The mineralization at the Arctic Deposit and at several other known occurrences within the Ambler Sequence stratigraphy of the Ambler District, consists of Devonian age, polymetallic (zinc-copper-lead-silver-gold) VMS-like occurrences. VMS deposits are formed by and associated with sub-marine volcanic related hydrothermal events. These events are related to spreading centres such as fore arc, back arc or mid-ocean ridges. VMS deposits are often stratiform accumulations of sulphide minerals that precipitate from hydrothermal fluids on or below the seafloor. These deposits are found in association with volcanic, volcaniclastic and/or siliciclastic rocks. They are classified by their depositional environment and associated proportions of mafic and/or felsic igneous rocks to sedimentary rocks.

In the Ambler District, VMS-like mineralization occurs in the Ambler Sequence schists over a strike length of approximately 110 km. These deposits are hosted in volcaniclastic, siliciclastic and calcareous metasedimentary rocks interlayered with mafic and felsic metavolcanic rocks. Sulphide mineralization occurs above the mafic metavolcanic rocks but below the Button schist, a distinctive district wide felsic unit characterized by large albite porphyroblasts after relic phenocrysts. The presence of the mafic and felsic metavolcanic units is used as evidence to suggest formation in a riftrelated environment, possibly proximal to a continental margin.

A sulphide-smoker occurrence has been tentatively identified near Dead Creek northwest of the Arctic Deposit and suggests local hydrothermal venting during deposition. However, the lack of stockworks and stringer type mineralization at the Arctic Deposit suggest that the deposit may not be a proximal vent type VMS. Although the deposit is stratiform in nature, it exhibits characteristics and textures common to replacement-style mineralization. At least some of the mineralization may have formed as a diagenetic replacement.

At Arctic, sulphides occur as semi-massive (30 to 50% sulphide) to massive (greater than 50% sulphide) layers, typically dominated by pyrite with substantial disseminated sphalerite and chalcopyrite and trace amounts of galena and tetrahedrite. The Arctic Deposit sulphide accumulation is thought to be stratigraphically correlative to those seen at the Dead Creek and Sunshine deposits up to 12 km to the west.

There is also an occurrence of epithermal discordant vein and fracture hosted base metal (lead-zinc-copper) mineralization with significant fluorite mineralization identified at the Red prospect in the Kogoluktuk Valley, east of the Arctic Deposit. Although not yet fully understood, the genesis of this occurrence is considered to be related to the regional system that formed the VMS deposits in the Ambler District.

Mineralization occurs as stratiform SMS to MS beds within primarily graphitic chlorite schists and fine-grained quartz sandstones. 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 bulk of the mineralization is within five modelled zones lying along the upper and lower limbs of the Arctic isoclinal anticline. All of the zones are 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 resources. The other three zones also occur at specific pseudo-stratigraphic levels, but are too discontinuous to confidently model as resources.

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

Mineralization is predominately coarse-grained sulphides consisting mainly of chalcopyrite, sphalerite, galena, tetrahedrite, arsenopyrite, pyrite and pyrrhotite. Trace amounts of electrum and enargite are also present. Gangue minerals associated with the mineralized horizons include quartz, barite, white mica, black chlorite, talc, calcite, dolomite and cymrite.

The mine will utilize a conventional truck-and shovel fleet and the proposed open pit operations will provide process plant feed at a nominal rate of 10,000 t/d. The pit design incorporates a bench height of 5 m and an inter-ramp angle of 45°. Pit optimization and design results are based on scoping level geotechnical and operational information.

Due to operational and geometric constraints, the mine will require a ramp-up period of three years before achieving full mill capacity. The mining schedule does not currently consider a low-grade stockpiling option but this can be assessed in more detail in future studies.

A 10,000 t/d process plant has been designed to process the massive and semi-massive sulphide mineralization of the Arctic Property. The main economic elements found in the deposit are copper, zinc, lead, and associated gold and silver. The process plant will operate two shifts per day, 365 d/a with an overall plant availability of 92%. The process plant will produce three concentrates: 1) copper concentrate, 2) zinc concentrate, and 3) lead concentrate. Gold and silver are expected to be payable at a smelter and are recovered in both the copper and lead concentrates. The process plant feed will be supplied from the Arctic open pit mine.

The crushed material will be ground by two stages of grinding, consisting of one SAG mill and one ball mill in closed circuit with hydrocyclones (SAB circuit). The hydrocyclone overflow with a grind size of approximately 80% passing 70 µm will first undergo pre-talc flotation, and then be processed by conventional bulk flotation (to recover copper, lead, and associated gold and silver), followed by zinc flotation. The rougher bulk concentrate will be cleaned and followed by copper and lead separation to produce a lead
concentrate and a copper concentrate. The final tailings from the zinc flotation circuit will be pumped to the TSF. Copper, lead, and zinc concentrates will be thickened and pressure-filtered before being transported by truck to a port and shipped to smelters.

Run-of-mine (ROM) material will be trucked to a 200 t receiving dump bin equipped with a stationary grizzly screen with an opening of 1,000 mm. The undersize of the screen will be fed to a primary jaw crusher driven by a 250 kW motor . The nominal crushing rate designed is 641 t/h. A rock breaker will be provided to break oversize rocks retained on the screen.

ROM material will be crushed by the jaw crusher to 80% passing 125 mm. The crusher product will be conveyed to a 10,000 t live capacity stockpile.

A dust collection system will be provided to control fugitive dust generated during crushing and transport of the materials. A magnet will be provided over the crusher discharge conveyor to remove any steel pieces, which will protect the downstream conveyor.

The primary grinding circuit will consist of one 7.3 m diameter by 3.8 m long SAG mill equipped with a trommel screen. The SAG mill will be driven by one 3,000 kW motor. Crushed material will be conveyed from the stockpile to the SAG mill feed chute and then to the SAG mill. The SAG mill will be equipped with 50 mm pebble ports for the removal of undersize material. Mill discharge from the SAG mill will be screened by the SAG mill trommel. The trommel undersize material will flow by gravity to a sump where the ball mill discharge will report to as well. The blended slurry will be pumped to the hydrocyclone cluster. Oversized material from the trommel will be sent to a surge bin and then trucked to the TSF for disposal. As required, steel balls will be added into the SAG mill to maintain grinding efficiency. The estimated primary grind size is 80% passing 800 to 900 µm.

The secondary grinding circuit will include one ball mill in closed circuit with one hydrocyclone cluster. The SAG mill product will be pumped to the hydrocyclone cluster along with the ball mill discharge. The hydrocyclone underflow will gravity-flow to the ball mill while the hydrocyclone overflow, with a solid density of 33% w/w, will be sent downstream to the copper and lead bulk flotation circuit. The proposed circulation load for the closed circuit is 300%.

The hydrocyclone overflow slurry will be processed by conventional flotation to recover the target minerals. Flotation will consist of talc pre-flotation, copper-lead bulk flotation to produce a bulk copper-lead concentrate, and subsequent copper and lead separation flotation to produce a copper concentrate and a lead concentrate. The copper-lead bulk flotation tailings will be conditioned and floated to produce a zinc concentrate. The zinc flotation tailings, which are the final tailings product, will be pumped to the TSF for storage.

The hydrocyclone overflow with 33% solids will be floated in two 100 m3 tank-type flotation cells to remove talc, which may impact copper and lead flotation and concentrate quality. Depending on the copper, lead, and zinc content of the talc product, the talc rougher flotation concentrate may be further cleaned to reject any entrained sulphide minerals. The pre flotation tailings will be directed to the copper-lead bulk flotation circuit, while the talc concentrate will be sent to the final tailings pump box to be pumped to the TSF, together with the zinc rougher scavenger flotation tailings and the first zinc cleaner scavenger flotation tailings.

The talc pre-flotation tailings will flow into the copper-lead bulk flotation conditioning tank where copper-lead mineral collectors will be added. The conditioned slurry will be floated in six conventional 100-m3 tank-type flotation cells, five cells for rougher flotation and one for rougher/scavenger flotation. The concentrate from the rougher bulk flotation circuits will be reground in a tower mill, which is driven by one 750 kW motor. The mill will be in closed circuit with ten 250 mm diameter hydrocyclones to grind the bulk concentrate to a particle size of 80% passing 35 to 40 µm prior to being further upgraded by two stages of cleaner flotation.

The first stage of copper-lead bulk cleaner flotation will be conducted in one 4.0 m diameter by 11.0 m high flotation column. The first cleaner flotation tailings will be scavenged to further recover copper-lead minerals in two 20 m3 tank-type flotation cells. The concentrate from the first cleaner/scavenger flotation will flow to the regrinding pump box for further regrinding, while the tailings produced from the flotation stage will be sent to the copper-lead rougher/scavenger tailings pump box where the combined slurry will be pumped to the zinc flotation circuit.

The copper-lead concentrate from the first cleaner flotation stage will be pumped to the second cleaner flotation stage. Secondary cleaner flotation will be conducted in one 4.0 m diameter by 10.0 m high flotation column. The concentrate from the copper-lead second cleaner flotation will be pumped to the copper lead bulk concentrate thickener prior to being pumped to the copper and lead separation flotation circuit. The overflow of the thickener will be sent to the process water tank for reuse. The tailings from the copper lead secondary cleaner flotation will return to the copper-lead first cleaner flotation.

The underflow of the copper-lead bulk cleaner concentrate thickener will be pumped to the copper-lead separation flotation conditioning tanks, where sodium cyanide and lime will be added to suppress copper minerals. The conditioned slurry will flow to lead rougher flotation cells. Lead collector will be added in the second conditioning tank to collect lead minerals. The lead concentrate produced from the lead rougher and rougher/scavenger flotation cells will be pumped to the first lead cleaner flotation circuit. The tailings of the rougher/scavenger flotation, which is the copper concentrate, will report to the copper concentrate thickener.

The lead concentrates from the lead rougher and scavenger flotation circuit will be further upgraded by three stages of cleaner flotation. The first stage of cleaner flotation will consist of first lead cleaner flotation and cleaner scavenger flotation. The lead concentrates from the first stage of cleaner flotation will be fed to the second stage of lead cleaner flotation. The lead cleaner/scavenger tailings, together with the lead rougher/scavenger concentrate, will report to lead rougher flotation. The second lead cleaner flotation concentrate will be further upgraded by the third stage of cleaner flotation to produce the final lead concentrate, which will report to the lead concentrate thick.