs do not consider them to be material.

NioCorp provided the following information to the QPs to support the preparation of this report:

● Information on land ownership and land agreements in the project area

● All assay, water quality, environmental, metallurgical, geochemical and geotechnical data analyzed by external labs

● Information on permitting requirements for the project and the status of the Company’s permitting efforts

● Information related to NioCorp’s relationships with the local community and community groups

● Market reports and market data related to niobium, scandium, titanium and the rare earth elements

● The economic model and associated calculations that define the prospective economic performance of the project

The following experts provided information to complete sections of this TRS:

Scandium Marketing and Pricing

Dahrouge relied on Dr. Andrew Matheson of OnG Commodities LLC (OnG) for input on scandium marketing and pricing. Dr. Matheson has extensive experience and expertise in the development and implementation of market assessments across a range of materials and industries over the course of 20+ years. He provides independent, expert judgment of the outlook for scandium markets and products.

The report referenced is titled “Scandium: A Market Assessment” by OnG Commodities LLC dated April 2017 updated by OnG in 2019.

Dr. Matheson’s expertise includes:

● Education: a Ph.D. from Cambridge University in theoretical physics, and an MBA from Harvard Business School.

● Consulting: management consulting for The Boston Consulting Group, and ten years as an independent consultant providing strategic, operating and market development advice to companies in the US and Asia, in industries including oil and gas, mining and metals processing, electronic materials, automotive materials. He has also served over the past five years as a contractor with Roskill Inc, a widely recognized and respected consulting firm in the field of minor metals.

● Electronics: Dr. Matheson served as general manager of a division of Cabot Corporation manufacturing consumable materials for the semiconductor industry, and also as COO of a UK-based company developing and licensing audio technology to the consumer electronics industry.

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● Mining: at Cabot Corporation. Dr. Matheson served for several years in roles overseeing the company’s global mineral development efforts in tantalum, as well as Cabot’s procurement efforts in tantalum. Thus, he has extensive experience working with smaller-scale and junior miners both as a customer and in a development role. As far back as 1998, he was investigating scandium recovery from tailings in the United States.

● Specialty Metals: Dr. Matheson is the founder and CEO of a company developing technology to produce metal powders that can provide benefits in aerospace and automotive markets. While the materials Dr. Matheson’s company is developing are not direct substitutes for competitors to scandium alloys, they are directed to the same major markets (aerospace and automotive). Commercial qualification and adoption paths are common to new materials in these industries, and Dr. Matheson’s experience is directly applicable to scandium.

Mine Engineering

Optimize Group relied on the following experts to complete his sections of this TRS. Optimize Group have reviewed the data supplied by other experts and in their professional judgement, have taken appropriate steps to ensure that the work, information, and advice from the noted experts below are sound for the purpose of this TRS:

Mineral Reserve Estimation

● Gavin Clow, P.Eng has over 10 years of experience in underground mining environments. He has contributed to the review of the current mineral reserves estimation of the report.

● Brett Stewart has been a design technician working in Mining Design for 16 years. He has a solid understanding of mining methods and is an expert in several software suites including 3D Mine Planning and Design, Datamine Block Model Import and Evaluation, Mine 2-4D EPS, and AutoCAD as well as the Microsoft Office Programs. Brett brings practical design experience allowing for the establishment of a workable mine design for the lifecycle of the ore body from feasibility through, operation and closure.

Mining Methods

● Gavin Clow, P.Eng. has over 10 years of experience in underground mining environments. He has a firm understanding of mining methods, design and mining-related development and construction. He has contributed to the updating of the current mining methods sections of this report.

● Brett Stewart has been a design technician working in Mining Design for 16 years. He has a solid understanding of mining methods and is an expert in several software suites including 3D Mine Planning and Design, Datamine Block Model Import and Evaluation, Mine 2-4D EPS, and AutoCAD as well as the Microsoft Office Programs. Brett brings practical design experience allowing for the establishment of a workable mine design for the lifecycle of the ore body from feasibility through, operation and closure.

Nordmin Engineering contributed to the design of the underground mine. In particular, Chris Dougherty, P.Eng. (Nordmin, Principal, Consulting Specialist and Civil Engineer), Gregory Menard, P.Eng. (Nordmin, Senior Mechanical Engineer) and Glen Kuntz, P. Geo. (Nordmin, Consulting Specialist Geology/Mining) are recognized for their extensive and valuable contributions.

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SRK

SRK completed the field work, technical evaluations and design work related to the tailings design, the underground geotechnical design and the environmental and permitting analysis. A2GC has relied on the previous work completed by SRK in the area of the underground geotechnical analysis in order to fulfill A2GC’s responsibility as a QP for that section of this TRS. A2GC has reviewed SRK’s work in this specific area of the project and agrees with SRK’s overall approach, the data collected, the tests conducted (including the laboratory and stress measurements), the characterization completed and the analyses done, and sees no fatal flaws.

Scott Honan, M.Sc. relied on the following experts to complete the designs summarized in this TRS related to tailings, associated infrastructure and salt management:

Project Infrastructure

● Clara Balasko, PE., is a registered professional engineer with 15 years of experience. She specializes in tailings storage facility design including slurry, paste and dry stack tailings disposal and has worked on numerous projects in the Americas, Australia and Asia.

● Dave Bentel, Pr. Eng, has more than 41 years of experience in the provision of engineering and environmental permitting services for mining facilities, including process fluid and stormwater management, tailings disposal, tailings recovery and re-treatment, heap leach, and open pit and waste rock disposal facilities. Dave has vast experience with establishing practical solutions towards mine design and closure and was involved throughout the design process.

Salt Management Cells

● Breese Burnley, PE., has more than 26 years of experience in engineering design, permitting and closure of facilities for mine water management, tailings disposal, heap leaching, and waste rock disposal. Breese brings experience with establishing a practical and innovative design for mine waste storage facilities thorough feasibility, operation and closure.

Olsson relied on the following experts to complete the initial work related to environmental and permitting matters for the project:

Environmental

● Filiz Toprak, MSc, is an SRK mining consultant with over 15 years of experience. She uses her training and background in mining engineering in projects focused on mine reclamation and closure cost estimation. She currently focuses on different types of closure cost estimates to address requirements based on financial assurance, financial reporting, and project planning. Ms. Toprak prepared the closure cost estimate for the Elk Creek Project.

Metallurgy Concept Solutions

Metallurgy Concept Solutions relied on the following experts to complete their Sections of this TRS. Metallurgy Concept Solutions have reviewed the data supplied by other experts and in their professional judgement, have taken appropriate steps to ensure that the work, information, and advice from the noted experts below are sound for the purpose of this TRS:

● Kingston Process Metallurgy (KPM) – the laboratory that performed the niobium oxide

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aluminothermic testing.

L3 Process Development

L3 Process Development relied on the following experts to complete their Sections of this TRS. L3 Process Development have reviewed the data supplied by other experts and in their professional judgement, have taken appropriate steps to ensure that the work, information, and advice from the noted experts below are sound for the purpose of this TRS:

Hydrochloric Acid Regeneration

● Mr. K. Michael (Mike) Sessions, PE, Chief Process Engineer, a chemical engineer (M.S., Tennessee Technological University, 1985) with 31 years of experience in process design, process simulation, process scale-up, plant operations, troubleshooting and management, of a variety of chemical processes including pharmaceuticals, foods, commodity and specialty chemicals, as well as in the specification and commissioning of a wide variety of process control instrumentation.

Sulphuric Acid Plant

● Mr. Douglas K. Louie, PE, Owner of DKL Engineering, an engineer with over 30 years of experience in process design, process simulation, process scale-up, plant evaluation, and troubleshooting in the sulphuric acid industry.

The Conservation and Survey Division (CSD) of the University of Nebraska Lincoln’s School of Natural Resources is a unique, multi-disciplinary research, service and data-resource organization that originated in 1893. It is Nebraska’s geological survey. CSD’s mission is to investigate and record information about Nebraska’s geologic history, its rock and mineral resources, the quantity and quality of its water resources, land cover and other aspects of its geography, as well as the nature, distribution and uses of its soils. CSD was actively involved in the discovery of the Elk Creek Carbonatite more than five decades ago. CSD continues to curate samples and data from the deposit, among its many other collections, for the benefit of stakeholders and in the public interest. CSD has been an invaluable source of data and expertise for minerals development and other Earth-science issues in Nebraska since its founding. CSD’s assistance and historic drill core repository were essential to establishing a mineral resource and a mineral reserve for the Elk Creek Project.

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26. DATE AND SIGNATURE PAGE

This TRS entitled “Technical Report Summary, Elk Creek Project, Nebraska” with an effective date of June 30, 2022 was prepared and signed by:

Dahrouge Geological Consulting USA Ltd. /s/ Dahrouge Geological Consulting USA Ltd.

Dated in Colorado, USA 

September 2, 2022

Understood Mineral Resources Ltd. /s/ Understood Mineral Resources Ltd.

Dated in Saskatchewan, Canada 

September 2, 2022

Optimize Group Inc. /s/ Optimize Group Inc.

Dated in Ontario, Canada 

September 2, 2022 

Tetra Tech /s/ Tetra Tech 

Dated in Utah, USA 

September 2, 2022

Adrian Brown Consultants Inc. /s/ Adrian Brown Consultants Inc.

Dated in Colorado, USA 

September 2, 2022

Magemi Mining Inc. /s/ Magemi Mining Inc.

Dated in Ontario, Canada 

September 2, 2022

L3 Process Development Inc. /s/ L3 Process Development Inc.

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Dated in Utah, USA 

September 2, 2022

Olsson /s/ Olsson

Dated in Nebraska, USA 

September 2, 2022

A2GC /s/ A2GC

Dated in Quebec, Canada 

September 2, 2022

Metallurgy Concept Solutions /s/ Metallurgy Concept Solutions

Dated in Oregon, USA 

September 2, 2022

Scott Honan, M.Sc, SME-RM, NioCorp /s/ Scott Honan, M.Sc, SME-RM, NioCorp

Dated in Colorado, USA 

September 2, 2022

Everett Bird, P.E., Cementation /s/ Everett Bird, P.E., Cementation

Dated in Utah, USA 

September 2, 2022

Matt Hales, P.E., Cementation /s/ Matt Hales, P.E., Cementation

Dated in Utah, USA 

September 2, 2022

Mahmood Khwaja, P.E., CDM Smith /s/ Mahmood Khwaja, P.E. CDM Smith

Dated in Massachusetts, USA 

September 2, 2022

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Martin Lepage, P.Eng, Ing., Cementation /s/ Martin Lepage, P.Eng, Ing., Cementation

Dated in Ontario, Canada 

September 2, 2022

Wynand Marx, M.Eng., BBE Consulting /s/ Wynand Marx, M.Eng., BBE Consulting

Dated in South Africa 

September 2, 2022

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27. QP RESPONSIBILITY MATRIX

Section #	Section	Company	Responsible QP
1	EXECUTIVE SUMMARY
1.1	Principal Outcomes	Dahrouge	Dahrouge
1.2	Property Description and Ownership	Dahrouge	Dahrouge
1.3	Geological Setting and Mineralization	Dahrouge	Dahrouge
1.4	History	Dahrouge	Dahrouge
1.5	Drilling	Dahrouge	Dahrouge
1.6	Mineral Resource Estimation	Understood	Understood
1.7	Mineral Reserve Estimation	Optimize Group	Optimize Group
1.8	Environmental Studies, Permitting and Social or Community Impact	Olsson	Olsson
1.9	Capital Cost Estimate	Dahrouge	Dahrouge
1.10	Operating Cost Estimate	Dahrouge	Dahrouge
1.11	Interpretation and Conclusions	Dahrouge	Dahrouge
2	INTRODUCTION	Dahrouge	Dahrouge
2.1	Terms of Reference and Purpose of the Technical Report	Dahrouge	Dahrouge
2.2	Sources of Information	Dahrouge	Dahrouge
2.3	Details of Inspection	Dahrouge	Dahrouge
2.4	History	Dahrouge	Dahrouge
3	PROPERTY DESCRIPTION	Dahrouge	Dahrouge
3.1	Property Location	Dahrouge	Dahrouge
3.2	Property Description and Land Tenure	Dahrouge	Dahrouge
3.2.1	Nature and Extent of Issuer’s Interest	Dahrouge	Dahrouge
3.3	Royalties, Agreements and Encumbrances	Dahrouge	Dahrouge
3.3.1	Required Permits and Status	Dahrouge	Dahrouge
3.4	Other Significant Factors and Risks	Dahrouge	Dahrouge
4	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY	Dahrouge	Dahrouge

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4.1	Physiography	Dahrouge	Dahrouge
4.2	Accessibility and Transportation to the Property	Dahrouge	Dahrouge
4.3	Climate and Length of Operating Season	Dahrouge	Dahrouge
4.4	Infrastructure	Dahrouge	Dahrouge
4.4.1	Personnel and Supplies	Dahrouge	Dahrouge
4.4.2	Electrical Power	Dahrouge	Dahrouge
4.4.2.1	Electrical Power Line & Substation	Dahrouge	Dahrouge
4.4.2.2	Electrical Power Distribution - Plant and Facilities	Dahrouge	Dahrouge
4.4.2.3	Electrical Power Distribution - Underground	Dahrouge	Dahrouge
4.4.2.4	Emergency Power Generation	Dahrouge	Dahrouge
4.4.3	Water	Dahrouge	Dahrouge
4.4.3.1	Process Water	Dahrouge	Dahrouge
4.4.3.2	Fire Water	Dahrouge	Dahrouge
4.4.3.3	Potable Water	Dahrouge	Dahrouge
5	HISTORY	Dahrouge	Dahrouge
5.1	Exploration History	Dahrouge	Dahrouge
5.1.1	USGS, 1964	Dahrouge	Dahrouge
5.1.2	Discovery, 1970-1971	Dahrouge	Dahrouge
5.1.3	Cominco American, 1974	Dahrouge	Dahrouge
5.1.4	Molycorp, 1973-1986	Dahrouge	Dahrouge
5.1.5	Geophysical Surveys	Dahrouge	Dahrouge
5.1.6	Drilling	Dahrouge	Dahrouge
5.1.7	Molycorp Data Verification, 1973-1986	Dahrouge	Dahrouge
6	GEOLOGICAL SETTING, MINERALISATION AND DEPOSIT	Dahrouge	Dahrouge
6.1	Regional Geology	Dahrouge	Dahrouge
6.2	Property Geology	Dahrouge	Dahrouge
6.3	Elk Creek Carbonatite	Dahrouge	Dahrouge
6.3.1	Age Dating	Dahrouge	Dahrouge

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6.4	Carbonatite Lithological Unit	Dahrouge	Dahrouge
6.5	Marine Sedimentary Rocks	Dahrouge	Dahrouge
6.6	Structural Geology	Dahrouge	Dahrouge
6.7	Mineralization	Dahrouge	Dahrouge
6.7.1	Niobium and Titanium Mineralization	Dahrouge	Dahrouge
6.7.2	Scandium Mineralization	Dahrouge	Dahrouge
6.7.3	Rare Earth Element Mineralization	Dahrouge	Dahrouge
6.8	Deposit Types	Dahrouge	Dahrouge
7	EXPLORATION	Dahrouge	Dahrouge
7.1	Geophysical Exploration	Dahrouge	Dahrouge
7.2	Drilling	Dahrouge	Dahrouge
7.2.1	Type and Extent	Dahrouge	Dahrouge
7.2.2	Molycorp, 1973-1986	Dahrouge	Dahrouge
7.2.3	Quantum, 2011	Dahrouge	Dahrouge
7.2.4	NioCorp 2014 Program	Dahrouge	Dahrouge
7.2.4.1	Procedures (NioCorp 2014 Program)	Dahrouge	Dahrouge
7.2.4.2	Collar Surveys	Dahrouge	Dahrouge
7.2.4.3	Downhole Surveys	Dahrouge	Dahrouge
7.2.5	Interpretation and Relevant Results	Dahrouge	Dahrouge
7.3	Geotechnical Design Parameters	A2GC	A2GC
7.4	Hydrogeology Design Parameters	ABC	ABC
7.4.1	Conceptual Geohydrology	ABC	ABC
7.4.2	Mine Inflow Control	ABC	ABC
8	SAMPLE PREPARATION, ANALYSES, AND SECURITY	Dahrouge	Dahrouge
8.1	Sample Preparation and Security	Dahrouge	Dahrouge
8.1.1	Molycorp, 1973 – 1986	Dahrouge	Dahrouge
8.1.2	NioCorp Drilling Program, 2011 - Current	Dahrouge	Dahrouge
8.1.3	Historical Re-Sampling Programs	Dahrouge	Dahrouge
8.1.3.1	NioCorp (Quantum 2010) Historical Re-Sampling Program	Dahrouge	Dahrouge
8.1.3.2	NioCorp (2014-2016) Historical Re-	Dahrouge	Dahrouge

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	Sampling Program
8.1.3.3	NioCorp (2021) Historical Re-Sampling Programs	Dahrouge	Dahrouge
8.2	Sample Analysis Procedures, 2011 – Current	Dahrouge	Dahrouge
8.3	Quality Assurance/Quality Control Programs	Dahrouge	Dahrouge
8.3.1	Re-Sampling/Verification of Historical Assays	Dahrouge	Dahrouge
8.3.2	NioCorp 2011 - Current	Dahrouge	Dahrouge
8.3.3	Quality Assurance & Quality Control Results	Dahrouge	Dahrouge
8.3.3.1	Field Quartz Blanks	Dahrouge	Dahrouge
8.3.3.2	Certified & Standard Reference Material	Dahrouge	Dahrouge
8.3.3.3	Reject Duplicates	Dahrouge	Dahrouge
8.3.3.4	Field 1/4 Core Duplicates	Dahrouge	Dahrouge
8.3.3.5	Third-Party Duplicate Check Analysis	Dahrouge	Dahrouge
8.4	Qualified Person’s Opinion on the Adequacy of Sample Preparation, Security and Analytical Procedures	Dahrouge	Dahrouge
8.5	Specific Gravity	Dahrouge	Dahrouge
9	DATA VERIFICATION
9.1	Understood and Optimize Group Data Validation	Understood	Understood
9.1.1	Site Visit	Understood	Understood
9.1.1.1	Drill Core Review	Understood	Understood
9.1.1.2	Collar Verification	Understood	Understood
9.1.1.3	Core Processing Protocols	Understood	Understood
9.1.2	Database Validation	Understood	Understood
9.1.3	Review of NioCorp QA/QC	Understood	Understood
9.2	Limitations	Dahrouge	Dahrouge
9.3	Qualified Person’s Opinion	Dahrouge	Dahrouge
10	MINERAL PROCESSING AND METALLURGICAL TESTING
10.1	Mineral Processing	Magemi Mining	Magemi Mining

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10.2	Hydrometallurgy	L3	L3
10.2.1	Testing and Procedures	L3	L3
10.2.2	Relevant Results	L3	L3
10.2.3	Significant Factors	L3	L3
10.3	Pyrometallurgy	MCS	MCS
11	MINERAL RESOURCE ESTIMATES	Understood	Understood
11.1	Introduction	Understood	Understood
11.2	Source Database	Understood	Understood
11.2.1	Drill Holes	Understood	Understood
11.3	Geological Domaining	Understood	Understood
11.4	Exploratory Data Analysis	Understood	Understood
11.4.1	Compositing	Understood	Understood
11.4.2	Declustering	Understood	Understood
11.4.3	Outlier Capping	Understood	Understood
11.4.4	Representative Distributions Statement	Understood	Understood
11.5	Exploratory Data Analysis	Understood	Understood
11.6	Variography	Understood	Understood
11.7	Block Model Resource Estimation	Understood	Understood
11.7.1	Estimation Overview	Understood	Understood
11.7.2	Block Model Definition	Understood	Understood
11.7.3	Estimation Strategy and Testing	Understood	Understood
11.7.3.1	Estimation Strategy	Understood	Understood
11.7.3.2	Testing and Strategy Refinement	Understood	Understood
11.7.4	Estimation/Interpolation Methods	Understood	Understood
11.8	Model Validation	Understood	Understood
11.8.1	Rare Earth Considerations	Understood	Understood
11.8.2	Global Checks	Understood	Understood
11.8.3	Visual Inspection	Understood	Understood
11.8.4	Swath Plots	Understood	Understood
11.8.5	Correlation Review	Understood	Understood
11.9	Mineral Resource Classification	Understood	Understood

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11.10	Reasonable Prospects of Eventual Economic Extraction	Understood	Understood
11.11	Cut-Off Grade	Understood	Understood
11.12	Mineral Resource Tabulation	Understood	Understood
11.13	Mineral Resource Uncertainty	Understood	Understood
11.13.1	Specific Identified Risks	Understood	Understood
11.13.2	Generic Mineral Resource Uncertainty	Understood	Understood
11.14	Mineral Resource Sensitivity	Understood	Understood
11.15	Relevant Factors	Understood	Understood
12	MINERAL RESERVE ESTIMATES	Optimize Group	Optimize Group
12.1	Conversion Assumptions, Parameters and Methods	Optimize Group	Optimize Group
12.1.1	Dilution	Optimize Group	Optimize Group
12.1.2	Recovery	Optimize Group	Optimize Group
12.1.3	Cut-Off Grade Calculation	Optimize Group	Optimize Group
12.1.4	Mine Design	Optimize Group	Optimize Group
12.2	Reserves	Optimize Group	Optimize Group
12.3	QP Opinion and Relevant Factors	Optimize Group	Optimize Group
13	MINING METHODS
13.1	Geotechnical Design Parameters	A2GC	A2GC
13.2	Hydrogeology Design Parameters	ABC	ABC
13.3	Mine Design	Optimize Group	Optimize Group
13.3.1	Selection of Mining Method	Optimize Group	Optimize Group
13.3.2	Stope Optimization	Optimize Group	Optimize Group
13.3.3	Stope Design	Optimize Group	Optimize Group
13.3.4	Development Design	Optimize Group	Optimize Group
13.3.5	Mine Access	Optimize Group	Optimize Group
13.3.5.1	Shaft Layouts	Optimize Group	Optimize Group
13.4	Production Schedule	Optimize Group	Optimize Group
13.4.1	Productivity	Optimize Group	Optimize Group
13.4.2	Shaft Sinking – Production Shaft and Ventilation Shaft	Optimize Group	Optimize Group

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13.4.3	Development and Production Schedule	Optimize Group	Optimize Group
13.5	Mining Operations	Optimize Group	Optimize Group
13.5.1	Production	Optimize Group	Optimize Group
13.5.2	Development	Optimize Group	Optimize Group
13.5.3	Truck and LHD Haulage	Optimize Group	Optimize Group
13.5.4	Backfilling	Optimize Group	Optimize Group
13.5.5	Ground Support	Optimize Group	Optimize Group
13.5.6	Grade Control and Reconciliation	Optimize Group	Optimize Group
13.5.7	Workforce	Optimize Group	Optimize Group
13.5.8	Equipment	Optimize Group	Optimize Group
13.6	Ventilation	BBE Consulting	Wynand Marx
13.6.1	Airflow Requirements	BBE Consulting	Wynand Marx
13.6.2	Ventilation Controls	BBE Consulting	Wynand Marx
13.6.3	Ventilation Model	BBE Consulting	Wynand Marx
13.6.4	Auxiliary Ventilation	BBE Consulting	Wynand Marx
13.6.5	Recommended Ventilation Infrastructure	BBE Consulting	Wynand Marx
13.6.6	Ventilation Power Consumption	BBE Consulting	Wynand Marx
13.6.7	Air Heating	BBE Consulting	Wynand Marx
13.7	Mine Infrastructure and Services
13.7.1	Material Handling System	Cementation	Everett Bird
13.7.2	Mine Dewatering System	Cementation	Everett Bird
13.7.3	Compressed Air System	Cementation	Everett Bird
13.7.4	Underground Water Supply	Cementation	Everett Bird
13.7.5	Underground Fuel Storage and Distribution	NioCorp	Scott Honan
13.7.6	Workshop, Maintenance Bays and Warehouse	NioCorp	Scott Honan
13.7.7	Explosives Storage	NioCorp	Scott Honan
13.7.8	Refuge Stations / Chambers	NioCorp	Scott Honan
13.7.9	Hoist House Substation Surface Electrical Distribution	Cementation	Matt Hales
13.7.10	Underground Electrical Distribution	Cementation	Matt Hales

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13.7.11	Overhead Pole Line Electrical Distribution	Cementation	Matt Hales
13.7.12	Hoisting Plants	Cementation	Martin LePage
13.7.13	Dust Suppression System	Cementation	Everett Bird
13.7.14	Communications System	Cementation	Matt Hales
13.7.15	Safety and Health	NioCorp	Scott Honan
14	PROCESSING AND RECOVERY METHODS
14.1	Process Plant Design Criteria
14.1.1	Surface Crushing, Ore Storage & Mineral Processing Plant	Magemi Mining	Magemi Mining
14.1.2	Hydrometallurgical Plant	L3	L3
14.1.3	Pyrometallurgical Plant	MCS	MCS
14.1.4	Acid Plant	L3	L3
14.2	Flowsheets and Process Description
14.2.1	Surface Crushing, Ore Storage & Mineral Processing Plant	Magemi Mining	Magemi Mining
14.2.2	Hydrometallurgical Plant	L3	L3
14.2.3	Pyrometallurgical Plant	MCS	MCS
14.2.4	Acid Plant	L3	L3
14.3	Process Equipment
14.3.1	Surface Crushing, Ore Storage & Mineral Processing Plant	Magemi Mining	Magemi Mining
14.3.2	Hydrometallurgical Plant	L3	L3
14.3.3	Pyrometallurgical Plant	MCS	MCS
14.3.4	Acid Plant	L3	L3
14.4	Power Requirements
14.4.1	Surface Crushing, Ore Storage & Mineral Processing Plant	Magemi Mining	Magemi Mining
14.4.2	Hydrometallurgical Plant	L3	L3
14.4.3	Pyrometallurgical Plant	MCS	MCS
14.4.4	Acid Plant	L3	L3
14.5	Plant Water	Tetra Tech	Tetra Tech
14.5.1	Water Treatment Plant	Tetra Tech	Tetra Tech

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14.5.1.1	Flow Equalization	Tetra Tech	Tetra Tech
14.5.1.2	Softening and Clarification	Tetra Tech	Tetra Tech
14.5.1.3	Multimedia Filtration	Tetra Tech	Tetra Tech
14.5.1.4	Reverse Osmosis (RO) System	Tetra Tech	Tetra Tech
14.5.1.5	Cooling Tower Makeup System (CTMU)	Tetra Tech	Tetra Tech
14.5.1.6	Sludge Handling	Tetra Tech	Tetra Tech
14.5.1.7	Evaporation and Crystallization System	Tetra Tech	Tetra Tech
14.5.1.8	Crystallizer Brine Flow	Tetra Tech	Tetra Tech
14.5.2	Process Water	Tetra Tech	Tetra Tech
14.5.3	Fire Water	Tetra Tech	Tetra Tech
14.5.4	Potable Water	Tetra Tech	Tetra Tech
15	INFRASTRUCTURE
15.1	Electrical Power
15.1.1	Electrical Power Line & Substation	Tetra Tech	Tetra Tech
15.1.2	Electrical Power Distribution - Plant and Facilities	Tetra Tech	Tetra Tech
15.1.3	Electrical Power Distribution - Underground	Cementation	Matt Hales
15.1.4	Emergency Power Generation	Cementation	Matt Hales
15.2	Natural Gas	Tetra Tech	Tetra Tech
15.2.1	Natural Gas Pipeline to Site	Tetra Tech	Tetra Tech
15.2.2	Natural Gas Distribution on Site	Tetra Tech	Tetra Tech
15.3	Plant Water	Tetra Tech	Tetra Tech
15.3.1	Water Treatment Plant	Tetra Tech	Tetra Tech
15.3.1.1	Flow Equalization	Tetra Tech	Tetra Tech
15.3.1.2	Softening and Clarification	Tetra Tech	Tetra Tech
15.3.1.3	Multimedia Filtration	Tetra Tech	Tetra Tech
15.3.1.4	Reverse Osmosis (RO) System	Tetra Tech	Tetra Tech
15.3.1.5	Cooling Tower Makeup System (CTMU)	Tetra Tech	Tetra Tech
15.3.1.6	Sludge Handling	Tetra Tech	Tetra Tech
15.3.1.7	Evaporation and Crystallization System	Tetra Tech	Tetra Tech
15.3.1.8	Crystallizer Brine Flow	Tetra Tech	Tetra Tech

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S-K 1300 Technical Report Summary

15.3.2	Process Water	Tetra Tech	Tetra Tech
15.3.3	Fire Water	Tetra Tech	Tetra Tech
15.3.4	Potable Water	Tetra Tech	Tetra Tech
15.4	Roads	Tetra Tech	Tetra Tech
15.4.1	Main Access Road to Site	Tetra Tech	Tetra Tech
15.4.2	Secondary Site Access Roads	Tetra Tech	Tetra Tech
15.4.3	Secondary Site Roads (to tailings, etc.)	Tetra Tech	Tetra Tech
15.5	Tailing Storage and Associated Facilities	NioCorp	Scott Honan
15.6	Salt Management Cells	NioCorp	Scott Honan
15.7	Paste Backfill Plant and Underground Distribution	Optimize Group	Optimize Group
15.7.1	Surface Plant	Optimize Group	Optimize Group
15.7.2	Backfill Testwork	Optimize Group	Optimize Group
15.7.3	Paste Plant Process	Optimize Group	Optimize Group
15.7.4	Underground Distribution of Paste Backfill	Optimize Group	Optimize Group
15.8	Freeze Plant	CDM Smith	Mahmood Khwaja
16	MARKET STUDIES	Dahrouge	Dahrouge
16.1	Market Studies	Dahrouge	Dahrouge
16.1.1	Niobium Market Overview	Dahrouge	Dahrouge
16.1.1.1	Niobium Supply	Dahrouge	Dahrouge
16.1.1.2	Niobium Demand	Dahrouge	Dahrouge
16.1.2	Titanium Dioxide Market Overview	Dahrouge	Dahrouge
16.1.2.1	Titanium Dioxide Demand	Dahrouge	Dahrouge
16.1.3	Scandium Trioxide Market Overview	Dahrouge	Dahrouge
16.1.4	Key Aspects of OnG Commodities Report	Dahrouge	Dahrouge
16.1.5	Rare Earth Market Overview	Dahrouge	Dahrouge
16.2	Contracts and Status	Dahrouge	Dahrouge
17	ENVIRONMENTAL STUDIES, PERMITTING, AND PLANS, NEGOTIATIONS, OR AGREEMENTS WITH LOCAL INDIVIDUALS OR GROUPS	Olsson	Olsson
17.1	Environmental Studies	Olsson	Olsson

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S-K 1300 Technical Report Summary

17.1.1	Soils	Olsson	Olsson
17.1.2	Climate/Meteorology/Air Quality	Olsson	Olsson
17.1.3	Cultural and Archeological Resources	Olsson	Olsson
17.1.4	Vegetation	Olsson	Olsson
17.1.5	Wildlife	Olsson	Olsson
17.1.6	Threatened, Endangered, and Special Status Species	Olsson	Olsson
17.1.7	Land Use	Olsson	Olsson
17.1.8	Hydrogeology (Groundwater)	Olsson	Olsson
17.1.9	Hydrology (Surface Water)	Olsson	Olsson
17.1.10	Wetlands/Riparian Zones	Olsson	Olsson
17.1.11	Geochemistry	Olsson	Olsson
17.1.12	Known Environmental Issues	Olsson	Olsson
17.1.13	Tailings	Olsson	Olsson
17.1.14	Project Waste Disposal	Olsson	Olsson
17.1.15	Site Monitoring	Olsson	Olsson
17.1.16	Water Management	Olsson	Olsson
17.1.17	Chemicals and Reagent Handling	Olsson	Olsson
17.2	Project Permitting Requirements	Olsson	Olsson
17.2.1	Nebraska Underground Injection Control (UIC)	Olsson	Olsson
17.2.2	DHHS Radioactive Materials Program and Licensing	Olsson	Olsson
17.2.3	Nebraska Air Quality Permitting	Olsson	Olsson
17.2.4	Nebraska Dam Permitting	Olsson	Olsson
17.2.5	Greenhouse Gas Permitting	Olsson	Olsson
17.2.6	Permitting Status	Olsson	Olsson
17.2.7	Post-Performance and Reclamation Bonding	Olsson	Olsson
17.3	Community Relations and Social Responsibilities	Olsson	Olsson
17.3.1	Safety and Health	Olsson	Olsson
17.4	Reclamation and Closure	Olsson	Olsson

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S-K 1300 Technical Report Summary

17.4.1	Surface Disturbance	Olsson	Olsson
17.4.2	Buildings and Equipment	Olsson	Olsson
17.4.3	Tailings Disposal Facility	Olsson	Olsson
17.4.4	Closure Cost Estimate	Olsson	Olsson
17.5	International Standards and Guidelines	Olsson	Olsson
17.6	Qualified Person’s Opinion	Olsson	Olsson
18	CAPITAL AND OPERATING COSTS	Dahrouge	Dahrouge
18.1	Capital Cost Estimate	Dahrouge	Dahrouge
18.1.1	Basis of Estimate	Dahrouge	Dahrouge
18.1.1.1	Mining, Process, and Infrastructure Capital Costs	Dahrouge	Dahrouge
18.1.1.2	Tailings and Tailings Water Management Capital Costs	Dahrouge	Dahrouge
18.2	Capital Cost Summary	Dahrouge	Dahrouge
18.2.1	Capitalized Pre-production Costs	Dahrouge	Dahrouge
18.2.2	Mining Capital Costs	Dahrouge	Dahrouge
18.2.3	Processing Plant Capital Costs	Dahrouge	Dahrouge
18.2.3.1	Processing Indirects	Dahrouge	Dahrouge
18.2.3.2	Process Commissioning	Dahrouge	Dahrouge
18.2.4	Tailings Water Management and Salt Management Cells	Dahrouge	Dahrouge
18.2.4.1	Temporary Waste Rock Storage Facility	Dahrouge	Dahrouge
18.2.5	Water Management and Infrastructure	Dahrouge	Dahrouge
18.2.6	Site Preparation and Infrastructure Capital Costs	Dahrouge	Dahrouge
18.2.6.1	Site Wide Indirects	Dahrouge	Dahrouge
18.2.7	Owner’s Costs	Dahrouge	Dahrouge
18.2.8	Closure and Reclamation	Dahrouge	Dahrouge
18.2.9	Sustaining Capital Costs	Dahrouge	Dahrouge
18.2.10	Contingency	Dahrouge	Dahrouge
18.3	Operating Cost Estimate	Dahrouge	Dahrouge
18.3.1	Basis of Estimate	Dahrouge	Dahrouge

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S-K 1300 Technical Report Summary

18.3.1.1	Mining Operating Costs	Dahrouge	Dahrouge
18.3.1.2	Process Plants Operating Costs	Dahrouge	Dahrouge
18.3.1.3	Tailings and Tailings Water Management Operating Costs	Dahrouge	Dahrouge
18.3.1.4	Site G&A Operating Costs	Dahrouge	Dahrouge
18.3.1.5	Owner’s Costs Capital Costs	Dahrouge	Dahrouge
18.3.1.6	Water Supply Operating Costs	Dahrouge	Dahrouge
18.3.1.7	Closure and Reclamation	Dahrouge	Dahrouge
18.3.2	Operating Cost Summary	Dahrouge	Dahrouge
18.3.2.1	Mining Operating Costs	Dahrouge	Dahrouge
18.3.2.2	Process Plant Operating Costs	Dahrouge	Dahrouge
18.3.2.3	Tailings, Salt and Tailings Water Management Operating Costs	Dahrouge	Dahrouge
18.3.2.4	Site G&A Operating Costs	Dahrouge	Dahrouge
19	ECONOMIC ANALYSIS	Dahrouge	Dahrouge
19.1	Methodology Used	Dahrouge	Dahrouge
19.2	Financial Model Parameters and Assumptions	Dahrouge	Dahrouge
19.2.1	Physicals	Dahrouge	Dahrouge
19.2.2	Revenue	Dahrouge	Dahrouge
19.2.3	Operating Costs	Dahrouge	Dahrouge
19.2.4	Capital Costs	Dahrouge	Dahrouge
19.3	Cashflow Forecasts and Annual Production Forecasts	Dahrouge	Dahrouge
19.4	Sensitivity Analysis	Dahrouge	Dahrouge
20	ADJACENT PROPERTIES	Dahrouge	Dahrouge
21	OTHER RELEVANT DATA AND INFORMATION	Dahrouge	Dahrouge
21.1	Project Implementation Plan	Dahrouge	Dahrouge
21.1.1	Project Cost Objectives	Dahrouge	Dahrouge
21.1.2	Project Schedule Objectives	Dahrouge	Dahrouge
21.1.3	Early Works	Dahrouge	Dahrouge
21.1.4	Project Team	Dahrouge	Dahrouge

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S-K 1300 Technical Report Summary

21.1.5	Project and Document Control	Dahrouge	Dahrouge
21.1.6	Engineering	Dahrouge	Dahrouge
21.1.7	Supply Chain and Procurement	Dahrouge	Dahrouge
21.1.8	Construction Management	Dahrouge	Dahrouge
21.1.9	Commissioning, Operational Readiness, and Early Operations	Dahrouge	Dahrouge
22	INTERPRETATIONS AND CONCLUSIONS
22.1	Introduction	Dahrouge	Dahrouge
22.2	Geology & Mineral Resource	Understood	Understood
22.3	Exploration, Drilling and Analytical Data Collection in Support of Mineral Resource Estimation	Dahrouge	Dahrouge
22.4	Processing and Metallurgical Testing	L3	L3
22.5	Mining & Mineral Reserve	Optimize Group	Optimize Group
22.6	Recovery Methods	L3	L3
22.7	Infrastructure	Tetra Tech	Tetra Tech
22.7.1	Tailings	NioCorp	Scott Honan
22.8	Environmental, Permitting and Social or Community Considerations	Olsson	Olsson
22.9	Market Studies and Contracts	Dahrouge	Dahrouge
22.1	Capital and Operating Costs	Dahrouge	Dahrouge
22.11	Economic Analysis	Dahrouge	Dahrouge
22.12	Opportunities and Risk Assessment	Dahrouge	Dahrouge
22.12.1	Opportunities	Dahrouge	Dahrouge
22.12.2	Risks	Dahrouge	Dahrouge
23	RECOMMENDATIONS
23.1	Recommended Work Programs
23.1.1	Geology and Resources	Dahrouge	Dahrouge
23.1.1.1	Quality Assurance/Quality Control	Dahrouge	Dahrouge
23.1.2	Hydrometallurgical Plant	L3	L3
23.1.3	Geotechnical	A2GC	A2GC
23.1.4	Mining and Reserves	Optimize Group	Optimize Group

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S-K 1300 Technical Report Summary

23.1.5	Recovery Methods	L3	L3
23.1.6	Infrastructure	Tetra Tech	Tetra Tech
23.1.7	Environmental and Social	Olsson	Olsson
23.1.8	Hoisting Plants	Cementation	Martin LePage
23.1.9	Summary of Costs for Recommended Work	Dahrouge	Dahrouge
24	REFERENCES
24.1	References
24.2	Glossary
24.2.1	Mineral Resource
24.2.2	Mineral Reserve
24.2.3	Definition of Terms
24.2.4	Abbreviations
25	RELIANCE ON INFORMATION PROVIDED BY THE REGISTRANT
26	DATE AND SIGNATURE PAGE
27	QP RESPONSIBILITY MATRIX

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