The Bisie tin deposit is a cassiterite-bearing stock-work or vein system adjacent and possibly distal to underlying source granite. The mineralization at Bisie is unusual and different from other classic tin deposits. The deposit has up to 0.5 % rare earth elements (REE) and very high grade tin (with some sample assays reaching greater than 60 % Sn).

The deposit can be simply described as a number of steeply dipping tabular sheets of variable grade mineralization consisting of irregular veins and disseminations of cassiterite that is complex on a small scale.

On a regional scale the metamorphic rock units generally strike northwest-southeast. The ridge hosting the Bisie mineralization strikes north-south as far as the Oso River in the north, after which the strike of the ridge changes to the northwest-southeast. Regional scale folding is evident in the satellite imagery.

The stratigraphic rock package hosting the Bisie deposit has been divided into five separate units.
From hangingwall to footwall, a general description of the major units is as follows:

• Carbonaceous shale (CBSH) - dark grey to black, thinly bedded (0.5 cm to 2 cm), fine grained, carbonaceous siltstone-shale greater than 150 m true thickness. Contains abundant quartz-tourmaline-carbonate veins and minor pyrite.

• Meta-sediment (METS) - pale grey, thinly bedded (0.5 cm to 2 cm), fine-grained siltstone-shale 30 m to 40 m true thickness. The host rock has been moderately silica-magnetite altered. Magnetite occurs as either discrete bands (1 cm to 2 cm), pervasive, disseminated alteration or stock-work veins.

• Quartz-sericite schist (QSSH) - in drill core appears more like a feldspar-rich, polymictic tuff. Pale to dark grey-green, thinly bedded (~1 cm) to massive, medium grained (1 mm to 5 mm), feldspar-sericite rich tuff 80 m to 90 m true thickness.

• Mica schist (MSCH) - pale to dark grey, laminated to moderately banded (0.5 cm to 5 cm), fine grained mica-rich schist 100 m to 150 m true thickness. Intensity of dark and light coloured bands varies according to biotite-muscovite content.

• Amphibolite (AMPH) - the current interpretation is that this unit was originally a separate mafic-ultramafic unit hosted within the MSCH overprinted with intense chlorite-talc-garnet alteration; however this has not been confirmed. Dark green to black, moderately banded (1 cm to 5 cm) to massive, fine grained to porphyroblastic (garnet), chlorite schist 20 m to 30 m true thickness. Hosts the tin and base metal mineralization at Bisie.

The stratigraphic package dips east between 60° and 75° and appears to steepen down-dip and towards the east. The Mpama North ridge crest is more or less defined by the AMPH which probably resists erosion due to the massive, coherent nature of the rock and high chlorite content. The QSSH appears more susceptible to weathering and erosion due to its' high feldspar-sericite content. The base of complete oxidation (BOCO) is approximately 10 m along the ridge crest and approximately 50 m in the OSSH. Overall the BOCO averages 30 m.
The units most affected by alteration are the QSSH and MSCH particularly in the hangingwall of the tin mineralization:

• QSSH - contains areas where the original rock has undergone intense silica alteration (SILZ) and also intense sericite alteration of the original QSSH. To a lesser degree, QSSH is effected by biotite alteration and named biotite schist (BSCH). Where intensely chloritised the OSSH is termed amphibolite (AMPH)

• MSCH - is frequently termed BSCH where it comprises intense biotite alteration of the original MSCH is also referred to as amphibolite schist (ASCH) in places which usually occurs as a weak, chlorite alteration halo surrounding the AMPH.

As the alteration results in the same rock types in different parts of the stratigraphy, there appears to be an overlap between the MSCH and QSSH making the lithological boundary poorly defined in places.

Mineralization.
The bulk of the tin mineralization at Bisie is hosted within the north-south striking, east dipping amphibolite unit over 1 km to 3 km along the Mpama ridge, east of the granite intrusive. Mineralization is multi-phase and the paragenetic sequence appears similar in nature to the San Rafael deposit in Peru (Pearl, 2011; Alphamin Report, 2013).

Structural and mineralogical evidence from drill core indicates cassiterite was emplaced first, followed by copper mineralization in the form of chalcopyrite and bornite, then by lead and zinc mineralization occurring as sphalerite and galena. There is also evidence of late-stage quartz-chalcopyrite veining which cross-cuts the mineralization with veins trending north-northwest.

Chlorite alteration is widespread and appears to be the result of late stage fluids entering the system. The host rocks are predominantly highly chlorite-altered amphibolites and fine to medium grained, mica-chlorite-garnet schists. The tin and copper mineralization is predominantly found in zones dominated by intense chlorite alteration, however, cassiterite mineralization with no chlorite alteration has been intersected in the hangingwall and footwall vein zones hosted in MSCH.

The most common style of cassiterite mineralization observed in drill core comprises discrete, massive veins ranging from 2 mm to 1.80 m true thickness. Finely disseminated cassiterite is also present though less visible to the naked eye. High grade cassiterite chutes 20 m wide by 8 m thick have been historically mined by artisanal miners in the upper parts of the deposit, which most likely comprised a number of closely spaced vein sets.

The individual cassiterite veins are massive, pinkish brown, fine-grained and often botryoidal, and show compositional layering thought to be due to variations in iron content. This form of cassiterite has often been referred to as "wood tin".

The dominant structural control on mineralization is the north-south trending, brittle-ductile shear zone that runs more or less parallel to the cassiterite mineralized zones. It occurs predominantly as a single structure, with minor hangingwall and footwall splays particularly in the upper, more brittle parts of the structure. In the upper parts of the structure, above the 700RL, brittle fracturing has resulted in the development of up to four, lower grade vein systems whilst below the 700RL, ductile deformation has resulted in the development of a single, narrower, higher grade vein system.

Both tin mineralization and copper mineralization seems to be concentrated in two high grade chutes, referred to by Alphamin as the upper high grade chute and lower high grade chute. Mineralization between these two chutes is lower grade as these areas contain narrower, more widely spaced cassiterite veins. Both chutes run parallel to each other and plunge to the north at approximately 35°.

Most of the copper mineralization occurs in the form of blebs, lenses and veins, with the latter two being sub-parallel to foliation, and a late-stage quartz-chalcopyrite vein set trending northwest. In addition to the quartz veins, chalcopyrite also occurs with pyrite and to a lesser extent arsenopyrite. Chalcopyrite and bornite also occur as fracture fillings within the cassiterite. Higher grade copper intercepts usually occur adjacent to and overlapping the high grade tin intercepts, particularly with the lower high grade chute.

In contrast to Mpama North, lead-zinc mineralization is better developed at Mpama South. Most of the zinc mineralization is hosted within massive to semi-massive pyrite in the hangingwall of the cassiterite bearing zone coincident with minor lead and silver mineralization. Small quantities of zinc mineralization are also found together with the tin and copper mineralization. The degree of galena and sphalerite replacement of the pyrite appears to be structurally controlled with replacement along late stage structures.

A sub-level caving (SLC) mining method has been selected for the extraction of the Bisie ore body. There are limited narrow sections of the orebody, which make up less than 10 percent of the mineable resource, in which SLC is not applicable. In these areas handheld overhand mining will be employed.

Access to the underground workings will be established via a single trackless decline developed at 4.0m wide x 4.5m high. The portal of the Main Decline will be located centrally along strike of the orebody and at an elevation of 710m AMSL. The decline will advance in an easterly direction at an inclination of +2% for the first 50m to prevent the inflow of water into the underground working from intense rainfall events. The ramp will then become the main trucking decline at an inclination of -8%.

Levels are spaced 15m apart vertically and from the trucking decline an access cross-cut will be developed to the orebody. A footwall strike drive is planned 10m in the footwall of the orebody on every level and serves to provide advance geological information and access for equipment to the retreating stopes.

A return airway (RAW) will be established on 740 m RL. The RAW will link the top of the stoping horizon with surface and will be equipped with the main ventilation fan on surface. The RAW will provide an alternative means of egress from the mine in case of emergency and loss of access to the main decline.
The levels will all be connected be a service raise which also acts as a return air raise and emergency means of egress.

All development will be completed using mechanised drill, blast, support, load and haul methods. Waste rock will be hauled to a designated waste dump situated approximately 1 km from the portal position.

Over the life of mine, some 15 836m of waste development and 15 922m of ore drive development is planned.

Once the lateral footwall and ore drive development on a level is completed and the service raise linking to the level above has been developed and equipped the level is ready for stoping to commence.

Stope preparation starts with the development of a slot raise at the widest point of the block of ground to be mined. The slot raise will be mined at 3.0m x 3.0m and will hole between sublevels. Slot raises will be mined using longhole raising techniques. The drilling of the holes will be completed with the long-hole production rig. Multiple slot raises are planned per level as the stoping plans call for retreating the stoping face from the widest area of the orebody to the narrowest area.

When the slot raise is completed the raise will be opened up to the full width of the orebody, to form an open face across the width extent of the stope. On completion of this slot the stope is ready for production stoping to commence.

During development of the sublevel information regarding the orebody geometry and grade is collected by two main methods:

• Development sampling. Ore development will be sampled with chip sampling at regular intervals of approximately 3.0m.

• In-fill diamond drilling. Short diamond drill holes will be drilled to intersect the orebody at 15m strike spacing.

Prior to the commencement of stoping on a level a 3-dimensional model of the stope will be developed by the technical services department using all data available to form the basis for stope grade estimation, stope design and blast hole ring design.

The preferred method of operation of the stope is for a number of rings to be pre-drilled and then blasted, one ring at a time. However if ground conditions do not allow for pre-drilling of holes, i.e. holes closure is excessive between the time of drilling and charging of the holes, then rings will be drilled one row at a time, immediately prior to blasting. Only one ring is blasted and loaded at a time as this has been found to produce the best ore recovery in sublevel caving. Blasting of excessive strike length in a single blast allows waste from above to enter the draw point before the ore from the back of the blast has been loaded, resulting in poor ore recoveries.

The sequence of stoping in sublevel caving is top down. The stope faces will be managed so as to leave a safe lag distance of at least 20m between stopes, with the upper stope leading the stope below.

Blasted ore and waste will be loaded in the face by a 10 t load haul dump unit (LHD), transported to the level access drive and tipped directly into a truck for hauling to surface. The trucks will move ore to the RoM pad on surface and waste to the designated waste re-handling area. Dump trucks will only operate in the decline and level access drives. They will not enter the lateral footwall or ore drive development ends. The LHDs will move the ore and waste from the face, along the lateral development to the central level access position.

All material will be transported into the mine by light vehicle (LDV) or by utility vehicle. Underground workers will travel from surface to their working place by light vehicle.

The mining equipment fleet at Bisie has been selected considering the orebody geometry, minimising dilution and maximising productivity. A trade-off was conducted to determine which would be the more efficient between a suite made up of 6 tonne loaders and 20 tonne trucks and one of 10 tonne loaders and 30 tonne trucks. The study showed that the larger fleet was more cost effective as less units were required. The expected benefit of smaller excavations when using the smaller fleet is not applicable as the size of the decline is determined by the intake ventilation requirement, which is in fact greater in the case of the smaller fleet due to increased diesel powered equipment in operation underground.

The previous flow sheet for the Bisie tin plant was designed to treat 500 000 t/a of ore, however it would be operated for a reduced number of hours initially to bring down its throughput rate to 360 000 t/a. The revised flow sheet is designed to treat a maximum of 360 000 - 430 000 t/a of ore at an average grade of approximately 3.6% tin and produce approximately 15 500 t/a of tin concentrate. The other major difference is that the oxide flotation process has been replaced with a gravity circuit instead. The plant comprises of the following processes:

• Crushing of Run of Mine ("RoM") ore to -8mm.

• The -8mm material is fed into jig separators.

• Screening post the jigs separates the crushed ore into -8mm +1mm and -1 mm fractions.

• The-8mm +1mm HG jig concentrate is milled to 80% -425pm and processed using gravity spiral concentrators and shaking tables.

• The-8mm +1 mm HG jig tailings is discarded to a tailings stockpile, part of the tailings storage facility (TSF).

• HG shaking table concentrate is bypassed directly to product recoveries.

• HG shaking table middlings is milled and directed to sulphide flotation circuit.

• HG spiral tailings is directed to the LG regrind circuit.

• The -1 mm screened from the jig floats and sinks is processed using gravity spiral concentrators and LG shaking tables. The concentrate from the circuit is combined with the HG shaking table middlings and ground ahead of sulphide flotation. The middlings from the LG circuit is combined with the HG spiral tails and reground to 80% -106µm. The combined stream is then directed over another stage of shaking tables whose concentrate is directed to sulphide flotation circuit.

• The -1 mm spiral tails, sulphide flotation and LG regrind shaking table rejects are thickened and discarded.

• Combined final gravity concentrates and sulphide float product is treated through a magnetic separator to remove iron, then filtered and bagged for sale.

The crushing circuit will have an utilisation of 5 200 hours annually and the remainder of the plant will have a utilisation of 7 000 hours annually.

Front End Crushing and Screening.
The ROM material will then be withdrawn from the bin by apron feeder and through a jaw crusher, which will reduce the material from nominally -450mm to nominally -150mm. The jaw crusher product will be conveyed to the secondary crusher sizing screen, whose oversize will be conveyed to the secondary crusher feed bin, while the screen undersize will feed the tertiary crusher screen. The tertiary crusher screen will also be fed by the products on the secondary and tertiary crushers in a closed circuit arrangement. The tertiary crusher screen oversize will be conveyed to the tertiary crusher bin while the screen undersize will be conveyed to the plant feed stockpile.

The secondary crusher will produce a product of nominally -32mm while the tertiary crusher will produce a product of nominally -8mm, both of which will be collected on a common conveyor and conveyed back to the tertiary crusher screen.

The crushing and screening circuit will operate on 5 200 hours annually and has a design throughput of 100 t/hr. The 1 900 tonne plant feed stockpile will serve as the buffer between the crushing circuit and the remainder of the plant, which will operate on 7 000 hours annually.

Jigs.
-8mm material is reclaimed from the plant feed stockpile via FEL, a small plant feed bin and a belt feeder at nominally 52 t/hr and is conveyed to the rougher jig. The jig effects a density separation, with the less dense reject material gravitating to the tailings dewatering screen and the more dense concentrate material gravitating to the cleaner jig. The cleaner jig also effects a gravity separation and the tailings is pumped back to the rougher jig while the concentrate is gravitated to the concentrate dewatering screen.

Both the jig concentrate and tailings is dewatered on a screen with a cut size of 1mm, with the combined -1mm slurry being pumped to the low grade gravity concentration circuit. The -8mm+1 mm jig tailings, at nominally 37 t/hr, are conveyed to a tailings stockpile for reload by FEL onto trucks for disposal at the TSF and the -8mm+1mm jig concentrate is conveyed to the HG mill feed bin.

High Grade Gravity Circuit.
Jig concentrate is withdrawn from the bottom of the HG mill feed bin at a controlled rate, nominally 4 t/hr, by weigh feeder and fed into the primary peripheral discharge ball mill. The mill circuit includes a sizing screen and produces a product with a p80 of 425µm which is pumped to the HG gravity concentration circuit. The high grade (HG) circuit is arranged such that a rougher spiral produces a concentrate and tailings, with the concentrate feeding a shaking table and the tailings feeding a scavenger spiral. The rougher shaking table concentrate ("HG gravity concentrate") is pumped directly to the product magsep feed tank due to its low contaminants and high tin values, and the shaking table tails are combined with the rougher spiral tailings and pumped to the scavenger spiral.

The scavenger spiral produces a concentrate which is gravitated to another set of shaking tables, and a tail ("HG gravity tails") which is directed to the LG regrind circuit. The shaking table concentrate ("HG gravity middlings") is gravitated to the HG regrind mill and the tailings is recycled to the feed of the scavenger spiral.

LG Gravity Circuit.
The -1mm stream from the jig product and discard dewatering screens, nominally 13 t/hr, is pumped to a rougher spiral which produces a concentrate that is gravitated to a bank of shaking tables and a tailings which is pumped to a scavenger spiral. The scavenger spiral produces a tailings stream ("LG gravity tails") which is pumped to the tailings thickener and a concentrate stream which gravitates to the shaking table bank mentioned previously.

The low grade (LG) shaking tables produce a concentrate ("LG gravity concentrate") which reports to the HG regrind circuit, and a tails ("LG gravity middlings") which reports to the LG regrind mill.

Low Grade Regrind Gravity Circuit.
The HG gravity tails along with the LG gravity middlings report to a regrind ball mill and screen, which produces a product with a p80 of 106µm. The stream reports to a bank of shaking tables which produce a concentrate ("LG regrind concentrate") that reports to the sulphide float feed, and tailings ("LG regrind tailings") which is pumped to the tailings thickener.

High Grade Regrind Circuit.
The feed to the HG regrind circuit consists of HG gravity middlings and LG gravity concentrate. The mill circuit consists of a ball mill in conjunction with a screen and produces a product with a p80 of 106µm. The ground stream is then combined with the LG regrind concentrate and pumped to sulphide flotation.

Sulphide Flotation, Magnetic Separation and Filtration.
Product from the HG regrind mill along with the LG regrind concentrate reports, via a 2 hour buffer tank, to a bank of conventional sulphide rougher flotation cells. The sulphide minerals, which are a contaminant in the final concentrate, are floated off with the froth while the material which did not float is combined with the HG gravity concentrate and pumped to the product magnetic separator. The magnetic separator removes any free iron minerals which may still be present along with any iron added to the stream during the various during the various grinding stages. The magnetic separator tails are combined with the flotation froth and pumped to the tailings thickener while the magnetic separator concentrate reports to the product filter via a 2 hour buffer tank. The single concentrate stream is dewatered with a vacuum belt filter before being bagged and set aside for storage.