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 ........