The Dasa Project is located in north-eastern Niger inside the Tim Mersoï sedimentary basin. The basin covers an area of some 114,000 km2 and is part of the much larger Iullemeden Basin (Palaeozoic-Tertiary) that stretches into Mali, Algeria, Benin and Nigeria.

All known uranium occurrences and deposits in Niger are all classified to belong to the sedimentary tabular, paleo channel and roll-front or sandstone types.

In the Dasa deposit, characteristics more consistent with the paleo channel tabular type seem to prevail.

The uranium in many of the deposits of the Tim Mersoï Basin is oxidized. Among the primary tetravalent minerals, coffinite is dominant and accompanied by pitchblende and silico titanates of uranium. Uranium hexavalent minerals such as uranophane and meta-tyuyamunite are present in the Imouraren and TGT-Geleli deposits.

The gangue is composed of quartz, feldspar, analcime and often illite, kaolinite and chlorite; with accessories such as some zircon, ilmenite, magnetite, tourmaline, garnet, anatase and leucoxene.

The uranium minerals are frequently associated with copper minerals (native copper, chalcocite, chalcopyrite, malachite, chrysocolla) and also with iron minerals such as pyrite, hematite and goethite. Organic plant materials are generally plentiful in un-oxidized facies of greyish-greenish colour.

The Dasa deposit site corresponds to a major structural intersection of the Adrar-Emoles flexure and the Azouza fault which resulted in the doming and creation of the Azouza Graben (Siebenthal, 2013). These are features that characterize other major uranium deposits in the Tim Mersoï Basin.

The geometry and the distribution of the uranium mineralization as seen in the Dasa drill core is to a large extent comparable with what has been reported from the uranium mines in the Arlit and Imouraren areas outside the Project:

There is a strong control by stratigraphy and lithology – with mineralization mainly hosted within the Tchirezrine 2 sandstones, particularly in the coarser-grained micro-conglomeratic facies of greyishgreenish colour containing frequent sulphides and organic matter such as plant remains.

The mineralized lenses are contained within northeast-southwest trending channels. The thickness of the mineralization may vary considerably between drillholes most likely an indication that channel stacking of favourable lithologies has increased the normal thickness of the sediment pile.

There are strong indications that the mineralization is influenced by a tectonic control along late northeast and southwest faults where some leaching has been observed.

Uranium mineralization is controlled by zones of oxidation – from surface (ground oxidation) and local/regional horizons at depth.

Groundwater circulation has created overtime discontinuities in the mineralisation as a result of tectonic movements.

Thin section work and petrographic studies by Activation Lab (2007) on Dasa samples has revealed that the uranium host rocks are sandstone and wacke which are variably oxidized. The main component is angular quartz, some plagioclase and lesser orthoclase. They are cemented by goethite, amorphous iron-hydroxides, and various secondary uranium-rich minerals.

The original cement between the grains of quartz and feldspar consisted of sericite and carbonate which were replaced during later stages by goethite and the amorphous iron-hydroxides. The quartz and the feldspar contain micro fractures partly filled with uranium-rich oxide. The latter also rim some of the silicates. Uranophane in form of radiating aggregates forms cement between the silicates and partly replaces them.

Global Atomic Corporation initiated a mineralogical study of the uranium mineralization on its property (Molebale, 2012) with five drill samples and five residue samples submitted for analysis. The samples were from drillholes ASDH 351, 353, 354(1), 354(2) and one DADH sample. The samples were split into representative portions and polished sections were prepared. Subsamples were pulverized for x-ray diffraction (XRD).

Five uranium-bearing minerals have been identified in Dasa samples (Molebale, 2012):
• Carnotite K2(UO2)2(VO4)2 x 3H2O;
• Uranophane Ca(UO2)2SiO3(OH)2 x 5H2O;
• U-rich titanite (U,Ca,Ce)(Ti,Fe)2O6;
• Coffinite U(SiO4)1-x(OH)4x;
• Torbernite Cu(UO2)2(PO4)2 x 11H2O;
• Autunite Ca(UO2)2(PO4)2 x 12H2O.

Majority of the mineralization is comprised of carnotite, uranophane and uranium-rich titanite and contribute to most of the uranium in the ASDH samples in terms of mass %, while torbernite is dominant in the DADH sample. The average grain size for the observed uranium-bearing minerals is -38 µm.

The source of the uranium is very likely leaching of the frequent volcanic tuff and ash blankets and intercalations now altered to analcimolite within the Wagadi and Dabla sediment packages. This has occurred over time in the geological history of the area and probably began as pre-uranium concentrations during the early sedimentation in favourable reducing environments such as organic matter-rich lower flow regimes and in favourable lithologies. The first stratiform mineralized bodies would have been formed during the early digenesis. Later, structural deformation and ground water movement within coarser grained organic-rich sediments aided by fluid movements and influenced by faults and tectonic activity, initiated roll front like redistribution of the uranium thus giving the mineralized bodies their present shape.

The Feasibility Study presented in this report presents an optimized mine plan and recovery process for the Dasa deposit based on the extraction 4.254 million tonnes of mineralised material from a sub-vertical section of the deposit on the flank of the graben, from an average depth of 300m below surface. The mine is designed to produce 1000 tonnes per day of Run of Mine (RoM) feed to the plant.

The selected mining method for the Dasa orebody is a transverse longhole open stoping (LHOS) method with a cemented hydraulic fill.

Access to the underground workings will be via a single access decline developed from surface. The decline system will be located in the footwall of the orebody.

Access to the orebody will be established at selected elevations by developing a crosscut from the decline towards the orebody, in a direction approximately perpendicular to the strike of the orebody. A footwall drive will be developed approximately 20m from the orebody, along strike. The stope access crosscuts spaced at 16.5 metre centres will be developed from the footwall drive across the orebody.

The level spacing is nominally 22.5 vertical metres. The general sequence of stoping will be bottom-up. Mining will commence on the lowest sublevel level in a mining block, progressing upwards towards a sill pillar, separating the mining block from the one above. Stopes will be filled with cemented hydraulic fill (CHF). The use of the cemented backfill will allow the mining of the sill pillars once the mining of a stope panel is completed.

The mine has been divided into 5 zones which are formed by areas of higher grade, which are targeted in the mine plan.

Mining will commence above the sill pillar on the 5275.5m elevation in Zone 1 and progress upwards which the development of the access to the bottom of Zone 1 and the other zones continues. Access to the top of Zone 5 is established via a haulage on the 5200 elevation. The haulage is twinned part of the way to provide an intake airway for Zones 4 and 5 and a return airway for Zone 5.

Standard mechanized underground mining equipment is proposed and will comprise electro-hydraulic long hole face drilling rigs and modern ground support drilling rigs. Proposed rock handling equipment will comprise diesel powered 14-tonne Load Haul Dump (LHD) units and 42-tonne articulated dump trucks.

Ancillary equipment will consist of diesel-powered charge-up vehicles, utility vehicles and other light vehicles such as Integrated Tool Carrier (ITC) units, man-carriers, Front End Loader (FEL), mobile rockbreaker, maintenance utility vehicle, and crane.

Primary Access
The primary access to the underground workings is via a single decline, developed from surface. The decline will facilitate truck hauling of rock, transport of personnel and material in and out of the mine as well as supplement the intake ventilation capacity.

In order to establish a safe mining face in hard rock to start the decline development a boxcut is excavated. The boxcut will be 13m deep and 31m wide at the high wall position with an inclination of minus 8 degrees. The ramp roadway is 5 metres wide and approximately 105 metres long, the roadway surface will be of a typical haul road design, with cast concrete drains running down either side of the boxcut excavation to interface with the cut-off drain that is positioned at the base of the highwall to direct water towards the collection sump. The top of the boxcut area will have a water retaining berm, approximately 3 metres wide at the base and 1 metre high, to lead the flow of excess run-off water away from the boxcut.

The portal area and decline will be sized to accommodate the trucks selected for ore and waste haulage, the selection of which is discussed in 16.15. The finished dimensions of the portal and decline will be 4.6m high x 5.1 m wide. To allow for a road base 30 cm thick, the excavation will be blasted at 4.6m wide x 5.4 m high. The decline will be developed at the same inclination as the boxcut roadway, an inclination of minus 8 degrees or 1 in 7.1 (14%).

The total length of the boxcut and decline system from surface to the lowest level of the mine is 4495 m.

Development
The decline will be located in the footwall of the orebody, at a minimum of 60 m from the footwall contact, to allow for the selected stand-off distance between the stopes and the footwall drive and adequate middling between the footwall drive, cubbies and the decline.

Access to the orebody will be established every 22.5 m vertically, by means of a level access crosscut. The level crosscuts will also be excavated at 4.6 m x 5.1 m. Between the decline and the footwall drive (strike drive) breakaway, a stub drive of 15 m length will be established. This will initially be used as muck-bays.

Stope crosscuts will be developed from the footwall drive at centre spacings of 16.5m and will be developed perpendicular to the strike direction intersecting the footwall contact of the orebody first but then advancing through the orebody to the hangingwall contact.

All development will be completed using mechanised drill, blast, support, load, and haul methods. Ore and waste will be hauled to surface by underground articulated dump truck (ADT), where it will be tipped on the RoM pad or wate rock dump. Once stoping commences and adequate voids are available underground for the tipping of waste into, waste will not be hauled to surface but will be placed in mined out stope voids as backfill.

The primary waste development which is made up of the ramps, level access crosscuts and footwall drives will all be 4.6 m x 5.1 m (finished dimensions), although the ramp is excavated 0.3 m higher to allow for a permanent road base to be established. Stope crosscuts are 4.5m x 4.5m.

All development drilling will be done by a twin boom electrohydraulic drill rig (Epiroc Boomer or equivalent). The blast-hole length will be 4.20 m, resulting in an expected advance per blast of 3.9 m. Holes will be charged with bulk emulsion. The initiation system will be shock-tube long period detonators (LPDs).

Stoping
Once the lateral development on a level is completed and the upcast ventilation raises linking to the level above have been developed, the level is ready for the stope cross cuts to be developed, after which stoping can commence. Stope preparation starts with the development of a slot raise at the end of the stope.

The slot raise will be mined at 2.0 m x 2.0 m and will hole between sublevels. Slot raises will be mined using drop raising techniques. The drilling of the holes will be completed with the long-hole production rig. Once the slot raise is completed, the raise will be opened up to the full width of the stope, to form a complete slot, using a pattern of fanned blast holes. Once this slot is complete the stope is ready for production stoping to commence.

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 at a time, immediately prior to blasting.

The sequence of mining is to retreat from the ore block extent, back towards the stope access. After the stope has advanced to the orebody limit or a maximum distance of 90 m from the initial slot raise, backfilling must take place and the backfill must cure, before re-slotting and recommencing the stoping operation. The backfill will be deposited into the stope from the stope access crosscut on the level above.

The sequence of mining is bottom-up from a main level. Conventionally, a permanent sill pillar would be left below the main level in this type of mining. In this case the sill pillar will be mined in a similar manner to a normal level, once all mining in the mining panel is completed.

Crushing Circuit
The primary crushing circuit consists of a single tip with a dedicated ROM bin and a single jaw crusher in open circuit. Primary crusher product reports to the crushed ore stockpile (COS). The ROM ore (F100 500 mm) is loaded into a 70 t ROM bin by means of a front-end loader (FEL), or by direct tipping by haul trucks.

A static grizzly will be fitted to the ROM bin to protect the downstream equipment from oversize material and will be inclined for easy removal of oversize to minimise stoppages.

The ROM ore is drawn from the ROM bin at a controlled rate by a single, variable speed vibrating grizzly feeder, and fed directly to the jaw crusher. The speed of the vibrating grizzly feeder is controlled to maintain crusher throughput. Undersize material from the vibrating grizzly feeder reports to the primary crushing conveyor, where it is combined with the primary crusher product (P100 250 mm, P80 80 mm). The primary crushing conveyor is fitted with a belt magnet to remove any tramp iron material. The primary crushing conveyor discharges the crushed material onto the COS.

Dry Grinding and Classification
The milling circuit is configured as a dry SAG mill in closed circuit with dry screens. Ore is withdrawn from the COS by vibrating feeders feeding onto the SAG mill feed conveyor. A weightometer indicates the instantaneous and totalised crushed ore mill feed tonnage and is used to control the SAG mill feed rate via the vibrating feeders. The SAG mill feed conveyor discharges directly into the SAG mill feed hopper. The SAG mill discharge is conveyed to the sizing screens via a bucket elevator conveyor system.

The crushed ore will be ground dry in an air heated semi-autogenous (SAG) mill to a product P95 600 µm. The SAG mill is in closed circuit with dry sizing screens. The screened ore oversize (>600 µm) is recycled to the mill. The < 600 µm material from the screens, together with dust recovered from a bag filter system (dust generated during the screening process and the SAG mill), is sent to the leaching process.

The SAG mill is equipped with a variable speed drive to assist with managing variations in feed hardness. When treating soft ore, the SAG mill will be operated at a lower speed and reduced ball charge to minimise overgrinding of the low competency ore. When treating hard ore, the ball mill speed and ball charge will be increased to apply more power to mill the competent ore.

The air heating and drying system feeding the SAG mill will use steam from the acid plant to preheat the air with diesel fed burners providing the final heating requirements. A ball loading system is used for loading of grinding media into the SAG mill (via the mill feed conveyor).

The Dasa Plant has a capacity to treat 365,000 t/a of run-of-mine (ROM) at a head grade of 0.45% U3O8.

Ore delivered to the crusher is reduced in size and delivered to a mill feed stockpile by conveyor belt – the stockpile providing a buffer capacity to ensure feed to the mill is constant.

Reclaim feeders under the stockpile feed ore via a conveyor belt to the SAG Mill, where the ore is dried and milled – the milling process includes screening and ore recirculating back to the SAG Mill until a final product size of 0.6 mm is attained. Ore discharging from the milling process reports to a revolving pug drum where sulphuric acid and nitric acid are introduced to start the uranium extraction process. The retention time in the pugging drum is approximately 15 minutes before the ore discharges and reports via a feed conveyor to the curing conveyor belt.

Ore discharges from the curing belt and is fed to the leach section where it discharges into re-pulping ta ........