TECHNICAL FIELD

The present disclosure relates to a method for processing ore. In particular, the present disclosure relates to an integrated method for processing copper and uranium containing ore which includes a heap leach process and an oxidative leach process. The present disclosure also relates to a plant for conducting the integrated method.

BACKGROUND ART

Mineral deposits in the Stuart shelf, Australia, generally contain three metal values, namely copper, uranium and precious metals (mainly gold and silver). The copper is principally in the form of sulphide minerals such as Chalcocite (Cu ₂ S), Bornite

(Cu ₅ FeS ₄ ) and Chalcopyrite (CuFeS ₂ ). The uranium is principally in the form of such minerals as Uraninite, Coffmite and Brannerite. The common gangue minerals may include quartz, hematite, feldspar, sericite, fluorite, siderite, chlorite and pyrite.

If the copper, uranium and gold content in the ore body is sufficiently high to be economically recovered then the treatment of the ore is complex as historically each value mineral requires within reason a separate processing route. A typical process employed to treat this ore type may contain the following processing steps:

1. Milling and flotation:

The run of mine ore is milled and reports to a sulphide flotation plant, which recovers approximately 93% of the copper sulphides and 70% of the gold to sulphide

concentrate. This concentrate accounts for approximately 3% of the total ore mass.

2. Concentrate pre-treatment

The concentrate is thickened and leached at elevated temperatures in a series of stirred tanks with sulphuric acid to solubilize fluoride and uranium as well dissolve carbonates. The resultant thickener underflow is filtered. The thickener decant as well as filtrates report to a tails leaching facility. The filter cake is re-pulped, neutralized, re-filtered and dried in preparation for smelting.

3. Smelting and Anode production

The concentrate is fed to a flash smelter along with oxygen and fluxing agents. The concentrate separates into a blister copper and slag phase. The blister copper is fire refined in an anode furnace and then cast into anodes.

4. Electro refin ing and precious metal recovery

The cast copper anodes are electro refined onto blank cathode sheets. The cathodes are removed, stripped and sold. Gold and silver from the concentrate report to the anodes and collect in the tankhouse as slimes. The slimes are treated with cyanide to recover gold and silver.

5. Flotation tails leaching and uranium recovery

The flotation tails contains the majority of the uranium. The flotation tails are thickened and leached at elevated temperatures in a series of stirred tanks with sulphuric acid and sodium chlorate to oxidise and solubilize uranium. The leached tails slurry undergoes counter current decantation. The decant is clarified and the clarified liquor, now called pregnant leach solution, reports to a copper solvent extraction facility (CuSX) to extract copper. The CuSX raffmate reports to a uranium solvent extraction facility where uranium is precipitated from solution as ammonium diuranate. The ammonium diuranate is calcined to form uranium oxide and packaged for sale.

The above type of process has a number of disadvantages when applied to this type of orebody:

• Robustness: the three major processing units 1, 3 and 5 are linked. Tails

leaching (unit 5) and smelting (unit 3) require the mill and flotation (unit 1) to operate in order to provide feed. Also, the Tails leach (unit 5) requires smelting (unit 3) to operate for acid supply. In addition, the mill and flotation (unit 1) requires tails leach (unit 5) to operate as otherwise the flotation tails are too voluminous to store. In general, the short process retention times dictates that stoppage of one processing unit requires stoppages of all units within a short time period. This in turn may reduce overall plant availability and continuity. Lack of expandability. These types of processing units are limited by the size of unit operations such as the counter current decantation circuit. This means that using this technology to treat large tonnages requires multiple parallel processing plants.

Smelting operations: Producing a high grade concentrate for smelting or export is at the expense of copper and gold recovery. The flash smelter energy balance limits the range of concentrate that can be treated and the smelting process would need modification to accommodate the change in mineralogy of the ore deposit over the life of the mine. Hence concentrates containing high proportions of chalcopyrite or pyrite would require a second stage of smelting to be added to the existing process. Moreover, smelting has health and hygiene risks because of the concentration of radio nuclides in certain smelter streams as well as because of the high temperature operation.

Tails leaching: Fluorite and feldspar dissolve in the tails leach tanks and respectively release calcium and potassium into solution. These ions re- precipitate as gypsum and jarosite scaling. The tails leach reactors each have a short individual retention time (~2hours) which is insufficient time for seed nucleation within the slurry. This results in the slurries remaining super saturated in potassium and calcium, meaning scale formation throughout the tails leach, CCD and solvent extraction facilities. This necessitates maintenance which can affect plant availability and efficiency. In addition, elevated chlorite concentrations in the ore can create spikes in the silica concentration which may polymerize into silica gel and bog the process. Furthermore, uranium dissolution during tails leaching is initially rapid but then has a very slow leaching tail. The vast majority of the uranium may be recovered using stirred reactors but this may not be cost effective. It would therefore be desirable to provide a method for processing ore that overcomes or at least alleviates one or more disadvantages of the existing process flowsheet. It would also be desirable to provide a method for processing ore which yielded higher total metal recoveries than the existing process flowsheet.

The above references to the background art do not constitute an admission that the art forms a part of the common general knowledge of a person of ordinary skill in the art. The above references are also not intended to limit the application of the processes and plant as disclosed herein.

SUMMARY OF THE DISCLOSURE

In a first aspect there is disclosed a process for recovering copper, uranium and one or more precious metals from an ore material, including:

a. forming a heap of the ore material;

b. subjecting the heap of the ore material to an acidic heap leach using an iron containing acidic leach solution in the presence of an oxygen containing gas, and producing a first pregnant leach solution and a ripios;

c. subjecting the ripios to flotation to produce a copper concentrate and tailings; d. subjecting the copper concentrate to an oxidative acid leach to produce a second pregnant leach solution and a leach residue;

e. treating at least the first pregnant leach solution in an extraction step to extract uranium and copper; and

f. treating the leach residue from step (d) to recover the one or more precious metals.

The relatively long response time of the heap leach step enables it to be effectively decoupled from the flotation step. This means that suspending or stopping the flotation step does not require suspending or stopping the heap leach step, thereby enhancing overall plant availability and continuity. Moreover, the concentrate from flotation is a relatively small flow and can be given a reasonable storage time and hence a reasonable surge capacity between the steps is possible. The second pregnant leach solution (PLS) arising from the oxidative acid leach in step (d) may be recycled for use as at least part of the iron containing acidic leach solution in the heap leach step (b). Alternatively, the second PLS may be directly treated in extraction step (e). The use of a single extraction step (e) for one or more pregnant leach solutions means that the copper recovery split between the heap leach and oxidative acid leach stages can be varied if required with little or no impact upon the operation of this stage.

In one embodiment of the process, the ore material contains copper sulphides and uranium minerals. The ore material may also contain one or more iron containing minerals. The iron containing minerals may include gangue minerals, such as one or more of hematite, siderite and chlorite. These minerals may partly or wholly comprise the source of the iron in the iron containing acidic leach solution.

In one embodiment of the process, the oxygen containing gas is air. In another embodiment, the oxygen containing gas is oxygen-enriched air.

In one embodiment of the process, each of the first and second pregnant leach solutions contains copper and uranium. In another embodiment, the second pregnant leach solution predominantly contains copper.

In one embodiment of the process, the iron containing acidic leach solution contains ferric ions which oxidise the ore material to dissolve copper and uranium, resulting in reduction to ferrous ions. The ferrous ions are then reoxidised to ferric ions by reaction with the oxygen containing gas.

The extraction step (e) may comprise a solvent extraction stage. The solvent extraction stage may comprise two solvent extraction processes, one to extract copper and the other to extract uranium from the pregnant leach solution. The extraction step may comprise a first solvent extraction process to recover copper from the pregnant leach solution/s, and the raffinate from the first solvent extraction process is then treated in a second solvent extraction process to extract uranium. The raffinate from the second solvent extraction process may be at least partly recycled for use in the heap leach step (b) and/or at least partly purged in order to control the build up of deleterious species (eg chloride and/or excessive iron) in solution.

5 In an embodiment the extraction step may include an ion exchange process to extract uranium from the raffmate from the first solvent extraction process.

The acidic heap leach in step (b) may be conducted in more than one stage. In one embodiment, the acidic heap leach is conducted in two stages. In another embodiment, l o the acidic heap leach is conducted in more than two stages, such as in three stages.

Where more than one heap leach stage is employed, the ore being leached in a subsequent stage may comprise partially leached ore from a previous stage. Moreover, the first heap leach stage may result in an intermediate leachate which may at least partially comprise the leachant for the second heap leach stage. The leachate from the

15 second heap leach stage may at least partially comprise the leachant for a third heap leach stage (if used) and so on.

The leachate from at least the second heap leach stage may comprise the first pregnant leach solution.

20

A portion of the first pregnant leach solution may pass to the extraction step and a portion may be recycled. This is done to minimise the volume of the first pregnant leach solution required to be processed in the extraction step and to therefore control the size of extraction equipment.

25

In one embodiment, the first and second pregnant leach solutions may be each directly treated in the extraction step. This may be done by providing a single (common) extraction step to service both the leach stages (b) and (d), leading to greater process and cost efficiency.

30 In another embodiment, the second pregnant leach solution is recycled for use as a leachant for the heap leach step (b) and only the first pregnant leach solution is subjected to extraction to recover copper and uranium.

The ripios produced in step (b) may be rinsed with water or recycled process water, such as from the flotation step, prior to step (c). Rinsing removes acid and soluble metals. The rinse water may then be used as a diluent such as in the heap leach step (b) (eg in the final heap leach stage) and/or in the oxidative acid leach of step (d). The use of heap leach derived rinse water as a diluent for the oxidative acid leach step, and its subsequent return to the heap leach step in the second PLS, assists to optimise the water conservation of the metal recovery process.

The ripios produced in step (b) is preferably milled prior to the flotation step (c), in order to maximise separation of ore minerals from gangue.

The copper concentrate produced in step (c) may comprise copper containing sulphides, uranium minerals and one or more precious metals. The precious metal may be gold and/or silver.

Prior to the oxidative acid leach step (also known as "direct oxidative leach"), the concentrate may be subjected to milling or further milling to reduce the particle size distribution. The finer particle size distribution may comprise P80 of 105 microns or below, such as a P80 of 75 microns or below. In an embodiment, milling is conducted to a particle size of P80 of 35 microns or below. It has been found that ultrafine grinding of the concentrate increases the reactivity of the concentrate to oxidative leaching by virtue of the increased surface area.

The oxidative acid leach of step (d) may be conducted at atmospheric pressure and at a temperature of up to the boiling point of the leach solution. In an embodiment, the temperature may be from 60 to 90° C. In another embodiment, the oxidative acid leach of step (d) is conducted in an autoclave at an elevated temperature and a pressure above atmospheric pressure. The temperature may be a maximum of 230 ° C, such as from 160 ⁰ C to 220 °C. the pressure may be a maximum of 35 bar, such as from 10 to 24 bar(a) .

5

The oxidative acid leach of step (d) may be conducted by treating the copper concentrate in one or more reactors with an acidic solution in the presence of oxygen gas to dissolve copper and uranium into the second pregnant leach solution leaving a leach residue containing the precious metal.. The copper sulphides and uranium

o minerals in the copper concentrate are oxidised by ferric ions in the acidic solution resulting in reduction of ferric ions to ferrous ions. The resulting ferrous ions are reoxidised to ferric ions by dissolved oxygen. The addition of the oxygen containing gas may conveniently be via a sparging device. The oxygen containing gas may be introduced into the reactor below an agitator, preferably a high solidity agitator, in order5 to better distribute dissolved oxygen in the reactor. The measured dissolved oxygen

(DO) in solution may be controlled at an optimum value, to optimize oxygen utilization and redox potential. In one embodiment, the optimum value of DO concentration is controlled by varying the flow rate of the oxygen containing gas introduced into said solution. However, alternative ways of controlling DO including adjusting the partial 0 pressure of oxygen in the oxygen containing gas or adjusting the amount of agitation of said solution, in particular the amount of power transferred to the solution by a motor driven agitator.

The oxidative acid leach of step (d) may be conducted in two or more reactors arranged 5 in series. The retention time of the oxidative acid leach in the first reactor may be such that the oxidative acid leach is autothermal.

In an embodiment, the oxidative acid leach of step (d) may be conducted in one or more autoclaves. Each autoclave may have two or more compartments. A quench solution, o such as comprising rinsewater from another stage in the process (eg, from rinsing

ripios), may be used to control the temperature in the autoclave. Oxygen overpressure is also preferably controlled. The precious metal recovery step (f) may comprise cyaniding a slurry of the leach residue and subjecting the slurry to a carbon in pulp or a carbon in leach process.

In a second aspect there is disclosed a process for recovering copper and uranium from an ore material, including:

a) forming a heap of the ore material;

b) subjecting the ore material to heap leaching using an iron containing acidic leach solution in the presence of an oxygen containing gas, and producing a first pregnant leach solution and a ripios;

c) rinsing the ripios to produce a rinsed ripios and rinsewater;

d) milling the rinsed ripios to a particle size of P80 of 105 microns or lower; e) subjecting the milled ripios to flotation to produce a copper gold

concentrate and tailings;

f) diluting the concentrate (if necessary) to form a slurry having a solids content of 5 - 20% w/w;

g) subjecting the concentrate slurry to an oxidative acid leach to produce a second pregnant leach solution;

h) optionally using the rinse water from step (c) as a diluent in the heap leach step (b) and/or in the oxidative acid leach of step (g);

i) recycling the second pregnant leach solution for use as a leachant for the heap leach step (b);

j) subjecting the first pregnant leach solution to solvent extraction to recover copper and uranium; and

k) purging a portion of the raffinate resulting from solvent extraction to

control the build up of deleterious species such as chloride.

In a third aspect there is disclosed an integrated plant for use in a method for recovery of copper and uranium from an ore material as disclosed above, including:

means for forming a heap of the ore material;

means for supplying an acidic leach solution and an oxygen containing gas to the heap to form a first pregnant leach solution and a ripios; extraction equipment for extracting copper and uranium from the first pregnant leach solution;

means for collecting and transferring the first pregnant leach solution arising from the heap leach to the extraction equipment;

flotation equipment for use in flotation of the ripios to produce a copper concentrate and means to transfer the ripios from the heap to the flotation equipment; one or more reactors for conducting an acidic oxidative leach of the copper concentrate to produce a second pregnant leach solution; and

means for collecting and transferring the second pregnant leach solution arising from the acidic oxidative leach to one or more of the heap or the extraction equipment.

Each aspect of the process for recovering copper and uranium may be conducted using saline water (such as sea water or saline ground water) for the process solutions, as opposed to fresh water. This may be beneficial in regions with water scarcity, or where water conservation is important. In addition, the presence of chloride ions in the process solutions may improve the leach rates of ore minerals, such as chalcopyrite.

BRIEF DESCRIPTION OF THE DRAWINGS

Notwithstanding any other forms which may fall within the scope of the process as set forth in the Summaiy, specific embodiments will now be described, by way of example only, with reference to the accompanying drawings in which:

Figure 1 shows a process block diagram illustrating the general unit operations for an embodiment of a process for recovering copper, uranium and one or more precious metals from an ore material;

Figure 2 shows a first embodiment of a flowsheet for a process for recovering copper, uranium and one or more precious metals from an ore material;

Figure 3 shows a second embodiment of a flowsheet for a process for recovering copper, uranium and one or more precious metals from an ore material; Figure 4 shows a third embodiment of a flowsheet for a process for recovering copper, uranium and one or more precious metals from an ore material; and

Figure 5 shows a fourth embodiment of a flowsheet for a process for recovering copper, uranium and one or more precious metals from an ore material.

DETAILED DESCRIPTION OF SPECIFIC EMBODIMENTS

In each of Figures 1 to 5, like reference numerals are used to denote similar or like parts.

Referring firstly to Figure 1, shown is a process block diagram 10 illustrating the general unit operations for an embodiment of a process for recovering copper, uranium and one or more precious metals from an ore material.

As-mined ore 12, comprising copper sulphides, uranium minerals and gangue minerals (siderite, hematite and chlorite), is delivered to an ore stock pile, then subjected to a crushing step 14. The crushing step 14 may comprise one or more stages of crushing, such as up to three stages of crushing.

The crushed ore 15 reports to an agglomeration step 16, where it is contacted with raffinate and/or acid in a drum to bind the fine particles of ore to the larger lumps.

The agglomerated ore 17 reports to a heap leach step 20. The ore is stacked in one or more heaps to a height of approximately 6 to 10m by a stacker (not shown), such as a moving bridge stacker. The ore is stacked on a lined re-usable pad (not shown) that has a drainage layer, drain pipes and aeration pipes, and is irrigated through a dripper system (also not shown) with acidified liquor 30. The heap/s are supplied with air 32 blown into the heap. During the heap leach step 20, the gangue minerals siderite and chlorite are leached with the acidic liquor 30 and they release ferrous iron into solution. This ferrous iron converts to ferric iron in the presence of the oxygen supplied by the air 32 blown into the heap. The gangue mineral Hematite also dissolves and releases ferric iron into solution. The ferric ions leach the copper sulphide minerals, (eg, chalcopyrite, bornite and chalcocite) liberating copper and more ferrous iron. The liquor is recirculated and the ferrous iron is re-oxidised to ferric ion by oxygen from the air. The ferric iron and acid also leach uranium into solution from the uranium minerals in the ore. The heap leach step 20 produces a first pregnant leach solution 40 containing dissolved copper and uranium and a ripios 42.

The resulting ripios 42 reports to a milling step 50. In this step, the ripios is milled in either a ball or pebble mill. Make up water 51 may be added if required.

The milled ripios 60 reports to a flotation step 62 which recovers approximately 93 - 95% of the copper sulphides and 70 - 73% of the precious metals to the sulphide concentrate 68. This concentrate 68 accounts for approximately 1-2% of the total ore mass. The concentrate 68 is thickened in concentrate thickening step 70 to about 40% - 50% solids. Excess water 72 is returned to the milling step 50.

The flotation tailings 74 are thickened to about 68-70% solids in the flotation tails thickening step 76. The decanted water 78 from the flotation tailings thickening step 76 is returned to the milling step 50. The underflow 80 from the flotation tails thickening step 76 reports to the tails storage facility 82.

The thickened concentrate 84 resulting from thickening step 70 may be re-ground to a finer particle size distribution before reporting to an oxidative acid leach stage 86 which may comprise multiple autoclaves or a series of stirred tanks operating at atmospheric pressure. The finer particle size distribution may comprise P80 of 105 microns or below, such as 75 microns or below. In an embodiment, the particle size distribution has a P80 of 35 microns or below. Oxygen containing gas 88 is injected into a slurry of the (optionally milled) concentrate 84 below high solidity gas transfer type agitators (not shown). In the oxidative acid leach stage 86, whether operated under pressurized, atmospheric, chloride based or bacterially catalysed conditions, copper sulphides are fully or partially oxidized which liberates copper into solution.

The oxidised slurry 89 from the oxidative acid leach stage 86 reports to a thickening and filtration step 90. The decant from the thickening/filtration step 90, comprising a second PLS 92, is recycled to either the extraction step 52 or to the acidic heap leach step 20 (not shown).

The first pregnant leach solution 40 and (optionally) the second pregnant leach solution 92 report to an extraction step 52 in which copper and uranium are extracted from solution. The extraction step 52 reclaims copper that is dissolved during the heap leaching step 20 and the oxidative leach step 86 of the concentrate.

Copper is extracted using solvent extraction and the copper depleted aqueous steam (copper SX raffinate) reports to a uranium solvent extraction facility. The copper loaded organic is scrubbed, stripped and copper is electrowon onto cathode plates. The uranium bearing organic is likewise first scrubbed and then the uranium stripped from the organic. The uranium is then precipitated from solution as ammonium diuranate, then calcined to forai uranium oxide and packaged for sale.

The raffinate from the uranium solvent extraction stage is recycled back to the heap. A portion of this raffinate is purged and reports to the tails storage facility.

The solid residue 94 from thickening/filtration step 90 reports to the cyanidation section 96 (which is more often referred to as Carbon in Leach (CIL)) to which lime and cyanide 97 are added for recovery of precious metals gold and silver 98.

Figures 2 to 5 are, respectively, first, second, third and fourth flowsheet embodiments showing further details of the unit operations illustrated generally in Figure 1. The following discussion will focus on those features that differ from corresponding features in the other embodiments.

Referring to Figure 2, there is shown a first embodiment of a flowsheet 1 10 for a process of recovering copper, uranium and one or more precious metals from an ore material. Agglomerated ore 117 reports to a heap leach step 120, in which the ore 117 is stacked in one or more heaps. The heap leach step 120 includes a stacking/curing step (not shown), followed by first, second and third (final) stages of leaching, 122a, 122b and 122c, respectively. In each of the second and subsequent stages, the ore being leached has been partially leached in a preceding stage. While three leach stages are shown, it should be appreciated that fewer or greater leach stages could instead be employed. A final rinse stage 123 removes acid and soluble metals from the final leached ore (ripios) 142.

Acidic leach liquor is irrigated over the first leach stage 122a and the resulting intermediate leachate 124 reports to a first intermediate leach solution pond 125. Acid (not shown) is preferably added to intermediate leachate 124 and the acidified leachate 124a is then pumped over the second leach stage 122b. The second leach stage 122b produces a first pregnant leach solution 140 which reports to a PLS pond 126. Because the first pregnant leach solution 140 has passed over partially leached ore (in the second leach stage 122b), it is therefore more predictable in terms of acidity and metals tenors.

A portion 140a of the first pregnant leach solution 140 reports to an extraction step 152, in which copper and uranium are extracted from solution, and another portion 140b is by- passed to a second intermediate liquor pond 127. The liquor from the second intermediate liquor pond 127 is recycled as leachant for the first leach stage 122a.

The extraction step 152 includes a copper solvent extraction/ electrowinning stage (CuSXEW) 153 and a uranium solvent extraction and refining stage (USX) 154.

In the CuSXEW stage 153, the first pregnant leach solution 140a is contacted in a counter current flow with an organic phase that loads copper to produce a loaded organic and a copper depleted aqueous steam (the copper SX raffinate). The loaded organic is scrubbed and then the copper is stripped off by a strong acid liquor (eg, spent electrolyte). The loaded strip liquor reports to an electro-winning facility where the aqueous copper is plated on to blank cathode plates. The copper is stripped from the plates and sold. The spent electrolyte and the organic phase are both recycled.

The copper raffinate 156 reports to a uranium solvent extraction and refining facility 154. This is mostly analogous to the CuSXEW stage 153, with the exception that stripping is carried out by aqueous ammonia and the uranium is precipitated from solution as ammonium diuranate. The ammonium diuranate is calcined to form uranium oxide and packaged for sale. The uranium and copper raffinate (uranium and copper depleted liquor) 157 reports to the raffinate pond 158.

The above described splitting of the first pregnant leach solution 140 into streams 140a and 140b ensures that the volume of the portion of first pregnant leach solution 140a going to the extraction step 152 is reduced, thus reducing the required size of the extraction facility in the extraction step 152. It is based around optimizing copper tenors of solution in the CuSXEW stage 153to as well as maintaining a fixed irrigation flux rate (l/m ² /h) over the heap leach step 120.

A purge of liquor 159a is bled from the raffinate pond 158 to control chloride tenors in the circuit.

A raffinate stream 159b is recycled to the final heap leach stage 122c. This is beneficial as it means that the final heap leach stage 122c is irrigated with a low copper and uranium containing solution prior to rinsing and removal. This minimizes losses of soluble Cu and U through the subsequent milling step 150. The raffinate stream 159b percolates through the third heap leach stage 122c, producing a second intermediate leach solution 161 that reports to the second intermediate liquor pond 127. The acid tenor is adjusted if required by the addition of sulphuric acid and a stream of the leach solution 163 is pumped to the first heap leach stage 122a. The ripios 142 is rinsed with water 164 prior to being offloaded. The volume of rinse water 164 is set to match the water losses through evaporation and the raffinate bleed but offset by the influent water streams from scrubbing and other sources. After rinsing, the rinse water 167 reports to rinsewater pond 165.

The rinsed ripios 142 is reclaimed by most likely a bridge mounted bucket wheel excavator (not shown) and transferred by conveyors (also not shown) to a milling step 150. The ripios is milled in either ball or pebble mills. Make up water is added.

The milled ripios 160 reports to a sulphide flotation plant 162. Froth flotation is used to generate a rougher concentrate. The rougher concentrate can either be used as is, cleaned further, or re-ground to improve mineral liberation. Regrinding may be conducted in either a ball mill, a stirred mill or a tower mill. The ground concentrate is then passed through cleaner flotation cells to improve the grade, but at the expense of slight gold and copper loss. The sulphide flotation plant 162 produces a concentrate 168 and tailings 174. Approximately 93 - 95% of the copper sulphides and 70 - 73% of the gold is recovered to the concentrate 168. This concentrate accounts for approximately 1-2% of the total ore mass. The concentrate 168 reports to a concentrate thickener 170 and is thickened to 40%- 50% solids. The excess water 172 is returned to the milling section 150.

The flotation tailings 174 report to one or more tailings thickener/s 176 and are thickened to 68-70% solids. The decanted water 178 from the flotation tailings thickener(s) 176 is returned to the milling section 150. The underflow 180 from the flotation tailings thickener 176 is diluted with the heap leach raffinate purge 159a and pumped to the tails storage facility 182.

The concentrate 184 reporting from the concentrate thickener 170 underflow may be re- ground to a finer grind size in mill 185, such as to a P80 of 40 microns or less, in order to increase the reactivity of the concentrate to oxidative leaching by virtue of the increased surface area availability. The ground concentrate 184a is slurried using water from the rinsewater pond 165. The dilute concentrate slurry 184a reports to the oxidative acid leach step 186, which comprises a series of stirred tanks 187a, 187b, 187c. The sulphide in the concentrate 184 is oxidized by the injection into the slurry of oxygen containing gas 188 below high solidity gas transfer type agitators 191. The temperature of the oxidative acid leach step 186 is maintained by the addition of steam if necessary and/or cooling. Cooling could be effected by in- tank pipes carrying a circulating cold liquor (eg water). The heat can be removed from this cold liquor by exchange with heap leach PLS liquor 140 or through the use of cooling towers (not shown). Alternatively, cooling can be effected by extra dilution (eg, cold shot cooling). The oxidative leach step may be carried out in accordance with the disclosure in Batty J.D & Rorke G.V., 2005, "Development and commercial demonstration of the BioCOP™ thermophile process", S.T.L. Harrison, D.E. Rawlings, J. Petersen (Eds.), Proceedings of the 16th International

Biohydrometallurgy Symposium, Cape Town. South Africa (2005), pp. 153-161, the entire disclosure of which is incorporated herein by reference.

An important consideration in the oxidative acid leach step 186 is the ratio between the retention time in the first stirred tank reactor 187a (or multiple reactors in parallel) and the rest of the tank reactors (187b, 187c) in the leach circuit. This is important for the following reasons: the reactions occurring in the tanks can be autocatalytic in nature and therefore optimizing the retention time ratio will minimize the total number of tanks required for a specific oxidative acid leach.

A short retention time in the first tank 187a (or multiple reactors in parallel) would result in steam being required to get to operating temperature. The energy inputed through the steam addition has to be removed by a commensurate increase in cooling water. It is preferred that the retention time in the first tank 187a is sufficient so that the oxidative acid leach step 186 may operate autothermally. This minimizes the need for cooling and removes the process reliance on steam availability except during start up.

A further benefit of relatively long retention time in the first stirred tank reactor 187a (or multiple reactors in parallel) is that sufficient time is provided for the formation and maintenance of gypsum and jarosite seed inventory. Because these minerals provide nucleation sites for future growth this minimizes scale formation on reactor and pipe walls.

The oxygen transfer equipment for use in the injection into the slurry of oxygen containing gas 188, and the control of the equipment, may be as disclosed in

PCT/ZA06/000108, the entire disclosure of which is incorporated herein by reference.

In the oxidative acid leach step 186, whether pressurized, atmospheric, chloride based or bacterially catalyzed, copper sulphides are fully or partially oxidized which liberates copper into solution.

The iron in solution reporting to oxidative acid leach step 186 will play an important role in optimizing the circuit. The conversion from ferrous to ferric (the required oxidant) is a function of ferrous ions squared multiplied by the dissolved oxygen.

Operation at higher concentration of iron will allow equivalent rates of ferric generation at lower dissolved oxygen. Operating at lower dissolved oxygen levels will improve oxygen utilization. In addition, dissolved iron in autoclaves and to a lesser extent in atmospheric leach vessels will precipitate and generate acid. This will offset some of the acid demand required in the oxidative acid leach step 186 and heap leach step 120 (for example, the acid demand of gangue reactions and sulphide oxidation)

The slurry 189 from the oxidative acid leach step 186 reports to a thickener 190. The decant 192 from the thickener 190 reports back to the second intermediate liquor pond 127.

The underflow 194 from the thickener 190 reports to a filter 193. In the filter 193 the solid residue 195 from the direct oxidation leach is further dewatered and rinsed. The solid residue 195 reports to the cyanidation step 196. The liquor 205 from the filter 193 mixes with the thickener decant 192 and the combined stream 206 also reports to the second intermediate liquor pond 127. Accordingly, the copper and uranium in the liquor 205 would be ultimately recovered through the extraction step 152. The solid residue 195 is re-pulped in a stirred preconditioning tank (not shown) with water to approximately 40 to 50% solids. A lime slurry is added to achieve a pH to 10 to 10.5. The slurry from the preconditioning tank will report to series of stirred tanks 197. In these tanks sodium or calcium cyanide will be added to dissolve the gold and silver that is in the solid residue 195. Activated carbon is introduced into the last tank of the series and is moved in a counter current fashion to the slurry via a series of pumps and screens. The soluble gold and silver loads onto the activated carbon, which is then removed, acid washed and eluted. The gold and silver 198 are recovered by

electrowinning and further refined through to bullion through smelting.

Other potential routes for gold recovery include cyanide leaching followed by carbon in pulp or possibly the Merrill-Crowe process.

The flotation tails 180 and the heap leach purge liquor 159a are combined into a slurry which is pumped to a tails storage facility 182. A small decant 199 (7 to 10%) from the tails storage facility 182 is pumped to a two stage neutralization plant 200 comprising a series of air sparged tanks 201 with high solidity gas transfer agitators 202. In the first section of the neutralization plant 200 (first 3 tanks) the pH of the slurry in the second neutralisation tank is raised to pH 5 to 5.5 with limestone. In the second series the pH is increased to pH 6 to 7 with lime.

A recycle of the final slurry is returned to the first tank to improve the limestone consumption rate and provide precipitate seeding to improve the settling characteristics of the sluny. Air 203 is supplied to the tanks 201 to convert the ferrous iron in the liquor to ferric iron and hence precipitate it at elevated pH as ferric hydroxide. Calcium concentrations will be at the gypsum saturation levels.

The final neutralised sluny 204 reports to the flotation tails thickener 176. The small amount of neutralized solids is co-thickened with the flotation tails to more efficiently dewater the neutralized slurry. The additional water that is decanted helps reduce the raw water requirement of the milling step 150. Calcium concentrations are reduced below saturation by the influent raw water into the milling circuit.

Referring now to Figure 3, the second embodiment of the flowsheet differs from the first embodiment shown in Figure 2 as follows. The combined stream 306 arising from the mixing of liquor 305 and the thickener decant 292 is mixed with the raffinate stream 259b before being recycled to the final heap leach stage 222c. In this embodiment, the higher ferric bearing liquors (292, 306) from the oxidative acid leach step 286 report to the final heap leach stage 222c.

Referring now to Figure 4, the third embodiment of the flowsheet differs from the first embodiment shown in Figure 2 as follows. In the third embodiment, at least some of the liquor required for the oxidative acid leach step 386 is provided from at least one intermediate leach ponds, in this case the first intermediate leach solution pond 325. In this manner, increased concentration of soluble iron is provided for the oxidative acid leach step 386, so as to optimise parameters within that step. In addition, the second intermediate leach solution 361 and the rinse water 367 are combined and recycled to the first heap leach stage 322a as a leach solution 363. In addition, the combined stream 406 arising from the mixing of liquor 405 and the thickener decant 392 is directly recycled to the first intermediate leach solution pond 325.

Referring now to Figure 5, the fourth embodiment of the flowsheet differs from the first embodiment shown in Figure 2 as follows. In the fourth embodiment, the heap leach step 420 includes only first and second stages of leaching, 422(a) and 422(b), respectively. As in the previous embodiments, the ore being leached in the second stage 422(b) has been previously leached in the first stage 422(a). The intermediate leachate 424 reports to a single intermediate leach solution pond 425 (as opposed to the embodiment in Figure 2 which includes two intermediate leach solution ponds). This modification is used to treat ores that have higher content of copper so as to keep the soluble copper tenor in the PLS reporting to the copper solvent extraction facility in optimal range. A further consequence of having a single intermediate leach solution pond is that the combined stream liquor 506 reports directly to the first heap leach stage 422 (a).

In the extraction step 452, the raffmate stream 459b is recycled to the first heap leach stage 422(a). In comparison, the embodiment of Figure 2 shows the equivalent raffmate stream (159b) as being recycled to the last heap leach stage (122c). This modification results in uncontrolled acid generated from electrowinning and transferred to the raffmate through copper SX reporting to stage one of the heap leach which has the highest acid demand. The acid tenor reporting to stage 2 of the heap can then be controlled to optimise acid consumption and optimise the acid tenor reporting to the PLS, as this is an important parameter for copper extraction.

Advantages of the disclosed process and plant are:

• Robustness: The response times of the heap leach step are very long. This means that the heap leach and mill/flotation steps can be decoupled. The copper concentrate arising from flotation is a relatively small flow and can be given a reasonable storage time and hence a reasonable surge capacity between the steps/equipment is possible. There are no other dependencies.

• Integration of the heap leach and oxidative acid leach stages: In ore processing plants where only a concentrate is produced, the direct oxidation of the concentrate whether through atmospheric or pressure leach results in the need to partially neutralize return liquors to control iron and acid tenors. This is expensive in lime cost. In the case of the present integrated heap leach and oxidative leach process, excess acid is absorbed by the acid demand of the heap leach step. Further, the excess soluble iron can exit the system by purging liquor from the raffmate after the extraction step, where 90 to 93% of it may be sequestered into tails storage facility meaning only 7 to 10% requires neutralization. The single, common extraction step (containing eg, CuSX/EW plant) means that copper recovery split between the heap leach and oxidative leach steps can change with no impact open the operation of the extraction step. The heap leach step provides iron tenor for the oxidative leach step, which if an atmospheric operation will allow the reduction in operating dissolved oxygen and hence an improvement in oxygen utilization.

Gypsum and jarosite: The formation of gypsum and jarosite seed crystal during the curing step promotes the formation of these precipitates within the heap and not in the pipe work used during subsequent processing. This reduces the maintenance requirements.

Plant size: Heap leach operations are readily expandable up to very large tonnages and are suited to large scale large open pit mining operations.

Uranium recovery: Testwork has suggested that employing a heap leach step may improve the uranium dissolution of ores samples by up to 15-20%.

Oxidative Leach Step: This type of technology is able to handle low grade and variable mineralogy concentrates. If an atmospheric type process is selected it is easily incrementally expandable.

Extraction step: the Copper SX/EW can receive a clean low TSS (total soluble salts) feed, low acid and copper tenors that result in efficient recovery of copper. Again the single, common extraction step enables changes in the copper recovery split with no impact open the operation of the extraction step.

Moreover, the uranium extraction stage may still get CRUD formation from zirconia, but data to date suggest that the concentrations of bismuth are significantly reduced.

Improved economics: The combination of minimal losses of soluble uranium and extended leach time enabled by the low cost heap leach results in higher uranium production. The combination of high copper recoveries from solution coupled with equivalent flotation performance on the ripios as is achieved on the ore also results in greater over all copper recoveries. Heap leaching provides a lower capital and less aggressive leaching environment. The latter leads to lower operating costs. The combination of the all the above generates a greater return on investment.

It has been found that in the disclosed method the rate of oxidation of ferrous iron is proportional to the square of the ferrous ion concentration. The disclosed process and plant are particularly efficient for the Olympic Dam ore body because the combination of gangue and value minerals in this ore body creates elevated iron tenors which allow this process to operate at an efficient rate.

Whilst a number of specific embodiments have been described, it should be appreciated that the process and plant may be embodied in many other forms.

References to the background art herein do not constitute an admission that the art forms a part of the common general knowledge of a person of ordinary skill in the art. Those references are also not intended to limit the application of the process as disclosed herein.

In the claims which follow, and in the preceding description, except where the context requires otherwise due to express language or necessary implication, the word

"comprise" and variations such as "comprises" or "comprising" are used in an inclusive sense, i.e. to specify the presence of the stated features but not to preclude the presence or addition of further features in various embodiments of the process and plant as disclosed herein.