The Hope Bay deposits are considered to be typical of orogenic gold deposits. In Canada, orogenic deposits represent the main source of gold and are typically located in the Archean greenstone belts of the Superior and Slave provinces.

GOLD MINERALIZATION AT THE DORIS AREA.
The Doris Vein system is characterized by a series of north–south striking, subvertical, goldbearing, ductile-brittle structures that commonly host wide, stylolitic, ribboned, or bull quartz veins. A total of nine sub-parallel structures have been identified within the Doris system. The most significant from west to east are: the West Valley Wall set, C/2 Stringer zone, Central, Lakeshore, and Island veins. Gold bearing structures have been traced by diamond drilling for over 2.3 km and to a depth of 650 m.

Within the veins, gold is commonly associated with narrow tourmaline chlorite septa oriented parallel to and along the vein margins. Veins are not consistently mineralized along strike. Gold mineralization includes visible and disseminated gold occurring primarily with quartz veins ranging from a few centimetres to approximately 10 m in scale. Visible gold includes coarse leafy free milling grains located along vein margins, tourmaline septa, and wallrock fragments which are commonly associated with pyrite. Gold is also associated with disseminated sulphides at the margins of the quartz veins, or with sulphide clusters within the vein. Occasionally, gold is present within brecciated zones adjacent to the quartz veins. Sulphide mineralization consists of trace to 2% pyrite, trace chalcopyrite, rare sphalerite, and pyrrhotite.

GOLD MINERALIZATION AT MADRID AREA.
The gold mineralization within Naartok West, Naartok East, Rand Suluk, and Patch 7 consists of quartz-carbonate stockwork veining, which overprints dolomite-sericite-albite-pyrite altered mafic volcanic rocks of the Patch Group. The gold mineralization is characterized by multistage brecciation and alteration with at least two separate gold mineralization events. Gold occurs within north-northeast, east, southeast, and north-northwest trending brecciated and carbonate altered zones and is associated with disseminated pyrite which has replaced brecciated mafic fragments.

GOLD MINERALIZATION AT THE BOSTON AREA.
Gold mineralization at the Boston deposit is present in sub-vertical horizons of extensive hydrothermal alteration within a large iron-rich carbonate-altered shear zone. Organic carbon graphite is a common component of the Boston mineralized system. Gold occurs within and around structurally controlled quartz-carbonate veins, which have been developed along lithological contacts. Gold is associated with sulphide mineralization, mainly as clots of the pyrite within veins as well as in the wallrock. Three major zones of the gold mineralization have been identified: B2, B3, and B4 zones. Each zone is over one kilometre in length and is composed of numerous narrow quartz-carbonate veins with disseminated pyrite. The veins are from 0.05 m to 3.0 m wide and of variable lengths within a 10 m to 40 m wide mineralized zone.

The mining methods that will be used for the Project include a bulk method consisting of Sublevel Longhole Retreat (SLR) and a more conventional method of Drift and Fill in narrow and more difficult ground conditions. The use of Transverse Longhole mining was considered initially in the PEA study for the Madrid North Naartok zone, however, a review has indicated that the large widths are made up of several veins with sufficient separation between them to allow for Longitudinal Longhole mining. The sub-level intervals were fixed at 20 m to provide the required control for the small diameter longholes that will be used. The nature of the veins did not appear conducive to use of increased sub-level intervals and larger borehole diameters. The Drift and Fill method will be used for all of the Madrid South (Patch 14 and Wolverine) zones and the Hinge portion of the Doris North zone while the Hinge limbs will be mined by SLR. All of Madrid North (including Naartok, Rand, and Suluk) as well as the Boston zones will be mined using SLR.

The stoping method for the bulk of the mining at Hope Bay will be the SLR method followed by a combination of CRF and UCF made up of the mine waste rock and surface quarry rock.

The undercut drift and overcut drift are driven on ore and mining would start from the lower access retreating out towards the main access drift. Stope strike lengths can vary from 20 m to 35 m. Once a stope block is finished, CRF is introduced from the upper sub-level to create a “plug” and the remainder of the stope is backfilled with unconsolidated backfill (waste rock from development or quarried rock). The CRF plug required will be optimized as experience is gained in each mine and an allowance of on average 40% of CRF backfill tonnage has been utilized in the current study (ranging from 25% to 48%).

To commence the next stope, a slot raise is drilled three to four metres from the CRF plug to serve as a slot for the initial start of blasting. Longhole drilling is done from the top sub-level, however, it can also be carried out from the bottom sub-level if required. Blasting consists of between two and four rings, with 1.5 m burden per ring, followed by mucking. Keeping the stopes fairly short on strike length will permit effective mucking as good visibility can be maintained. The stope strike length is controlled in order to respect the HR values and prevent excessive dilution from the HW and FW. The use of a portion of unconsolidated backfill will result in re mucking of some of this backfill when mucking the stopes above. This stoping system has the advantage of reducing the waste development required in preparation of the stopes, however, this necessitates having more sub-levels opened up to provide sufficient faces to maintain the planned production rate of the mine.

Drill holes will be from 2.5 in. to 4.0 in. in diameter (63.5 mm to 100 mm) averaging approximately 17 m in length given the sub level interval of 20 m floor to floor. At this length, the drill hole deviation is expected to be negligible and breakthrough holes can be located and picked up by survey to confirm accuracy.

Subsequently, the longholes will be loaded with ANFO blasting agent and primed and blasted in a series of three to four rings at a time.

Stoping operations at the Madrid North’s Naartok and the Boston mine will use the SLR method followed by a combination of CRF and UCF backfill. For these veins, however, there are frequent parallel zones encountered thereby increasing the level of CRF that will be required to permit effective mining of stopes located in close proximity. While this will increase the stoping cost with respect to the backfill, it will offer flexibility in the longhole drilling phase of the stoping. CRF and sill mats will also be utilized to create separate mining horizons to permit achieving production tonnages. Small pillars left below the mats will be drilled and blasted on retreat as the last phase of a stope cycle.

In areas of Madrid North and Naartok, the width of the vein is large (greater than eight metres) and RPA recommends assessing the use of larger diameter drill holes and increased drill pattern to lower stoping costs. This needs to be assessed taking into consideration the blasting design and the effect on the stope wall damage (increased peak particle velocity) and potential increased dilution.

The Drift and Fill stoping method will be used at the Doris North mine in the larger part of the Hinge zone. In this case, drifts will be driven on both contacts due to the width of the vein. Subsequently, primary drifts will be driven 4 m high by 4 m wide from the hangingwall contact drift across and break through to the footwall contact drift. The primary drift will be filled with CRF. Secondary drifts will then be driven at 4 m high by 4 m wide on centre and the two metres on either side slashed out on retreat. The secondary drift will then be filled with UCF.

For the subsequent lift, driven above the first lift, the drifts will be angled opposite to create a herring bone effect designed to reduce dilution as drifts are driven across less UCF and not directly over top of the initial drive.

Drift and Fill method will be used at the Madrid South for the Patch 14 and Wolverine veins. The access drift will be offset about 50 m from the vein and an attack ramp driven at -20% gradient to intercept the vein and through to the HW contact. Subsequent cuts will be taken at 4 m heights with a total of five cuts and an approximate two metre backslash taken on retreat under the CRF drift of the upper stope. The veins will be accessed at approximately 150 m centres and the cuts driven in both directions to provide two working faces. This will be repeated on several levels to enable achieving the production levels.

The plant recovers gold from the ore using a combination of continuous gravity concentration, flotation, intensive cyanidation, and resin absorption unit processes to achieve an overall process plant recovery of 91% over the life of mine. In Gekko terminology, the plant is divided into two sections – Python and CTP (Concentrate Treatment Plant).

ROM material from the mine is trucked to the ROM pad. A front end loader recovers the ore from the ROM pad and tips onto the Primary Grizzly to scalp out +600 mm materials. The +600 mm materials are presented to the Rock Breaker to be broken and stockpiled for reclaim, while the -600 mm materials flow onto the vibrating grizzly.

The vibrating grizzly scalps out minus 100 mm ore and feeds the oversize into the primary jaw crusher to be crushed to P95 125 mm. The vibrating grizzly undersize and crusher discharge are conveyed by the primary crusher discharge conveyor under a belt magnet to the stockpile feed conveyor and transferred to the stockpile.

The Python plant feed conveyor ultimately discharges into the secondary jaw crusher via the grizzly feeder, scalping out minus 50 mm material.

The P80 ~1.7 mm undersize slurry from the secondary screen reports to the secondary screen undersize sump; from which it is pumped to the gravity/grinding circuit.

The secondary screen underflow pump delivers slurry to one side of the Primary Mill Discharge sump, where they are mixed with the discharge from the primary mill to the rougher InLine Pressure Jig (IPJ). The concentrate from the rougher IPJ is sent to the cleaner IPJs (IPJ 1000), while the tails report to the other side of the Primary Mill Discharge sump, from where it is combined with excess mill discharge slurry and is pumped to the cyclone cluster. The cyclone clusters are banks of 6 Cavex 250CVX20, with four operating and two standbys. The underflow slurry from the cyclones return to the primary mill for grinding, while the overflow reports to the agitated rougher flotation conditioning tank. 
 
The cleaner IPJ tail returns to the secondary screen and the cleaner IPJ concentrate is pumped to the concentrate thickener.

The cyclone overflow reports to the agitated rougher flotation conditioning tank; process water is added as required to reach the required %w/w solids and reagents (collector, promoter, and frother) are added as required. The slurry overflows from the conditioning tank to the rougher flotation cells. The rougher flotation tails are pumped to the plant tails thickener, while the rougher flotation concentrate is pumped to the cleaner flotation cells.

The cleaner flotation tails flow by gravity to the rougher flotation cells for re-processing, while the cleaner concentrate is pumped to the concentrate thickener.

CONCENTRATE TREATMENT PLANT (CTP).
Concentrate filter cake, or reclaimed concentrate material, is discharged to the regrind mill discharge sump, mixed with cyanide laden barren solution and regrind mill discharge, and is then pumped to the regrind cyclone cluster, consisting of six 100CVX cyclones with four operating and two standby.

The underflow from the regrind cyclones reports to separate batch centrifugal separators (BCS) operating in parallel. The cyclone overflow reports to the agitated continuous InLine Leach Reactor (ILR) feed tank. The tail from the centrifugal separators reports to the regrind mill, while the concentrate reports to the BCS concentrate transfer tank and is pumped to the batch ILR by the BCS concentrate transfer pump.

The concentrate from the BCS is accumulated in the Batch ILR feed cone in preparation for transfer into the ILR drum for leaching on a 24 hour cycle. The gold is leached into solution using sodium cyanide and oxygen as reagents, with sodium hydroxide used for pH adjustment. Flocculant is added at the end of the leach cycle to the batch ILR discharge to aid in solid settling and to obtain clarified pregnant leached solution (PLS) which is pumped to the batch ILR EWC feed tank.

The overflow from the cyclones flows to the continuous ILR feed tank, and pumped to the continuous ILRs. The gold is leached into solution using sodium cyanide and oxygen as reagents with sodium hydroxide used for pH adjustment. The discharges from the continuous ILRs are sent to the leach residue thickener.

The contents of the leach residue filter feed tank are pumped to the leach residue filter. The filtrate from the filter is sent to the PLS tank via the trash screen. After the filtration phase, the cake is washed with barren solution to recover moisture laden gold in the filter cake. The discharges from the wash cycle, drips, and the cake drying are also sent to the PLS tank via the trash screen, while the cake is discharged onto the leach residue filter discharge conveyor, into the detox holding tank.

The PLS tank is pumped to the resin absorption column circuits in parallel. Each circuit consists of a pulsed absorption column, consisting of a series of six compartments loaded with approximately 800 L of AuRIX®100 Resin per compartment. The column stages are separated by 500 µm screens which hold the resin in place.

Leach solution is pumped as feed into the bottom of the column, while the resin is progressed down the column. The gold loaded resin is educted out of the resin columns’ 1st stage over the loaded resin screen into the resin stripping vessels. The resin is stripped by desorption using a heated sodium hydroxide and sodium benzoate solution in closed circuit with electrowinning cells for gold recovery.

The stripped resin is then transferred back to the resin column via the barren resin tank or periodically acid washed before it is returned to the resin column. The acid wash solution is neutralized with sodium hydroxide, then transferred to the tailings transfer tank.

The Batch ILR PLS is recirculated through a dedicated electrowinning cell to plate out the gold from the PLS onto the Stainless Steel (SS316) mesh cathodes. On completion of the electrowinning process, the barren solution from the cells, is pumped to the barren solution tank.

The gold loaded strip solution from the resin column circuit is pumped through two stainless steel electrowinning cells in parallel for gold recovery, with the return barren solution recycled back the resin stripping columns.

The electrowinning cells can be stripped at regular intervals by hand in the cathode wash station (70-CWS-01). The precipitates are then filtered, dried, and smeltered to produce doré in the gold room.

After drying, the precipitate is mixed with flux before being smelted in the diesel fired furnace. The doré produced is kept in a safe until secure cartage from site is arranged.