VMS deposits are also known as volcanic associated, volcanic-hosted, and volcano sedimentaryhosted massive sulphide deposits. They typically occur as lenses of polymetallic massive sulphide that form at or near the seafloor in submarine volcanic environments, and are classified according to base metal content, gold content, or host-rock lithology.

The Horne deposit is the largest Canadian Au rich VMS deposits in the Noranda district (331 tonnes of Au produced historically from 54.3 Mt of ore at 6.1 g/t Au; Kerr and Gibson, 1993).

The typical morphology of Au-rich VMS deposits consists of a lenticular massive sulphide body with associated underlying discordant stockwork-stringer feeders and replacement zones. At Horne, zones of auriferous sulphide veinlets with Fe-chlorite selvages account for some of the Aurich mineralization (Kerr and Mason, 1990), however, the deposit lacks a well-defined stringer zone (Poulsen et al., 2000).

The vertical extent of the stockwork is typically larger than its lateral extent. The lateral extent of the deposit is typically a few hundred metres, but in some cases where the deposits are overturned, the mineralization has more than 2 km of known vertical extent (Horne and LaRonde Penna deposits). The thickness of the massive sulphide lenses is highly variable, especially when subjected to deformation (shortening), but is commonly on the order of a few tens of metres.

Mineralization is typically hosted by felsic volcanic flows and volcaniclastic rocks (or their metamorphosed equivalents) near or at the interface with basaltic andesite, andesite or clastic sedimentary strata. The Horne deposit is contained within a fault-bounded block of tholeiitic rhyolite flows and pyroclastic breccias and tuffs in contact with andesite flows to the east. It is juxtaposed against andesite flows and a diorite intrusion to the south, and rhyolites to the north that contain the Quemont deposit, another auriferous massive sulphide deposit (Poulsen et al., 2000) potentially related to the same giant hydrothermal system responsible for the formation of the Horne deposit.

At the Horne deposit, most rhyolitic rocks within the fault-bounded block have been affected by weak sericitization and silicification that become more intense near the sulphide mineralization, where alteration is characterized by a quartz-sericite-pyrite assemblage (Poulsen et al., 2000). Chlorite alteration, which locally contains elevated Cu and Au values, is largely restricted to the immediate footwall and sidewall of the deposit, except for local discordant zones in the footwall (Barrett et al., 1991).

The Horne Mine orebodies (Upper H, Lower H and Horne 5 deposit) dip subvertically within rhyolitic flows, breccias, and tuffs that are bounded by the Andesite and the Horne Creek faults. Least-altered rhyolites have low K2O contents and other geochemical features that place them within the FII tholeiitic series (Lesher et al., 1986). Graded volcaniclastic beds, metal zoning in the orebodies, and locations of chloritized-mineralized rhyolites indicate that the volcanic sequence youngs to the north. The volcanics in the fault wedge are variably silicified and sericitized, and local zones in the orebody sidewalls and footwall are chloritized.

The H orebodies formed podiform masses up to 120 m wide, 100 m thick, and 300 m in downplunge extent, consisting of chalcopyrite pyrrhotite-pyrite gold ore.

A semi-continuous Cu-rich base (up to 15 m thick) exists above the footwall and adjacent to the sidewalls of the orebodies. The ore changes stratigraphically upwards from a chalcopyrite-rich base, through middle pyrrhotite-pyrite-rich zones, to upper pyrite rich zones. Gold enrichments occur in some of the Cu-rich ores but also in overlying pyritic ores and in adjacent host volcanics. Cu-Au bearing chloritized rhyolites occur mainly in the western and eastern sidewalls and at down plunge terminations of the H orebodies.

The Lower H orebody is stratigraphically overlain by a massive to semi-massive sulphide body, referred to as the Horne 5 deposit (Sinclair, 1971). This tabular zone consists of numerous lenses of massive pyrite interbedded with intensely altered felsic volcaniclastic rocks. The Horne 5 deposit extends for a strike length of more than 1,000 m to a depth of at least 2,650 m and ranges from approximately 30 m to 140 m in thickness (Sinclair, 1971; Fisher, 1974; Gibson et al., 2000).

Pyrite is the predominant sulphide in this zone, but scattered portions contain sphalerite, chalcopyrite, and gold in economic concentrations (Sinclair, 1971). Sphalerite is the second most abundant sulphide in the Horne 5 deposit, although it is virtually absent in the Lower H, but it is far less abundant than pyrite. Chalcopyrite is a common mineral in many parts of the Horne 5 deposit but usually is present in only very minor amounts. Pyrrhotite is even less abundant than chalcopyrite and galena is rare in the Horne 5 deposit. A set of drill-core samples of massive sulphides have been analyzed for Cu, Au, Zn, Ag by Barrett et al., (1991). Based on twenty-five (25) samples (thirteen (13) samples from level 27 and twelve (12) samples from level 49), the Horne 5 deposit contains, on average, 0.12% Cu, 1.8% Zn, 1.4 g/t Au, and 26.6 g/t Ag.

The Horne 5 deposit extends from above level 21 down to below level 65. It attains significant thickness and breadth only below level 23. Below level 23, the Horne 5 deposit ranges from 15 to 125 m thick.

The main mining methods selected for Horne 5 Project are transverse long hole stoping with primary and secondary sequence. This selection is based on the geometry of the mineralized zones, their vertical dip and the competency of the rockmass to allow high production rate. Some sectors are also mined using longitudinal long hole retreat and pillarless sequences.

This mining method involves accessing stopes using two transverse draw points: one above the stope for drilling and loading of bulk explosives, and one below for extracting blasted material. The empty stope is then backfilled to allow mining of the adjacent stopes. Stopes will be mined according to a primary/secondary sequence, thus increasing flexibility and productivity while ensuring rock stability. For levels L1310 to L1030, the thickness of the ore zones (approximately 100 m) permits the haulage drifts to be excavated inside the ore zone thus reducing waste and initial capital cost, and accelerate the mining schedule. So, a longitudinal retreat mining sequence will be used for the last stope of each draw point on these levels.

Stope dimensions will vary with depth to reduce the potential dilution and the risk of seismicity.

The transverse long hole mining method allows for a high degree of mechanization and automation and is amenable to tele-operation. Production technicians will be able to operate LHDs from a control room at surface, resulting in higher equipment usage rates, better productivity and less maintenance. This approach adds four extra hours of effective production per day by allowing to extract ore in between shifts and at lunch time. It is also estimated that the use of automation will reduce both the number of pieces of equipment and the number of personnel required to reach the production target.

The time required to mine a given stope, from the time it is drilled to the time its backfill has cured, is expected to be 65 days. However, the effective cycle time is 50 days as it will be possible to drill an adjacent stope while the backfill is curing. To achieve the production target averaging 15,500 tpd over the LOM, 16 stopes must be active (drilling, extraction or backfilling) at any given time. A total of 9 stopes averaging 50,000 t each are projected to be mined every month in Phase 1, while 18 stopes averaging 26,600 t will be needed in Phase 2. It could be possible to shorten the cycle time as preliminary results show that the necessary backfill strength could be achieved within 21 days.

The mineralized zones will be divided into mining horizons ranging from four to nine stopes high. Each horizon will in turn be subdivided into pyramids, providing many available workplaces to achieve the production target.

Two key aspects that influence the design of the underground distribution systems for both the paste backfill and tailings slurry are the capacity of the paste plant and secondly, the mix design. The plant capacity will determine the flow rates for all of the process streams, whereas the mix design is determined in order to achieve the required backfill strength at the lowest binder content while maintaining a material that can be pumped or delivered by gravity.

The processing facility will be used to process an approximate average of 15,500 tpd of mineralized material over the LOM. The flowsheet consists of a semi-autogenous ball milling (“SAB”) circuit to grind primary crushed material to a target P80 size of 55-60 microns.

Crushing is performed in the underground mine. The crushed material is hoisted to the surface and is discharged via a stockpiling conveyor onto an elongated slot established in backfilled and compacted material, covered with rip-rap.

As for the copper and zinc cleaning circuits, consideration was given to the yearly variability of the feed grades of copper and zinc, respectively. The design feed grades were established by considering the peak years for copper and zinc and adding a margin 25% above these values, to reflect short-term peaks to be expected in the plant feed stream during the course of said years. The design flows within the cleaning circuits indicated by such consideration were matched with a plant operating at its nominal throughput, not the design value, to avoid oversizing of the circuits in question.

The overall flotation section of the process plant is divided into three major circuits dedicated to recovering either copper, zinc or pyrite. Gold and silver units are also reporting to each of these concentrates, with payable contents placed in the copper and zinc concentrates sent to smelters. Additional precious metals units reporting to the pyrite concentrate and tails are then directed to their respective leaching circuits, following the pyrite flotation section.

The copper and zinc flotation circuits are further sub-divided in two major sections: a roughing and a cleaning circuit, whereas the pyrite circuit only comprises a roughing section.

The critical streams needed to provide on-going feedback to the operators, in terms of achieved grades and calculated recoveries, are fitted with samplers continuously extracting a representative sample of the targeted flows. These samples are typically reporting, through vertical froth pumps, to an X-ray diffraction based on-stream analyzer (“OSA”) system that provides assay data for each stream. Accounting samples are also collected on a shift basis at the OSA sample feed multiplexing stations for assessing metallurgical performance with composite samples to be processed in an assay laboratory. These composite samples are created through the accumulation of multiple individual increments taken throughout each shift. Each increment is fed to a lab-scale filtration bench fitted with a vacuum pump so as to permit the accumulation of a few kilograms of filtered cake per shift, thus yielding a representative sample of the whole shift duration without operator intervention during the period.

Eleven flotation streams report to one five and one six-position multiplexers located above the OSA room and are sequentially presented to the OSA for measuring the copper, zinc and iron content of the streams, as well as providing an indication of the slurry density (percent solids).

Pumping of intermediate flows between flotation stages is provided with centrifugal pumps tied to dedicated pump boxes. Variable speed capability is provided where needed to cope with large expected inflow ranges. The pumps are duplicated, with the provision for one stand-by unit per duty, with all the required automated valving to complete a switch to remotely operated unit.

The fine fraction of the cyclones (cyclone overflow) flows to a row of five copper rougher flotation cells. The flotation cells are 130 m3 tank cells, providing a design retention time of 20 minutes. The rougher concentrate is sent to the cleaning circuit, while the tailings proceed to the zinc rougher feed.

The cleaning circuit is arranged along a conventional countercurrent configuration, with three stages of cleaning closed by a cleaner scavenger. The concentrate from each step is brought as the fresh feed source to the following stage, while the tailings are recycled to the previous stage as a circulating load. The cleaning circuit is opened at the cleaner-scavenger, with the tailings making their way to the zinc rougher circuit feed, and at the third cleaner with its concentrate brought to the copper concentrate dewatering circuit.

The first cleaning stage comprises three 10 m3 tank cells, the second and third stages are made up of two and one 10 m3 cells, respectively, while the cleaner-scavenger has three 10 m3 cells. The design retention times provided thus equate to 12 min, 17 min, 11 min and 17 min, respectively.

The copper rougher and cleaner-scavenger tails are combined to form the feed to the first of two conditioners arranged in series. These are used to activate the sphalerite and adjust the pH. The final conditioner overflows into a row of five zinc rougher flotation cells. The flotation cells are 70 m3 cells providing a nominal retention time of almost 11 minutes. The rougher concentrate is sent to the zinc cleaning circuit, while the tailings proceed to the pyrite rougher feed.

Similar to the copper cleaning circuit, the zinc cleaning circuit is arranged in a conventional countercurrent configuration, with three stages of cleaning closed by a cleaner scavenger. The zinc cleaning circuit is opened at the cleaner-scavenger tails, sent to the pyrite rougher circuit, and at the third cleaner concentrate, brought to the zinc concentrate dewatering circuit.

The zinc rougher and cleaner-scavenger tails are fed to the six pyrite rougher flotation cells. The flotation cells are 130 m3 cells providing a nominal retention time of 24 minutes. The rougher concentrate and tails are sent to their respective dewatering step, ahead of cyanide leaching for gold recovery.

The leached slurries of pyrite concentrate and tails flow through their respective line of CIP tanks, each with an inventory of 6 mesh x 12 mesh activated carbon used to adsorb the dissolved gold and silver values (along with a portion of other base metals that may be present, mainly copper), as leached by the cyanide. The CIP train for the pyrite concentrate includes eight tanks of 300 m3 each, while the one for the pyrite flotation tails has six tanks of 130 m3. All the tanks in a series are of the same diameter and height.

For transfer of the loaded carbon inventory of the first tank of a given CIP tank row to the elution circuit, a recessed impeller pump (one operated, one stand-by per row) is used to drain the targeted tank of its slurry and carbon while it is bypassed by the fresh slurry flow. The slurry extracted from the tank is presented to one of two parallel vibrating screens for separating the coarser carbon from the solids in the slurry. These and most of the slurry water are returned to the head of the CIP circuit where they came from while the washed loaded carbon is accumulated in the acid wash vessels, located at the head of the gold elution circuit.

Carbon elution, or stripping, is following the high-pressure Zadra process. It is initiated when a barren strip solution of 2% NaOH and 0.2% NaCN circulates at an elevated temperature and pressure through the two elution columns, arranged in a serial configuration, at a flow rate of 2 bed volumes per hour for 8 hours. Sequential stripping can be implemented, where the a single column is first submitted to stripping and, as the pregnant solution grade is decreasing, the second one is brought in line to boost again the grade presented to the electrowinning (“EW”) circuit downstream, keeping the EW circuit efficiency higher for a longer period of time without changing the amperage of the EW cells.

Three parallel trains, of two EW cells each, recover gold and silver from the pregnant strip solution. The solution exiting the cells reports to the EW cell discharge tank and is pumped to the barren stripping solution tank.