Magmatic sulphide deposits are sulphide mineral concentrations in mafic and ultramafic rocks derived from immiscible sulphide liquids.

There are two main subset deposit types of magmatic sulphide deposits: first nickelcopper (PGE) deposit and second PGE deposits. In nickel–copper sulphide deposits, Ni and Cu are the main economic metals. Nickel constitutes the main economic commodity, generally at grades of about 1–3%. Copper may be either a co-product or by-product. Cobalt, PGE, and Au are common by-product metals.

The second major deposit type is PGE-rich, with the PGE elements associated with sparsely dispersed sulphides in very large to medium sized, typically mafic/ultramafic layered intrusions.

Host rocks are predominantly mafic to ultramafic igneous rocks. Significant amounts of ore are located in footwall country rocks of diverse meta-sedimentary or metaigneous origin and composition.

Some Ni–Cu–PGE deposits occur as individual sulphide bodies associated with magmatic mafic and/or ultramafic bodies. Others occur as groups of sulphide bodies associated with one or more related magmatic bodies in areas or belts up to tens, even hundreds, of kilometres long.

Sulphide minerals may be concentrated in structurally low areas at the base of intrusions or flows or may be in zones where silicate magma interacted with xenoliths. Sulphide mineral concentrations in layered, cumulate sequences may be related to major lithological features such as cyclic-unit boundaries, unconformities, chromite seams, pegmatoids or stratigraphic intervals characterized by major changes or discontinuities in cumulus minerals.

The location of sulphide concentrations in conduits at Talnakh–Noril’sk and Voisey’s Bay, and near conduits in particular komatiitic deposits, suggests that sulphides accumulated where the flow rate of magma was reduced and the entrained sulphides settled gravitationally to form rich basal concentrations.

Gangue mineralogy is the same as that of the host rock and consists primarily of plagioclase, orthopyroxene, clinopyroxene and olivine. Minor secondary phases are serpentine minerals, talc, magnetite, calcite, epidote, sericite, actinolite, chlorite, tremolite and clay minerals.

In general, within the Current Lake and Bridge Zones, sulphide mineralization is disseminated, ranging from a few percent to >25% sulphides.

Disseminated sulphide grains can range in size from 0.5 mm to as much as 1 cm in size and consist of pyrrhotite, chalcopyrite, pentlandite, minor pyrite and rare cubanite and violarite.

Basal net-textured (25–50%) sulphide and massive sulphide intervals have been intersected in core drilling locally and are more common in the Bridge Zone than the Current Lake Zone. In both of these zones the main concentrations of mineralization occur as elongated high-grade pods connected by narrower medium- to low-grade zones.

Significant massive sulphide veins, generally 1–2 cm wide, occur within the Current Lake Zone. These veins are typically either sub horizontal or near-vertical, and are interpreted by Magma Metals geologists to be the result of segregation of molten massive sulphide during the cooling of the intrusion. Plagioclase occurs in these veins, which indicates that plagioclase had also not yet crystallized completely and was still partially molten.

The Beaver Lake sulphide mineralization is disseminated, ranging from a few percent to >25% sulphides, and is also interstitial to the silicate gangue. Disseminations can range in size from 0.5 mm to as much as 1 cm in size. Blebby sulphides are common and classic net textured and massive sulphide mineralization has been intersected regularly in core drilling within the western portions of the Beaver Lake Zone, particularly where it merges with the Bridge Zone.

In the Beaver Lake Zone, sulphide grades are generally consistent between the mineralization along the upper and lower contacts. However, the Cloud Zone generally has higher-tenor mineralization. The basal mineralization, particularly the western and “Spine” areas, forms a complex mesh of mineralized subzones that occur in depressions on the floor of the intrusion. These intersecting depressions appear to coincide with conjugate fracture sets within the underlying Archean meta-sedimentary rocks and may have formed by thermal erosion along the structurally-weakened fracture zones.

The conceptual mine plan was developed using only open pit methods. Mining would be at a rate of 1.5 Mt/a over a seven-year mine life.

Initially the company studied two mining scenarios, a standalone open pit and a hybrid case with a smaller open pit mine and underground operation, utilising a drift-and fill mining method. As the study progressed and the economic parameters became better defined, it became apparent that the stand alone open pit presented better economic results then the hybrid scenario.

The overall slope angles account for the likely future ramp configuration. All of the ramps in the Current Lake area will be placed on the east wall. An 18° overall slope angle was used in the overburden material.

The Thunder Bay North deposit displays very low sensitivity to the slope angles with respect to the contained amount of mineralization; however, the stripping ratio varies considerably with the change in the slope angles. The relatively shallow Current Lake area would not be particularly sensitive to pit slope changes; however, the high strip ratio Beaver Lake area would be more sensitive to slope changes.

Phase volumetrics by bench were manually scheduled, using maximum descent rates of ten 2.5 m benches per period in selective mining areas. Selective and bulk zones were tracked to allow individual costing on an annual basis.

Initial mining at the north end of Current Lake will require small equipment, which can mine selectively.

The three-stage process selected as the preferred process route is as follows:
• Crushing, grinding and flotation to extract the sulphides from the ore to produce an initial bulk concentrate. A gravity circuit would extract a significant proportion of the gold, which is output to the bulk concentrate.
• The bulk concentrate is then treated by PlatsolTM pressure oxidation to produce a pregnant leach solution (PLS), which contains the dissolved metals.
• The PLS is then treated via relatively simple and commercial process routes (reduction with metal (cementation) to produce both precious metals bullion and copper, followed by electrowinning of nickel and cobalt).

Four high value products will be made and either sold to metal refineries or direct to end-users:
• A high-grade precious metals (PGM and gold) bullion product to a PGM refinery.
• A high-grade silver product to a silver refinery
• Copper briquettes to market, and
• A nickel-cobalt alloy to market.

A sample of flotation concentrate from the Thunder Bay North project was tested in a 2 L Parr titanium batch autoclave, under standard Platsol™ pressure oxidation operating conditions. The concentrate was oxidized for two hours at a temperature of 225°C, under an oxygen partial pressure of 100 psi and in the presence of 10 g/L chloride ions in solution.

Greater than 90% of the platinum and about 90% of the palladium in the concentrate were solubilized, along with >90% of the nickel and cobalt. Four tests were carried out in which the effects of fine grinding the concentrate, increasing the acid concentration in solution, and lowering the chloride concentration to 5 g/l were examined. None of these changes had a significant effect on metal recoveries, except for lowering the chloride strength, which had an adverse effect on platinum recovery.

Several methods were considered at the conceptual level for recovery of revenue metals from the Platsol™ pregnant leach solution (PLS) within the constraint of keeping the hydrometallurgical operation simple and economical but providing upgraded products which would improve Project revenue due to reduced impact of smelter deductions.

The Platsol™ pregnant leach solution is treated with copper powder (an internal recycle stream) to reduce solution Eh by reduction of ferric iron, PGM and part (approximately 5%) of the copper. The resulting dilute suspension of PGM and cuprous chloride is thickened and filtered to produce a cake containing essentially all of the PGM (about 10 kg/day) with about 200 kg/day of copper chloride. The thickened pulp is contacted with fresh Platsol™ PLS to dissolve most of the CuCl. The precipitate leach solution recycles to the PGM precipitation.

The PGM–AgCl–CuCl residue from the PLS leach (15–20 kg/day net product) is heated in a strong chloride solution (salt or ferrous chloride from copper precipitation) to dissolve silver and copper. Filtration produces a PGM + gold cake which is vacuum dried, de agglomerated and packed for shipment to a refinery.

Filtrate is cooled to precipitate an AgCl–CuCl precipitate, >65% silver, for sale to a silver refinery or chemical producer.

Copper remaining in the Platsol™ solution after PGM precipitation is precipitated quantitatively as CuCl by reduction on copper powder. CuCl is thickened and further reacted with scrap iron to produce copper powder (“cement” copper). Copper powder equivalent to net production, about 7.8 t/d, is filtered, washed and briquetted for sale. Approximately 8 t/d of copper powder is recycled to reduce ferric, precipitate PGM, and CuCl. Net ferrous chloride product is recycled to the Platsol™ feed, partly replacing salt as a source of chloride ion.

Precipitation of copper as copper chloride from an acid sulphate–chloride leach solution using copper powder and cementation of copper from the precipitate slurried in ferrous chloride brine was commercially practiced for many years at Chuquicamata (McArthur and Leaphart, 1961).

Essentially copper-free solution passes to limestone neutralization and iron oxidation/jarosite precipitation; iron precipitate and gypsum are filtered and pass to tailing. Nickel (and cobalt) is then precipitated as hydroxides/basic sulphates, along with gypsum, by lime adjustment to pH 8. The gypsum-nickel hydroxide precipitate is thickened, filtered and repulped with a nickel sulphate pH 6 electrolyte and passed as a slurry through EMEW® cells to plate nickel cobalt alloy which is stripped and sold. For revenue computation, cobalt is valued at the price of nickel and marketing is limited to manufacturers of non-nuclear grades of stainless steel. The precedent for this is cobalt recovery from leach solutions in the Congo (Bouchat and Saquet, 1960).

Acid generated by plating of nickel is consumed by hydroxide so the pH and Ni(Co) contents of the circulating solution remain constant. Gypsum, with any residual nickel hydroxide, is removed from the circulating electrolyte by filtration and recycled to the jarosite precipitation. The final step in solution treatment is pH adjustment with lime to precipitate magnesium as its hydroxide and the associated sulphate as gypsum to prevent discharge of an excessive level of soluble magnesium sulphate.