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.

Mineable resources will be accessed through the Quemont No. 2 shaft, which will be deepened to a final depth of 1,910 m. The mining sequence was split into two phases to allow the deepening of the shaft. Phase 1 will be from level L1310 to level L710. Phase 2 will be from level L2060 to level L1340. Each phase will be composed of several horizons in which pyramid sequences will be established. The pyramids will allow many stopes to be available at any given time to allow flexibility, and the pyramids are sequenced to minimize stress issues that could potentially disrupt production.

The mine plan for the Horne 5 Project proposes a production rate averaging approximately 15,500 tpd of ore over the LOM. Calculations were performed assuming Sandvik LH625 LHDs with 25 t payloads and an expected average travelling distance of 200 m. The cycle times assume one minute of loading time per bucket, plus an additional one minute of unloading time. Each day assumes 20 productive hours of tele-operation, with 85% usage and availability rates. Under these assumptions, it is expected that each LHD has an average extraction capacity of 5,500 tpd for a combined production capacity of 16,500 tpd. One spare LHD has been considered to ensure the required LHD availability. Additionally, LH517 development LHDs will be equipped with teleoperation capabilities to ensure the required availability of LHD to serve as an additional backup or to temporarily increase production.

Production drilling will be carried out using four Sandvik DU412I drills. The purchase of one spare drill has been considered. The proposed 4.5 m by 5.0 m drill pattern results in a production drilling ratio of 32.0 t/m. Each drill is expected to drill an average of 126 m per day, for a capacity of 16,128 tpd.

Once the pyramidal mining sequence has achieved maturity, it is expected that a steady output of hoisted material will be maintained due to increased mining flexibility afforded by higher stope availability and thus allowing an approximate LOM average of 15,500 tpd to be mined.

In Phase 1, the mine plan provides for 866 stopes with an average of 50,000 t each (including dilution and a 95% mining recovery). Phase 1 will sustain the full production rate during the first six years. During the seventh year, Phase 2 stope extraction will start up progressively to establish the pyramidal production shape. Production will continue simultaneously in both Phase 1 and Phase 2 until Phase 1 is depleted. For Phase 2, a total of 1,299 stopes are planned at an average tonnage of 26,600 t. Dimensions for stopes in Phase 2 are smaller than in Phase 1 to accommodate the higher stress levels at greater depth.

The proposed hoisting system is designed to achieve a production of 23,180 tpd during Phase 1 of production for the loading station at level Q1180 and 16,530 tpd during Phase 2 for the loading station at level Q1851. The shaft is designed for high-speed skip operation. If required, the system could hoist ore using one skip only from both loading stations at level Q1180 and Q1851 in the transition period between Phases 1 and 2., resulting in a production rate of 11,590 tpd. Production rates shown in the following tables are based on a hoist availability of 19 h/d for 335 d/y.

Crushing
Crushing is performed in the underground mine. The crushed material is hoisted to the surface and is discharged via a stockpile feed conveyor onto a stockpile located in a domed structure.

The plant feed is reclaimed using two apron feeders located beneath the covered coarse ore stockpile. The feeders are equipped with variable speed drives to control the tonnage. Cameras, coupled to a particle size analysis software, are installed between the two feeders and after the second one. Their output allows adjustment of the mixture of ore size distribution extracted from the stockpile and fed to the SAG mill. The feeders are sized to allow operation of the plant at the design throughput with only one unit in operation, although both would typically be running to provide a better size distribution to the SAG mill. A sump pump is installed near the tail end of the reclaim tunnel, which is the low point within the sloping tunnel. An emergency exit is located at this end. Rails with electric hoists are mounted strategically around the feeders to facilitate maintenance and handling of spalling bars, used to isolate the feeders from the piled material above them, when required. Because of the stockpile configuration, the only ore available from it is in the live portion of the pile since no mobile equipment could be safely working to displace a portion of the dead load to increase ore availability. A live volume of approximately 5,000 m 3 is provided, or about 16 hours of operation at nominal throughput.

Grinding Mills Sizing
Grindability data for establishing the required grinding power requirements was obtained from a series of samples representing the different main rock types in the orebody: rhyolitic tuff, rhyolitic breccia, massive (“MS”) and semi-massive sulphides (“SMS”). The outcome of the comminution testwork outlined strong correlations of all the grinding parameters measured against the sulphur content of the samples tested.

These inverse correlations (e.g. ore getting softer as sulphur content is increasing) allowed to select a design sulphur feed grade, for the purpose of sizing the grinding mills, based on the analysis of the sulphur grade distribution expected from the mine plan. Both the yearly and LOM variability were considered to select the appropriate hardness design criteria to be used for grinding mill sizing. Based on the sulphur distribution statistics shown and the harder (e.g. lower sulphur content) material thus indicated for the latter years of the mine plan, a design sulphur content of 10%, per the yearly 20 th percentile values, was selected for determining the values of the grinding indices to be used for establishing the required grinding mill power.

The configuration of the retained grinding circuit includes a semi-autogenous (“SAG”) mill operated in closed-circuit with a vibrating screen located at its discharge end. The oversize fraction of the screen is returned to the SAG mill feed through a series of conveyors. There is no pebble crusher included in the basic configuration, based on the low rod mill to ball mill work indices ratio, but a pebble bypass capability is included to facilitate its inclusion or benefit from their rejection if they were to be shown as barren. The SAG mill circuit is followed by a ball mill, operated in closed-circuit with hydrocyclones arranged in a circular cluster.

SAG Mill Circuit
The reclaimed coarse ore is conveyed to the SAG mill. Water is added to the mill feed chute to achieve the desired slurry density within the mill. Nominal tonnage is rated at 679 tph (15,000 tpd at 92% availability), with the design ore hardness, whereas the plant downstream equipment sizing can readily accommodate 15% more tonnage with softer ore. A SAG mill size of Ø10.97 m x 5.41 m (Ø36 ft x 17.75 ft) EGL was selected for providing the power draw indicated of 8.5 MW at the pinion with the design ore. The mill drive train includes a twin-pinion arrangement powered by two low- speed synchronous motors of 6.7 MW (9,000 hp) each, making these units equal in size to those required to cover the ball milling duty, for spare commonality. The mill is operated with a charge of Ø125 mm steel balls intermittently added to the mill feed chute so as to maintain the desired media filling required to uphold the power draw and throughput capability.

The coarse fraction obtained from the 3.66 m x 7.32 m (12 ft x 24 ft) screen installed below the SAG mill discharge trunnion is returned to the SAG mill feed conveyor via a series of transfer conveyors. The pebble flow could also be bypassed and accumulated in the process plant or re- circulated to the process plant feed stockpile. Nevertheless, the initial SAG mill lining set is expected to include slotted discharge grates only, without pebble ports. The screen, under these conditions, would be receiving only material passing through the 12 mm wide slots and the size of its oversized material serves as an indication of the gradual wear of the grates, as well as a protection for the pump box and pumps underneath in case of a catastrophic grate failure. The screen undersize drops into a pump box receiving this stream as well as the ball mill discharge. A centrifugal pump feeds this slurry to the cyclone cluster. The transfer size to the ball mill (T 80 ), under design conditions, is estimated at 1,200 µm, with a screen deck featuring apertures of 3.7 mm.

Ball Mill Circuit
A ball mill of Ø7.62 m x 11.43 m (Ø25 ft x 37.5 ft) EGL was selected for the power draw requirement of 11.5 kWh/t indicated for the design ore (at the pinion). The drive train is based on a twin-pinion configuration fitted with low-speed synchronous motors of 6.7 MW (9,000 hp).

The ball mill is operated in closed-circuit with a cluster of Ø400 mm cyclones producing an average product P 80 of 55 µm as flotation feed, with a pulp density of 35% solids (weight basis). A sufficient number of cyclones are provided to cover a circulating load of up to 450%, whereas the nominal flow is based on 350%.

The cyclone cluster is fed with the combination of the SAG screen undersize and ball mill discharge slurry via a variable-speed centrifugal pump (with one stand-by unit). The ball mill is charged up to 30% of its volume with a mixture of Ø50 and Ø63.5 mm steel balls, added intermittently via a bucket placed on the ball mill feed chute with the overhead crane of the grinding bay.

The grinding efficiency of the circuit is monitored by an on-line particle size analyzer, receiving a sample of the cyclone feed and cyclone overflow slurry, thus providing a regular feedback of the P 80 to the flotation as well as the size distribution of the circulating load around the ball mill.

The processing facility will be used to process an approximate average of 15,500 tpd of mineralized material over the LOM.

The process plant’s major areas consist of the following:
- Ore reclaiming and Grinding;
- Differential flotation of copper, zinc and pyrite;
- Base metal concentrates dewatering;
- Fine regrinding of the pyrite concentrate: gold leaching from pyrite concentrate and flotation tails, followed by their respective CIP and cyanide destruction circuits;
- Gold elution and refinery;
- Paste backfill preparation with a mixture of PCTs and PFTs;
- Reagent preparation circuits;
- Process and fresh water, as well as low-pressure and compressed air distribution systems. All these areas are geared to produce copper and zinc concentrates and gold doré for delivery to buyers.

Flotation Circuit Description
The overall flotation section of the process plant is divided into three major circuits dedicated to ........