Overview

Deposit type

• VMS

Summary:

Previous workers at the Arctic deposit (Russell 1995 and Schmidt 1983) describe three mineralized horizons: the Main Sulphide Horizon, the Upper South Horizon and the Warm Springs Horizon. The Main Sulphide Horizon was further subdivided into three zones: the southeast zone, the central zone, and the northwest zone. Previous deposit modelling was grade-based resulting in numerous individual mineralized zones representing relatively thin sulphide horizons.

Observations and interpretations at the Arctic deposit such as: 1) the tectonic setting with Devonian volcanism in an evolving continental rift; 2) the geologic setting with bimodal volcanic rocks including pillow basalts and felsic volcanic tuffs; 3) an alteration assemblage with well-defined magnesium-rich footwall alteration and sodium-rich hanging wall alteration; and 4) typical polymetallic base-metal mineralization with massive and semi-massive sulphides, are indicative of a Volcanogenic Massive Sulphide (VMS) deposit that has undergone high strain and complex folding and faulting.

The tonnage, grades, and stratigraphic setting of the Arctic deposit, and its broader tectonostratigraphic setting, are similar to other felsic siliclastic VMS environments globally. The deposit has strong similarities to deposits found the Finlayson Lake VMS district, Yukon, Bathurst district, New Brunswick, and some parts of the Iberian Pyrite Belt, Spain-Portugal (Piercey, 2022).

A VMS model is considered applicable for use in exploration targeting in the project area.

Recent work by Ambler Metals defines the Arctic deposit as two or more discrete horizons of sulphide mineralization contained in a complexly deformed isoclinal fold with an upright upper limb and an overturned lower limb hosting the main mineralization. Nearby drilling suggests that a third upright lower limb, likely occurs beneath the explored stratigraphy.

Arctic Deposit Mineralization
Mineralization occurs as stratiform semi-massive sulphide (SMS) to massive sulphide (MS) beds within primarily graphitic schists and fine-grained quartz mica schists. The sulphide beds average 4 m in thickness but vary from less than 1 m up to as much as 18 m in thickness. The sulphide mineralization occurs within eight modelled zones lying along the upper and lower limbs of the Arctic isoclinal anticline. The zones are all within an area of roughly 1 km2 with mineralization extending to a depth of approximately 250 m below the surface. There are five zones of MS and SMS that occur at specific pseudo-stratigraphic levels which make up the bulk of the Mineral Resource estimate. The other three zones also occur at specific pseudo-stratigraphic levels but are too discontinuous.

Unlike more typical VMS deposits, mineralization is not characterized by steep metal zonation or massive pyritic zones. Mineralization dominantly consists of sheet-like zones of base metal sulphides with variable pyrite and only minor zonation, usually on a small scale.

Mineralization is predominately coarse-grained sulphides comprising chalcopyrite, sphalerite, galena, tetrahedrite-tennantite, pyrite, arsenopyrite, and pyrrhotite. Sulphides occur as disseminated (sulphide) to massive (greater than 50% sulphide) layers. Trace amounts of electrum are also present. Gangue minerals associated with the mineralized horizons include quartz, barite, white mica, chlorite, stilpnomelane, talc, calcite, dolomite and cymrite.

Deposit Model
The mineralization at the Arctic deposit and at several other known occurrences within the Ambler Sequence stratigraphy of the Ambler Mining District consists of Devonian age, polymetallic (zinc-copper-lead-silver-gold) VMS-like occurrences.

Observations and interpretations at the Arctic deposit such as: 1) the tectonic setting with Devonian volcanism in an evolving continental rift; 2) the geologic setting with bimodal volcanic rocks including pillow basalts and felsic volcanic tuffs; 3) an alteration assemblage with well-defined magnesium-rich footwall alteration and sodium-rich hanging wall alteration; and 4) typical polymetallic base-metal mineralization with massive and semi-massive sulphides, are indicative of a Volcanogenic Massive Sulphide (VMS) deposit that has undergone high strain and complex folding and faulting.

Mining Methods

• Truck & Shovel / Loader

Summary:

The Arctic Project is designed as a conventional truck-shovel operation with 144 t trucks and 15 m3 shovels. The pit design includes four nested phases to balance stripping requirements while satisfying the concentrator requirements. The design parameters include a ramp width of 30 m, road grades of 10%, bench height of 5 m, targeted mining width between 70 and 100 m, berm interval of 20 m, variable slope angles by sector and a minimum mining width of 30 m.

According to SRK’s geotechnical recommendations, a large talc zone located in the northeast pit slope must be completely removed. To achieve this, a pit optimization giving dummy values to the talc zone blocks was performed. This pit optimization was used as the basis for the mine design.

The smoothed final pit design contains approximately 46.7 Mt of ore and 340.2 Mt of waste for a resulting stripping ratio of 7.3:1. Within the 46.7 Mt of ore, the average grades are 2.11% Cu, 2.90% Zn, 0.56% Pb, 0.42 g/t Au and 31.83 g/t Ag.

Interim Phase Design
The deposit is mined in four nested phases, including the ultimate pit limit. During pre-production, the phasing strategy maximizes waste mining while minimizing the ore tonnage. This is primarily due to the lack of space to build ore stockpiles. In addition, the phasing strategy maintains enough exposed ore to guarantee the continuous operation of the process plant during the production.

A stockpile is required to store the ore mined during the pre-production period. Because this stockpile is depleted at the beginning of the operation, it is located within the WRF footprint and has a total storage capacity of 223,000 m3 . This volume is enough to satisfy the maximum stockpiling capacity of approximately 776 kt.

Production Schedule
The production schedule includes the processing ramp up. The processing plant ramp-up considers the normal inefficiencies related to the start of operations, and includes the tonnage processed as well as the associated recoveries, which increases the design capacity during the second quarter of operation. The mine requires two years of preproduction before the start of operations in the processing plant.

The deposit is mined in four nested phases, including the ultimate pit limit. The schedule was developed in months for the pre-production period and the first two years of production, in quarters from Year 3 to Year 5 of production, and annually thereafter. The scheduling constraints set the maximum mining capacity at 35 Mt per year and the maximum number of benches mined per year at ten in each phase.

The production schedule based on the Probable Mineral Reserves shows a LOM of 15 years, including two years of preproduction. The amount of rehandled mill feed is 776 kt, which is the ore mined during the pre-production period. The average grades to the mill over the LOM are 2.11% Cu, 2.90% Zn, 0.56% Pb, 0.42 g/t Au, 31.83 g/t Ag and 2.92% Talc.

Waste Material Handling
Waste is hauled to the WRF using 144 t trucks. The construction sequence starts at the bottom of the dump by dumping the material in 5-m lifts, leaving a 23.5-m bench every four lifts. The resulting overall slope angle of the dump face is 2.5H:1V.

Mining Equipment
A conventional owner-operated truck fleet utilizing a combination of hydraulic shovels and front-end loaders (FELs) is planned. The truck fleet consists of large trucks for waste stripping and for mining the ore zones. The trucks are diesel powered with a combined capacity to mine a maximum of 40Mt/a operating on a combination of 5 m and 10 m benches. The loading fleet is also diesel powered. Blasting is contract performed with explosives manufactured off-site and delivery to site by trucks.

Blasting
Blasting operations are contracted to a blasting explosives provider who is responsible for the blast design, loading, stemming, and initiation. Bulk emulsion is manufactured off-site. The explosive provider transports explosive material from Fairbanks, Alaska. The explosive is stored in four explosive silos with 80-ton capacity each: two silos for storing Ammonium Nitrate and the other two for storing emulsion produced on site.

In performing the explosive services, the blasting contractor is proposing to provide: • Two five person blasting crews providing seven day a week coverage; • Transportation of the explosive from Fairbanks to site.

Additionally, two (2) Mobile Manufacturing Unit (MMU) trucks and one stemming truck are provided by the mine. The MMU explosive trucks delivers a bulk emulsion product when required.

Two types of Heavy ANFO blend (HA) are used: 70% emulsion/30% Ammonium Nitrate Fuel Oil (ANFO) for wet material, and 30% emulsion/70% ANFO for dry material, with specific gravity of 1.25 g/cm3 and 1.23 g/cm3, respectively.

The wall control design patters are based on the 2022 Arctic Drill and Blast Study recommendations There are three main types of geotechnical zones within the pit: one group in the North and East walls, where the slope design is governed by the shallower foliation, with IRA lower than 38° and with geotechnical berms every 60 m, other group with IRA of between 38° and 47°, and other group with IRA greater than 47°. Pre-split blasting only used for the third group with IRA greater than 47°. For other zones, multiple stab trim holes are used for face outlining.

Although material is mined on a combination of 5 and 10 m benches, all material is drilled and blasted on 10 m benches.

The powder factor for ore is 0.3 kg/t and 0.26 kg/t for waste. For trim, variable powder factors were used depending on the characteristics of the geotechnical zone.

Drilling
Throughout the project life, drilling is required for both ore control and production blasting. Rock fragmentation achieved through blasting is the overriding design criteria for the drill hole pattern design.

Drill penetration is a function of bit size, bit load, drilling method, and rock strength properties. SRK completed unconfined compressive testing (UCS) and point load testing on Arctic’s primary rock types. The weighted average UCS value for the ore hosting rocks is 83 MPa and 59 MPa for the waste areas.

Loading
From the trade-off study, the primary loading units selected are two 15 m3 hydraulic shovels. After Year 11 only one hydraulic shovel is required. To assist the hydraulic shovels, two 12 m3 front end loaders (FEL) are scheduled until Year 7 of production, dropping to one until the end of the LOM. The FEL is also used for stockpile rehandling, all of which is scheduled in Year 1.

Hauling
From the TCO analysis, the primary hauling unit selected for ore and waste mining is a mechanical drive truck with a payload capacity of 144 t wet, assuming a standard body with a full set of liners.

Fourteen trucks are commissioned during pre-production. During Year 1 the fleet is ramped up to 15, reaching its peak. After Year 9, truck requirements drop to 14. Following Year 10, the 144 t truck requirements decline with declining mining rates.

Support equipment includes excavators, track dozers, rubber-tired dozers (RTDs), sand trucks, graders, water trucks, fuel/lube trucks, and water trucks.

Comminution

Crushers and Mills

Summary:

Run-of-mine ore will be trucked to the crushing station and dumped into a 200-t receiving bin protected with a 1,000-mmaperture stationary grizzly. Ore will be reclaimed from the bin with an apron feeder and scalped of fines with a 75-mmaperture vibrating grizzly. The grizzly oversize will be passed to the primary jaw crusher. The selected conventional jaw crusher will operate with a closed side setting of 100 mm. The crushed material together with the grizzly undersize, with a P80 of 80 mm, will be discharged to the conveyor. A tramp metal magnet will be installed at the head end of the conveyor to remove tramp metal.

Crushed ore will be sent to a single covered conical stockpile via the stockpile feed conveyor.

The major equipment in the crushing circuit will include:
• One apron feeder, installed power 30 kW;
• One vibrating grizzly, installed power 30kW; and
• One single toggle jaw crusher, installed power 160 kW.

Coarse Ore Storage
The coarse ore stockpile will have a live capacity of 5,000 t, equivalent to approximately 12 h of mill feed at the nominal mill feed rate. The stockpile total capacity will be approximately 20,000 t. The coarse ore stockpile will be a covered facility to mitigate freezing of the stockpiled material, which will be also equipped with a dust collecting system. The coarse ore stockpile building will have sufficient space to allow for the operation of mobile equipment as required. Reclaim of ore from the stockpile will be accomplished using two 1,000-mm-wide by 4,800-mm-long apron feeders at a nominal rate of 226 t/h per feeder. Reclaimed material from the apron feeders will be discharged onto the SAG mill feed conveyor.

The major equipment of the coarse ore storage area will include:
• Two apron feeders, unit installed power 11 kW.

Primary Grinding and Classification
The grinding circuit will consist of a SAG mill followed by a ball mill arranged in a closed circuit with a hydrocyclone cluster. The nominal feed throughput of the circuit will be approximately 453 t/h. The SAB circuit will reduce the crushed ore particle size from a P80 of 80 mm to 70 µm.

SAB circuit is selected by considering the low competence of the ore materials. The SAG mill will be a grate discharge type with 12 mm apertures and no pebble ports. As required, steel balls will be added to the SAG mill to maintain mill power. The SAG mill product will discharge onto a trommel screen. Trommel screen undersize will report to a hydrocyclone feed pumpbox and the oversize to a scats bunker.

The ball mill discharge will report to the hydrocyclone feed pumpbox, where it will be combined with the SAG mill trommel undersize prior to feeding the hydrocyclone cluster. Process water will be added to the SAG mill feed chute and hydrocyclone feed pumpbox to maintain a target slurry density. The hydrocyclone underflow will feed the ball mill by gravity, while the hydrocyclone overflow, with a solids content of 37%, will gravitate to the flotation circuit. The ball mill circulating load will be 350%.

The major equipment of the grinding and classification circuit will include:
• One SAG mill, 6.1 m in diameter (20 ft) by 4.9 m (16 ft) EGL, installed power 3,000 kW;
• One ball mill, 6.1 m diameter (20 ft) by 9.1 m (30 ft) EGL, installed power 6,000 kW;
• Two 55 kW slurry pumps to pump hydrocyclone feed, with one pump in operation and one in standby; and
• One hydrocyclone cluster with fourteen 400 mm hydrocyclones, nine in operation, four in standby and one blank.

Regrinding circuits will include:
One regrind ball mill, 3.0 m diameter by 5.3 m EGL, installed power 600 kW;
One regrind ball mill, 2.7 m diameter by 4.1 m EGL, installed power 355 kW;

Processing

• Jameson Cell Flotation
• Crush & Screen plant
• Flotation
• Dewatering
• Filter press
• Cyanide (reagent)

Summary:

The process plant is designed for a throughput of 10,000 t/d, equivalent to 3,650,000 t/a.

The mined material will be processed via a primary crushing circuit and a grinding circuit configured in semi-autogenous mill and ball mill (SAB) configuration. Due to the significant levels of talc contained in the mined materials (2.92% in the plant feed), a talc flotation is required to remove the mineral. This is then followed by a bulk flotation of a copper and lead concentrate, and the subsequent separation of copper and lead concentrate. Bulk flotation tails are processed through zinc flotation to produce the three flotation concentrates. The talc flotation concentrate and the zinc tailings will be combined and thickened before discharged to the tailings storage facility.

Flotation
The flotation plant will produce a talc concentrate for disposal, as well as three payable base metal concentrates for shipment to market and sale. The talc concentrate will be produced prior to base metal flotation to minimize dilution of the base metal concentrates. A standard bulk copper and lead concentrate followed by the flotation of a zinc concentrate will be used in the flotation circuit. The copper and lead in the bulk concentrate will be separated by floating the lead from the bulk concentrate to produce individual copper and lead concentrates.

Talc Pre-Flotation
The ball mill hydrocyclone overflow will feed the talc rougher circuit. The talc flotation will be composed of rougher and cleaner flotation stages to float the talc mineral and reject any entrained payable sulphide minerals. The rougher flotation will be performed in the conventional forced-air cells and the cleaner flotation will be performed in a Jameson cell. The talc cleaner concentrate will be pumped to the final tailings.

Copper – Lead Bulk Flotation and Regrinding
The talc pre-flotation tailings will be pumped to the copper/lead bulk flotation conditioning tanks where copper and lead mineral collectors will be added. The conditioned slurry will undergo rougher flotation in conventional tank flotation cells for recovery of a bulk copper and lead concentrate.

The bulk rougher concentrate will be reground in a 600 kW regrind ball mill configured in reverse, closed circuit with hydrocyclones. The bulk rougher concentrate will be reduced to a particle size of 80% passing 40 µm prior to being further upgraded by two stages of cleaner flotation. The 1st bulk cleaner and cleaner scavenger flotation will be conducted in conventional tank cells.

The 1st cleaner flotation tailings will gravitate for scavenging of residual copper and lead minerals. Concentrate from the cleaner scavenger flotation will be pumped to the regrind hydrocyclone feed pumpbox for further regrinding, while the cleaner-scavenger tailings will join the rougher flotation tailings for pumped transfer to the zinc flotation conditioners. The 2nd bulk cleaner flotation will be performed in a Jameson cell to upgrade the concentrate obtained from the last stage. The 2nd cleaner tailings will be reprocessed in the 1st cleaner flotation stage.

Copper-lead bulk cleaner flotation dilution water and launder sprays will be supplied by a dedicated process-water system. The supply for this system will be a combination of copper/lead bulk thickener overflow and main plant process water.

Copper and Lead Separation Flotation Circuit
The bulk 2nd cleaner concentrate will be pumped to the copper-lead separation flotation conditioning tank, where sodium cyanide and lime will be added to supress copper minerals. Lead flotation will be conducted at a pH ranging from 9.0 to 9.5. The conditioned slurry will flow to the conventional lead rougher flotation cells. Frother will be added in the rougher and cleaner flotation cells and lead promoter will be added in the lead rougher concentrate pumpbox to collect lead minerals.

The lead rougher concentrate will be further upgraded in three stages of cleaner flotation to produce the final lead concentrate that will report to the lead concentrate thickener. Tailings from the 2nd and 3rd stages of cleaner flotation will gravity flow to the proceeding flotation cell feed boxes. Tailings from the 1st cleaner flotation will be combined with the tailings from the lead rougher flotation to produce the final copper concentrate, which will report to the copper concentrate thickener.

Zinc Flotation and Regrind
The copper-lead bulk rougher tailings and copper-lead bulk cleaner scavenger tailings will be processed to recover zinc in the zinc flotation circuit. The circuit will consist of feed slurry conditioning, rougher/scavenger flotation, zinc rougher concentrate regrind, and two stages of cleaner flotation. A cleaner scalper stage will be installed when the plant feed zinc grade are expected to be higher.

Tailings from the copper-lead flotation circuit will be conditioned with lime (to a pH above 10.5, to depress pyrite) and copper sulphate (to activate zinc minerals), and then sent to the head cell of a bank of zinc rougher flotation cells. Zinc rougher concentrate will report to the regrinding circuit; zinc rougher tailings will be combined with the Talc concentrate and pumped to the tailings thickener. Thickened tails will be pumped to the Tailings Storage Facility (TSF).

The zinc rougher and cleaner scavenger concentrates will be reground in a 355 kW regrind ball mill configured in, closed circuit with hydrocyclones. The zinc concentrates will be reduced to a particle size of 80% passing 40 µm prior to being further processed by the zinc 1st cleaner flotation circuit.

The zinc cleaner circuit will operate at a pH of 11 or above to reject pyrite. The 1st cleaner concentrate will be further upgraded in the 2nd stage of cleaner flotation to produce the final zinc concentrate. The 1st cleaner flotation tailings will be processed in the cleaner scavenger flotation cells. Concentrate from the cleaner scavenger flotation will return to the zinc regrind circuit. The cleaner scavenger tailings will report to the final tailings pump box, joined by the zinc rougher tailings and talc flotation cleaner concentrate.

Product Dewatering
The copper, lead, and zinc concentrates will report to separate higher-rate thickeners, where flocculant will be added to assist in the settling of the solids. The thickened slurries will be dewatered by using pressure filters to a design moisture content of 6%. Two identical tower press filters are recommended for the concentrate filtration duties, including one common unit for lead and zinc concentrates dewatering; while another unit is dedicated to dewater the copper concentrate.

The main equipment in the copper concentrate dewatering circuit will include:
• One 15-m-diameter high-rate thickener, installed power 3.7 kW;
• One 84 m2 copper concentrate tower filter press, installed power 105 kW; and
• Two 44 m3 cyanide destruction tanks.

The main equipment in the lead concentrate dewatering circuit will include:
• One 6-m-diameter high-rate thickener, installed power 1.5 kW; and
• One 84 m2 tower filter press, shared with zinc concentrate filtration, installed power 105 kW.

The main equipment in the zinc concentrate dewatering circuit will include:
• One 14-m-diameter high-rate thickener, installed power 5.5 kW; and
• One 84 m2 tower filter press, shared with lead concentrate filtration, installed power 105 kW.

Tailings Disposal
The final flotation tailings consist of the talc flotation concentrate, zinc rougher flotation tailings and zinc cleaner scavenger flotation tailings. The final flotation tailings will be directed to a 32 m diameter high-rate thickener. Flocculant will be added to improve settling of the tailings. Thickener underflow with a solid density of 47% will be pumped towards the tailings storage facility. Thickener overflow will be delivered to the Process Pond.

Pipelines and Water Supply

Summary:

Water Supply
There will be three separate water supply systems: a fresh water system, a WRCP water supply system, and a process water supply system. The contact water will originate principally from the Waste-Rock Control Pond. The total water usage for the process plant is 42,240 m3/d.

Fresh Water Supply
System Fresh water will be used for fire water and potable water applications that will be supplied from ground water wells. A 350 m3 fresh water/fire water storage tank will hold well water, with a live capacity of 4 hours.

WRCP Water Supply System
WRCP water will be fed to the TMF during winter months and treated and discharged in the summer. Additional process water will be supplied from the WRCP to the TMF through the winter months to meet mill water demands and provide water stored in the TMF as a contingency for dry years. Contact water will also be supplied form the WRCP, which will feed the potable water and gland water tanks.

Process Water Supply System
Process water will mainly come from the Process Pond which is fed by tailings thickener overflow and reclaimed water from TMF. In addition, contact water is also another process water source as makeup water. Process water will be stored in a 14-m-diameter x 14-m-high tank, from where the water will be distributed to the process plant and other service locations. The process water consumption is 40,326 m3/d.

Tailings Delivery and Return Systems
The tailings delivery system will transport slurried tailings from the processing plant to the TMF. This will consist of one three-kilometre pipeline, with two kilometres of this being 500 mm (20 inch) diameter carbon steel rubber lined pipeline and the remaining one kilometre being 600 mm (24 inch) HDPE. This pipeline will transport up to 1,843 m3 /h of tailings to the TMF.

The return water delivery system for recycle water from the TMF has been sized on the basis of 1,308 m3/h of water being pumped from the TMF to the process water pond. This system will consist of a barge pump and a 3 km-long pipeline, run adjacent to the tailings pipeline. The pipeline will consist of 2 kms of 450 mm (18 inch) diameter carbon steel pipeline, with the remaining 1 km being 500 mm (20 inch) HDPE.

Both pipelines will be heat-traced to prevent freezing.

Production

Operational metrics

Personnel

Job Title	Name	Phone	Email	Profile	Ref. Date
Consultant - Infrastructure	Calvin Boese				Jan 20, 2023
Consultant - Mining	Bruce Murphy				Jan 20, 2023
Consultant - Mining & Costs	Piers Wendlandt				Jan 20, 2023
Consultant - Recovery Methods & Costs	Kevin Murray				Jan 20, 2023
Director of Environment & Sustainability	Cal Craig	(907) 317-7383	cal.craig@amblermetals.com		Oct 19, 2024
Site Manager	Jim Male				Oct 10, 2024