The majority of gold deposits on the West African Craton are orogenic-style deposits (Robertson and Peters, 2016). Orogenic-style deposits are sited proximal to major accretionary structures within, or at the boundaries of, composite metamorphosed volcanic- plutonic or sedimentary terrains. The majority of orogenic gold deposits occur within rocks that have been metamorphosed to greenschist facies, within a metamorphic pressure-temperature regime broadly corresponding to the brittle-ductile transition (McCuaig and Kerrich, 1998; Moritz, 2000).

The Markoye Fault is interpreted to be the main conduit for the gold mineralisation found at Sanbrado, accommodating large scale fluid migration. Sanbrado hosts shear zone type quartz vein gold mineralisation, similar to that found elsewhere in late Proterozoic Birimian terrains of West Africa. These orogenic type deposits exhibit strong relationships with regional arrays of major shear zones.

The structural environment can be described as brittle-ductile, coincident with a greenschist-amphibolite metamorphic facies. Arrays of transtensional features are apparent within a preferred S2 shear corridor and a strong shear controlled folding component is evident. Gold mineralisation is hosted in the sheared diorite and in the deformed mylonitic metasediment, and there is also a spatial relationship with the contact zones. It appears that the competency contrast caused by more brittle bodies of diorite within zones of more highly strained metasediment has played a role in partitioning the brittle deformation and dilation which hosts gold mineralisation.

Typically, the high-grade gold is observed within foliation parallel quartz lenses/veinlets often as short thin stringers of pure gold parallel to the foliation. In places gold can be seen along fractures in the quartz and in some cases the “vein material” that hosts the gold is observed to have relict rock fabric (i.e. foliation) indicating almost complete replacement of the protolith by silica.

Gold mineralisation is associated with the main hydrothermal event which produced strong silicification of the surrounding rock during opening and reactivation of the pre-existing foliation. Veinlets and lenses of quartz formed at this stage are mostly parallel to the foliation and can be observed as the siliceous bands in some of the banded mylonite. Subsequent local remobilisation would explain how much of the coarser free gold is in late fractures within the quartz.

This interpretation places gold mineralisation as post peak metamorphism after the bulk of the deformation, which means late during the regional D2 Birimian deformation within a roughly WNW-ESE (to NW-SE) stress field. Deformation and shearing along the high strain corridors has resulted in a pressure shadow, south of the main northern granitoid as the M1 and M5 high strain zones peel away (trending SE and SW respectively) from the same granitoid body. Conjugate movement along these two corridors - sinistral along M1 and dextral along M5 - is consistent with the late D2 stress field and has resulted in dilational opening and high grade steeply plunging ore shoots - along left-hand flexures at M1 and right-hand flexures at M5.

Late D3 deformation is orthogonal to D2 and reactivates D2 structures with an opposite sense of shear. While D3 reactivates, and in some cases offsets D2 structure and mineralisation, it is generally not associated with gold mineralisation, except possibly at M3, where the geometry of mineralisation is consistent with dextral shear across a north- west trending shear couple.

The mineralised structures occasionally display syn to post mineralisation displacement due to plunging fold noses and minor late faulting. Known mineralisation at M1 extends along strike for approximately 1km, is up to 50m wide and 550m in depth in M1S. The M5 mineralisation extends along strike for approximately 3km, is up to 200m wide and up to 300m in depth. The M3 mineralisation extends along strike for 750m, is up to 50m wide and 75m in depth. Mineralisation at all prospects remains open at depth.

The majority of the defined mineral resources at the Project are within 200m depth from the surface and of a lode style mineralisation. The material to be excavated will be predominantly free dig from surface with blasting required deeper in the oxidation profile. Given these conditions, conventional open pit mining techniques using drill and blast with material movement by hydraulic excavator and trucks will be employed. The project scale suits 120t – 200t class excavators in a backhoe configuration matched to 95t class mine haul trucks.

For this study a bench height of 5m has been used for all deposits. The benches will be excavated on 2 x 2.5m flitches in free dig material. The flitch height will increase to approximately 3m in blasted material when swell is considered.

In addition to the open pit resources, the steeply dipping, high grade zone at M1 South will be exploited by underground methods. The open pit has been limited to a depth of 120m from surface as optimisation of the mine schedule and project cashflow indicated that this provided the best interface with the underground mine. The underground mine will be accessed via a box cut and portal constructed 150m to the south west of the open pit.

The selected mining method is long hole open stoping, progressing upwards from the base of each lift. Depending on the geometry of the lodes two variations of this method are used. The majority of stoping is longitudinal open stoping. In some areas the 4 lodes have been combined together into larger stopes due to how close they are to each other. In this case transverse long hole open stoping is used. Stope filling uses a combination of cemented aggregate fill, cemented rock fill and development waste rock depending on whether or not the fill needs to be exposed to mine adjacent stopes.

The layout of the mine design for M1Sth was based on:
- Locating the portal in a separate boxcut adjacent to the M1Sth pit to commence in fresh rock. The portal is located in fresh rock and has 10 m of fresh rock above the portal.
- The deposit has been divided into four lifts each of 4 or 5 levels. Within these lifts stoping progresses upwards from the base of the lift. The final level in each lift is mined as a pillar recovery beneath the lift above, or the open pit in the case of lift 1. The crown between the underground mine and the open pit is mined after the completion other stopes.
- A level interval of 25 m was chosen to fit with the geotechnical conditions, the variable geometry of the orebody and to ensure maximum recovery of what is a valuable resource.
- Geotechnical analysis using the Mathews method has recommended the unsupported span on the hangingwall be limited to a hydraulic radius of <7 metres. For the 25m level interval this implies a strike length of approximately 25m. Where the stope strike length is designed greater than 25m then 10m long cable bolts will be installed in the hangingwall to ensure the unsupported span is within the recommendation of hydraulic radius < 7m.
- A single decline access based on a standard footwall 1:7 decline (5.5mW x 6.0mH) designed to accommodate 50t trucks is proposed.
- Ore drives of nominal size 5.0mW x 5.0mH are proposed, with widths adjusted locally to suit ore thickness.
- The initial exhaust ventilation will be through a 4.5m diameter raise bore from surface. This raise bore will also contain the initial stage of the second egress ladderway. The exhaust ventilation system will be extended from level to level with long hole raises (4.5m x 4.5m).

The main M5 pit is 2km long and averages 300m wide and 160m deep at the southern end. The pit has been designed so the southern higher grade portion can be mined independently of the northern portion of the pit. Both the northern and southern pit will be mined in two stages (an initial starter pit and then a cutback to final limits) in order to target higher grade earlier in the schedule and defer waste movement until later in the mine life.

The M1 deposit will be mined in two pits, a north and a south pit. The M1 South deposit will be mined by both an open pit and an underground mine exploiting the deeper ore. The open pit has been limited to a depth of 120m from surface as optimisation of the mine schedule and project cashflow indicated that this provided the best interface with the underground mine. The M1 South pit designs incorporates an underground portal in the south east wall. To enable the earliest possible development of the underground workings, the southern section of the pit will be mined to a depth of 30m to enable the portal to be developed in fresh rock. The final M1 South pit is 570m long by 290m wide by 120m deep. The M1 North pit is 350m long by 240m wide and 90m deep.

The primary aim of the production schedule is to provide the highest value ore to the mill as early as possible in order to maximise the value to the Project. This has been achieved by prioritising the underground mine and the highest grade open pits and pit stages. Additionally, low grade oxide ore from the M5 pit between the cutoff grade and 0.8g/t Au is stockpiled during the first 3 years of production to be treated in the final years of the mine plan.

Crushing
Haul trucks operating directly from the open pit mine will deliver run-of-mine (ROM) ore to the ROM pad. The gold bearing ore from the open pit mine will be stored on the ROM in separate stockpiles of varying ore types (oxidation stages) and grades to facilitate blending of the feed into the crushing plant. The estimated maximum particle size of ore on the ROM pad will be 800 millimetres in any dimension. Any oversized rock will be placed to one side and reduced to minus 800 millimetres on the ROM pad.

The primary crushing plant provides single stage crushing to feed the SAG mill. The primary crushing plant includes a 144 tonne capacity ROM Bin (121-BN-001), an apron feeder (121-FE-001), a vibrating grizzly (121-SC-001), a primary jaw crusher (121-CR-001), a stockpile feed conveyor (121-CV-001), a rock breaker(121-RB-001) and associated electrical equipment, steelwork and plate work.

The primary crusher will be installed on structural steel with the ROM bin, apron feeder and rock breaker, located adjacent to a 12 metre high concrete retaining wall against the ROM pad. Walkways and stairs will provide full operational and maintenance access throughout the primary crusher building.

The ROM bin will be fed blended ore from the ROM stockpiles using a front end loader (CAT 988 or equivalent). The ROM bin will be sized to accommodate direct tipping from 91 tonne sized haul trucks (CAT 777 or equivalent) but it is not expected that this will be the usual method of feeding the plant. The ROM bin will be lined with replaceable steel wear resistant liners . Feeding of the ROM bin will be controlled by a ‘dump – no dump’ traffic signal mounted adjacent to the ROM bin. The traffic signal will be controlled by a radar level sensor mounted above the ROM bin.

The ROM bin discharge will be controlled by a 1,500 millimetres wide apron feeder, which will direct feed into the primary jaw crusher. The primary jaw crusher will be a single toggle jaw crusher C130 with a feed opening of 1,300 x 1,000 millimetres that accepts nominally minus 800 millimetres rocks. The rock breaker mounted adjacent to the jaw crusher will break any oversized rocks that lodge in the crusher and would otherwise not be passed. The crusher jaw plates will be adjusted to a 110 millimetres closed side setting and will reduce the rock to a nominal minus 110 millimetres, although there will be some rocks larger than this that pass through the crusher. The material flow path will be conservatively sized to accommodate rocks to up to 350 millimetres in size.

The crushed product from the primary jaw crusher will discharge onto the 1,000 millimetres wide stockpile feed conveyor. A weightometer (121-WE-001) installed on the stockpile feed conveyor will provide information on the tonnage of crushed ore passing through the circuit and onto the stockpile. Dust control will be achieved using the dust suppression system (121-DC-01) consisting of fan forced dust collectors mounted in the discharge chute work of the primary crushing station.

Grinding
Primary crushed ore will be fed via 131-CV-002 to the SAG mill from the crushed ore stockpile. A 7.32 metre diameter x 4.27 metre long effective grinding length (EGL) SAG mill (132-ML-001) along with a 5.50 metre diameter x 7.85 metre long EGL Ball Mill (132- ML-002) make up the two stage grinding circuit. The SAG mill will operate with a duty ball charge of 7% (125mm balls at a maximum 15% ball charge) with an expected pinion power draw of 2.735 megawatt. The SAG mill will have an installed motor power of 4 megawatt. A variable speed drive (VSD) will be installed on the mill to vary the mill speed so that it can be adjusted if necessary for changes in ore characteristics. It should be noted that the VSD will be utilised to start the ball mill and once the ball mill is up to full speed, electric changeover will enable the SAG mill to be started and speed controlled. The Ball mill will operate with a duty ball charge of 28% (65mm balls at a maximum of 33% ball charge) with an expected pinion power draw of 3.396 megawatt. The Ball mill will have an installed motor power of 4 megawatt.

SAG mill slurry will discharge via a 2.4m diameter 3m long trommel. The SAG mill trommel (132-ZM-010) is installed to separate pebbles from the discharge slurry, control the particle size of the slurry reporting to the classification circuit and to adequately rinse the oversize ore, prior to it reporting to the pebble crushing circuit. The undersize slurry from the discharge screen will fall into the cyclone feed hopper(132-HP-001).

Cyclone feed pumps (132-PP-017 and 018) will be installed in a duty/standby arrangement, each having separate suction lines from the cyclone feed hopper. Pneumatically actuated suction inlet knife gate valves, pneumatically actuated knife gate dump valves on the pump suction pipework and pneumatically actuated knife gate valves on the discharge pipework will facilitate pump operations and maintenance of the off-duty pump when the system is operating the duty pump.

The slurry in the cyclone feed hopper will be pumped to a 15 pack classifying cyclone cluster (132-CS-001). The cluster will have 13 operating cyclones and two standby cyclones. The cyclones will classify the slurry feed into an overflow stream with a P 80 of 75 micron, which will be directed to the leaching circuit and coarse cyclone underflow stream. From the cyclone underflow, a portion of the slurry (approximately 100% of new feed) will report to the gravity circuit and the remainder fed back to the ball Mill (132-ML-002). The cyclone underflow will be split in the cyclone underflow boil box (132-BX-002).

New feed from the SAG mill feed conveyor will be added to the recirculating loads in the cyclone feed hopper (132-HP-001). The recirculating load has been designed to a nominal 400% of new mill feed with a 32% solids overflow. Proportional controllers will provide the mill operator with density control in the circuit by varying water addition to either the mill feed chute or cyclone feed hopper in fixed proportion to the SAG mill feed rate. The oversized material from the SAG mill trommel (132-ZM-010) will be directed either to a scats bunker or to the pebble crusher circuit. The pebble feed conveyor (136-CV-003) will deliver pebbles to a Terex TC 1000 Pebble Crusher (136-CR-001) at a design rate of 120tph with a design feed size of 75 millimetres. The crusher will operate at a close setting of 16 millimetres and the product from the pebble crusher will return to the SAG mill feed conveyor (131-CV-002).

A tramp metal magnet (136-MA-001) on the pebble feed conveyor and a metal detector (136-MD-001) will protect the pebble crusher from tramp metal. A diverter gate (136-ZM-003) at the end of the pebble feed conveyor will divert any tramp metal detected by the metal detector to bypass the crusher. The crusher station will also have an overflow to allow excess pebbles to recirculate to the SAG mill if the crusher is overloaded. The discharge from the Ball Mill will pass through the trommel screen (132-ZM-020) along with water addition, with the oversize reporting to the scats bunker and the undersize to the cyclone feed hopper.

The plant design proposed is simple, but robust, and broadly comprises the following:
1 Primary Crushing;
2 Crushed Ore Stockpile;
3 Grinding and Classification;
4 Gravity Recovery;
5 Leaching and Adsorption;
6 Elution;
7 Electrowinning; and
8 Smelting.

The gravity scalping screen (134-SC-001) will receive a portion of the cyclone underflow slurry. The gravity screen underflow will fall into the single centrifugal gravity concentrator (134-GC-001) which will remove concentrate into the Inline leach Reactor (ILR) storage feed hopper (134 HP-001). This hopper is part of the intensive leach reactor (134-ZM-001) which is located in a secure area adjacent to the milling building. The tailings from the gravity concentrator will be fed back into the cyclone feed hopper. The oversized stream from the scalping screen will be gravitate back to the ball mill for further grinding via the ball mill feed boil box (132-BX-003).

The ........