INTEGRATED HYDROMETALLURGICAL AND PYROMETALLURGICAL METHOD FOR PROCESSING ORE

Abstract

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 pregnant leach solution and a ripios; c. subjecting the ripios to flotation to produce a copper containing ripios concentrate and tailings; and d. subjecting the ripios concentrate to a smelting process to produce a smelted copper product. e. recovering copper and uranium from the pregnant leach solution.

Claims

1. A process for recovering copper, uranium and one or more precious metals from an ore material, the process comprising: a. forming a heap of an ore material; b. subjecting the heap of the ore material to an acidic heap leach using an iron comprising acidic leach solution in the presence of an oxygen containing gas, and producing a pregnant leach solution and a ripios; c. subjecting the ripios to flotation to produce a copper comprising ripios concentrate and tailings; and d. subjecting the ripios concentrate to a smelting process to produce a smelted copper product. e. recovering copper and uranium from the pregnant leach solution.

2. The process of claim 1, further comprising: f. electro-refining the smelted copper product to produce cathode copper and precious metal-containing anode slimes; and g. recovering one or more precious metals from the precious metal-containing anode slimes.

3. The process of claim 1, wherein the ore material comprises copper sulphides and uranium minerals.

4. The process of claim 1, wherein the smelting process comprises two smelting stages.

5. The process of claim 1, wherein the smelting process comprises a primary smelting stage and a secondary smelting stage.

6. The process of claim 5, wherein the primary smelting stage is conducted in a matte smelting furnace.

7. The process of claim 5, wherein the secondary smelting stage is conducted in a direct to blister furnace.

8. The process of claim 1, wherein the ore material also comprises an iron containing mineral which partly or wholly comprises any iron containing acidic leach solution.

9. The process of claim 8, wherein the iron containing mineral comprises a gangue mineral such as hematite, siderite, or chlorite.

10. The process of claim 1, wherein the oxygen containing gas is air or oxygen enriched air.

11. The process of claim 1, wherein the pregnant leach solution comprises copper and uranium.

12. The process of claim 1, wherein the iron containing acidic leach solution comprises ferric ions which oxidise the ore material to dissolve copper and uranium, resulting in reduction to ferrous ions which are reoxidised to ferric ions by reaction with the oxygen containing gas.

13. The process of claim 1, wherein the acidic heap leach is conducted in more than one stage.

14. The process of claim 1, wherein the ripios is milled prior to flotation in process (c).

15. The process of claim 14, wherein the ripios is milled to a particle size comprising a P80 of approximately 75 microns, preferably approximately 35 microns.

16. The process of claim 1, wherein the ripios concentrate comprises copper containing sulphides, uranium minerals and one or more precious metals.

17. The process of claim 16, wherein the one or more precious metals metal is gold and/or silver.

18. A process for smelting a first copper concentrate that has a ratio of Cu:S below 1.4, the process comprising: (a) smelting a first copper concentrate that has a ratio of Cu:S below 1.4 in a first stage comprising a matte smelter to produce a matte copper product containing at least 40% copper; (b) combining the matte copper product with a second copper concentrate to produce a feed material having a ratio of Cu:S that is higher than that of the first copper concentrate; and (c) smelting the feed material in a second stage comprising a direct to blister furnace to produce blister copper having at least 95% copper.

19. An integrated plant for use in a method for recovery of copper and uranium from an ore material as disclosed above, comprising: means for forming a heap of an ore material; means for supplying an acidic leach solution and an oxygen containing gas to the heap of the ore material to form a pregnant leach solution and a ripios; extraction equipment for extracting copper and uranium from the pregnant leach solution; means for collecting and transferring the pregnant leach solution 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 of the ore material to the flotation equipment; and one or more furnaces for smelting the copper concentrate to produce a smelted copper product and means to transfer the copper concentrate to the one or more furnaces.

20. The integrated plant of claim 19, further comprising electro-refining equipment for electrolytically processing the smelted copper product to produce cathode copper.

Description

BRIEF DESCRIPTION OF THE DRAWINGS

[0071] Notwithstanding any other forms which may fall within the scope of the process as set forth in the Summary, specific embodiments will now be described, by way of example only, with reference to the accompanying drawings in which:

[0072] FIG. 1 shows an embodiment of a flowsheet for a process for recovering copper, uranium and one or more precious metals from an ore material;

[0073] FIG. 2 shows an embodiment of a flowsheet for a process for smelting copper.

[0074] FIG. 3 is a graph of the average Cu and U extraction (%) versus leach time (days) for a range of ore mineralogies during the heap leach stage of the disclosed recovery process.

[0075] FIG. 4 is a bar graph showing the total copper recovery using a combination of heap leaching and flotation and smelting.

[0076] FIGS. 5(a) and (b) plot copper and gold recovery, respectively, from feed ore and floated ore.

[0077] FIG. 6 is a plot of copper recovery from ripios concentrate floated in seawater as compared with ripios floated using tap water.

DETAILED DESCRIPTION OF SPECIFIC EMBODIMENTS

[0078] In each of FIGS. 1 and 2, like reference numerals are used to denote similar or like parts.

[0079] Referring firstly to FIG. 1, shown is a flowsheet 10 illustrating an embodiment of a process for recovering copper, uranium and one or more precious metals from an ore material.

[0080] 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: primary, secondary and tertiary.

[0081] The crushed ore 15 reports to an agglomeration step 16, where it is contacted with water and raffinate and/or acid and/or a solution containing saline conditions in a drum to bind the fine particles of ore to the larger lumps.

[0082] 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 10 m 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. The stacked ore may be cured for a period of up to 30 days or more prior to the commencement of acid leaching.

[0083] 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 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.

[0084] The pregnant leach solution 40 reports to an extraction step 52 in which copper and uranium are extracted from solution. The extraction step 52 reclaims copper and uranium that are dissolved during the heap leaching step 20.

[0085] Copper is extracted using a copper solvent extraction/electrowinning (CuSXEW) stage 53. In the CuSXEW stage 53, the pregnant leach solution 40 is contacted in a counter current flow with an organic phase that loads copper to produce a loaded organic and a copper depleted aqueous stream (the copper SX raffinate, 56). 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.

[0086] The copper raffinate 56 reports to a uranium solvent extraction and refining facility 54. This is mostly analogous to the CuSXEW stage 53, 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 as the final product. The raffinate 57 from the uranium solvent extraction stage is recycled back to the heap 20.

[0087] The ripios 42 produced from the heap leach 20 reports to a milling step 50. In this step, the ripios is milled in either a ball or pebble mill. Make up water 51 and lime may be added if required.

[0088] 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 ripios concentrate 68. This ripios concentrate 68 accounts for approximately 1-2% of the total ore mass. The ripios concentrate 68 is thickened in concentrate thickening step 70 to about 40%-50% solids. Excess water 72 is returned to the milling step 50.

[0089] 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.

[0090] The thickened ripios concentrate 84 resulting from thickening step 70 primarily contains copper sulphides, pyrite, elemental sulphur, hematite and silica. It may also contain naturally occurring radionuclides, rare earth elements and gangue materials. In some instances calcium fluoride is present and may require removal through leaching or additional flotation cleaning. The particle size is typically a P80 of 35 m, however may range from 20 m to 105 m. The copper concentrate grade is typically around 25 wt % Cu, however may range from 15 wt % to 45 wt %.

[0091] The solids density of the ripios concentrate slurry may be increased from approximately 25 wt % up to 60 wt % (range 45 wt % to 65 wt %) in the thickening step 70. The thickener underflow is dewatered using a pressure filter 85 to produce a filter cake having a moisture content preferably less than 15%, and may be in the range of 6 wt % to 12 wt %, such as approximately 10 wt %. Filter cake squeezing and air purging may be used to ensure low moisture content. In the event of the primary smelting furnace being shutdown, the thickener underflow may be pumped to a concentrate storage pond (not shown) from which the ripios concentrate may be reclaimed and processed by the thickeners when required.

[0092] The filter cake 86 from the filter 85 may be conveyed to a shuttle conveyor (not shown) which distributes the filter cake 86 as ripios concentrate feed into concentrate surge bins (also not shown). These bins may be padded with inert gas (typically nitrogen) to minimise oxidation and self-heating of the ripios concentrate feed.

[0093] The ripios concentrate feed is supplemented with a proportion of the conventional concentrate from the pre existing ore processing plant.

[0094] The ripios concentrate feed and conventional concentrate are metered from the concentrate bins onto a collection conveyor which feeds the wet blended concentrate to a drying stage.

[0095] The drying stage may comprise using a rotary steam dryer. The rotary steam dryer may use steam produced in the subsequent smelter stage or acid plant boilers (described below). The steam pressure may be lowered to less than 10 Bar, potentially as low as 7 Bar to reduce the temperature and corresponding risks of overheating the elemental sulphur in the concentrate. The dryer may have an inert gas purge, to minimise the oxygen content. The preferred inert gas is nitrogen.

[0096] An alternative to the rotary steam dryer is to use a flash dryer where a fossil fuel is burnt and the combustion off-gas is used to convey and dry the concentrate over a short period of time. The dried concentrate from either drying option preferably will have a moisture content less than 0.5% and typically around 0.2%.

[0097] Dust from the dryer may be collected. This may be done in a cyclone, baghouse or particulate scrubber. In one embodiment, the preferred configuration has the rotary steam dryer with a baghouse. Baghouse insulation may be installed to reduce elemental sulphur deposition as offgas streams cool.

[0098] The dried concentrate feed 87 may be then pneumatically transferred to concentrate storage silos (not shown) to provide additional residence time. The pneumatic transfer may be completed with nitrogen to minimise the risks of self-heating the concentrate. The dried concentrate feed 87 is then pneumatically conveyed, again with nitrogen, to a dry concentrate day bin prior to being introduced into the primary matte smelter 88. The smelter 88 could be a bath type or flash type of smelter, with the flash smelter as the preference. If the primary matte smelter 88 is a flash smelter, the blended concentrate should be dried to <0.5% moisture prior to flash smelting. However, if self-heating of the blended concentrate causes issues during drying and the concentrate can not be dried to <0.5%, then bath smelting may instead be used in the first smelting step (a).

[0099] Should a bath type of smelter be installed, the dryer would potentially be replaced with an agglomerator and the moist feed would be transferred with conveying.

[0100] Silica flux may be transferred to a flux day bin and conveyor system (not shown) which feeds the flux to the primary matte smelter 88.

[0101] A portion of sulphated dust produced in the waste heat boiler or electro-static precipitator of the smelter 88 may be recycled back into the primary matte smelter 88.

[0102] The existing feed preparation circuit will process the remaining conventional concentrate. The concentrate will be processed through the existing thickener, filters and steam coil concentrate dryers to produce dry high grade concentrate. The dry concentrate is pneumatically conveyed to concentrate silos.

[0103] With reference to FIG. 2, there is shown an embodiment of a flowsheet 90 for smelting of the dried concentrate feed 87. The primary matte smelter 88 comprises a flash smelter.

[0104] In the embodiment shown in FIG. 2, the dried concentrate feed 87 is distributed evenly to the concentrate burner in the primary matte smelter 88 in order to achieve uniform distribution of dry charge to the reaction shaft. Oxygen enriched air 92 (typically 80 vol % oxygen, ranging from 50% to 90%) is blown through the concentrate burner as process air. Distribution air is blown horizontally through the concentrate burner tip to evenly distribute the charge.

[0105] The primary matte smelter 88 produces copper matte 94 (typically containing 70 wt % Cu, range 60% to 75%), iron silicate slag 96 (with typical Fe:SiO.sub.2 ratio of 1.2 and 2 wt % Cu) and sulphur dioxide bearing off-gas.

[0106] Matte is tapped periodically from the primary matte smelter 88 and laundered a short distance from the smelter directly into the matte granulation system.

[0107] For the alternative in which the primary matte smelter 88 consists of a bath smelter, the blended flux, concentrate and dust is fed into the primary matte smelter 88 via a conveyor. Enriched oxygen is introduced below the molten metal bath through a lance which mixes and oxidised the materials to produce a copper matte, iron silicate slag and sulphur dioxide bearing off-gas. The matte and the slag from the primary matte smelter 88 will be tapped together and separated in a rotary holding furnace before matte granulation.

[0108] In both cases additional heat can be supplied to the primary matte smelter 88 through the introduction of heavy fuel oil or liquid petroleum gas (LPG) via burners. The overall heat balance may also be controlled through adjusting the oxygen enrichment of the air entering the furnace.

[0109] Slag 96, typically containing 2 wt % Cu (ranging from 0.5% to 8%, typically from 1% to 6%, such as from 1% to 5%) is tapped from the primary matte smelter 88 (typically the uptake shaft end of the flash smelter settler (or bath smelter rotary holding furnace) and sent to the slag concentrator 100 for further copper recovery.

[0110] Off-gas from the primary matte smelter 88 is directed to the primary smelter waste heat boiler to generate saturated steam (eg, at 70 bar). Should the steam production be excess to requirements, it may be treated in an air cooled condenser or a steam turbine and generate electricity from the excess steam to be fed back into the grid.

[0111] Off-gas exiting the primary smelter waste heat boiler may be sent to an electrostatic precipitator then to wet gas cleaning. The wet gas cleaning system includes a quench tower, radial flow scrubber, gas cooling tower and wet electrostatic precipitators. This gas stream then enters an acid plant 102 where the SO.sub.2 in the gas is converted to sulphuric acid.

[0112] The dust collected from the primary smelter waste heat boiler and electrostatic precipitator may contain high levels of some radionuclides and can either be recycled to the primary matte smelter 88 or bled to a dust dissolving tank. The copper solution from dust leach may be recycled to either the heap leach CuSX/EW 52 or existing tails leach CuSX/EW facilities.

[0113] The granulated matte 94 is processed through a mill which grinds and dries the matte. The dry ground matte and dry conventional concentrate together with silica or lime flux are subsequently fed into the secondary smelter 98 comprising a Direct to Blister Hybrid Furnace (DBF). The DBF furnace would conventionally only treat concentrate. However, in the present process, the converting of the matte to blister and the treatment of the concentrate direct to blister in the same furnace at the same time enables it to be viewed as a hybrid DBF application (treatment of both matte and concentrate products). The simultaneous treatment of matte and concentrate together in a single DBF furnace avoids the need for a converting furnace to treat the matte alone.

[0114] The secondary smelter 98 produces blister copper 104a, slag 106 and sulphur dioxide bearing off-gas. The blister copper from the DBF Hybrid will be tapped and laundered to anode furnaces.

[0115] The slag 106a from the secondary smelter 98 is batch-tapped to an electric furnace 108 to recover copper via reduction of copper oxide using coke. The typical target electric furnace slag copper content is 4%, but may range from 2% to 8%.

[0116] Blister copper 104b from the electric furnace 108 is laundered to the existing anode furnaces. Electric furnace slag 106b is sent to the slag concentrator 100 for copper recovery.

[0117] The blister copper 104a, 104b from the secondary smelter 98 and the electric furnace 108 will enter an anode furnace (not shown). Plant air is injected through tuyeres into the molten bath of the Anode Furnace to oxidise the sulphur in the blister copper 104a, 104b. Following oxidation, a reduction cycle is completed by injecting LPG/Nitrogen into the tuyeres to control the oxygen content of the blister copper. Slag from the anode furnaces can be skimmed and recycled to recovery the copper.

[0118] Anode copper is cast into anodes in a casting wheel. The anodes are sent to an electro-refinery 110 for production of cathode copper 112. The copper anodes are loaded into cells filled with an electrolyte solution of sulphuric acid & copper sulphate. Next to each copper anode is a stainless steel cathode. A direct current of up to 31,000 amps is passed through the electrolyte solution, from the anode to the mother plate. The copper in the anode dissolves and redeposits as 99.99% pure copper on the stainless steel mother plate.

[0119] After a period of time (typically around 10 days) the copper is stripped from the cathode plates and bundled for shipment. Once a portion of the copper anodes have been dissolved (typically about 16% of their original weight), the remnant scrap is removed, washed & remelted.

[0120] Anode slimes that are generated by insoluble impurities in the copper anodes fall to the bottom of the cell. These slimes are further collected, treated and pumped to a gold room to extract gold & silver 114.

[0121] The slimes are first leached to remove the copper through leaching in aerated strong acid. This stage also removes some selenium and tellurium. The de-copperised slimes are neutralised to a pH of 8 or greater before being transferred to cyanidation tanks where cyanide is added at a pre-determined ratio and oxygen is injected. The gold and silver are dissolved in the cyanide solution, leaving behind insoluble impurities. The solution is filtered and pumped to the pregnant liquor storage tank. The residue is repulped with barren solution and ferrous sulphate to stabilise remaining cyanide before being pumped to the tails circuit.

[0122] The pregnant solution is pumped to a zinc precipitation tank where the gold and silver are precipitated out of solution. The precipitate is then filtered in a pressure filter and washed and re-pulped in potable water. The cake then undergoes a further leach step, where the zinc is leached in sulphuric acid. This also leaches the selenium through oxidation with nitric acid and/or hydrogen peroxide.

[0123] Water and salt are then added to the solution to leach silver which is further precipitated as a silver chloride. The silver chloride and gold metal are then filtered and collected.

[0124] The solids are repulped in a small amount of potable water to maintain a paste consistency. The material is fed into a conveyor dryer to remove the majority of the moisture before it reaches the roaster.

[0125] Sodium carbonate, silica flux and borax are added to the roaster are added to convert the silver chloride to silver metal and to remove any lead that is present in the precipitate. The molten material from the roaster is poured into Dore buttons and slag trays. The Dore is fed into the Dore furnace. The Dore furnace will remove the remaining lead, including Pb210 and selenium through oxidation and volatilisation. The Dore is then poured into moulds which cast the metal into anodes.

[0126] The Dore anodes are subjected to electrolysis in order to remove any impurities. 99.9% silver is produced at the electrolysis cathode and the gold within the anodes reports as mud. The mud is directed to the gold cage for further processing and the silver is centrifuged, weighed and cast into silver ingots.

[0127] The gold mud is dried and reagents and fluxes are added. The gold is then smelted into electrolysis anodes where they are placed into an electrolysis cell and current is applied. A resultant gold cathode is produced, stripped, washed and dried. The gold cathode is then melted and cast into gold bullion.

[0128] Advantages of the disclosed process and plant are:

[0129] The use of a multi-stage smelting operation on the copper concentrate is an efficient and economical means of recovering copper from a concentrate having a relatively low Cu:S ratio. In particular, the use of a first smelting stage to produce copper matte from low Cu:S concentrate feed, followed by a second stage in which the matte and concentrate are smelted in a DBF furnace provides a more efficient and economical means of recovering copper when compared to hydrometallurgical processing of the ripios concentrate. The process therefore allows utilisation of existing equipment on-site to smelt ore concentrates despite changing concentrate composition over the life of the mine, which enables significant cost advantage.

[0130] 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.

[0131] 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.

[0132] Plant size: Heap leach operations are readily expandable up to very large tonnages and are suited to large scale large open pit mining operations.

[0133] Uranium recovery: Testwork has suggested that employing a heap leach step may improve the uranium dissolution of ore samples [0134] 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. The uranium extraction stage may still get CRUD formation from zirconium and bismuth, but data to date suggest that the concentrations of zirconium and bismuth are significantly reduced

[0135] 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.

[0136] 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.

EXAMPLE

Heap Leaching and Flotation of Ripios

[0137] Ore samples, having a diverse range of mineralogy that reflected the actual range of ore mineralogy at Olympic Dam, were subjected to heap leaching (ie, step (a) of the disclosed process) in columns using an iron containing, acidic saline leaching solution in the presence of air. The Cu:S ratios of the ore samples ranged from approximately 3 to less than 1. The ore contained a number of iron containing minerals, such as siderite and chlorite, which provided a source of ferric in the leaching solution. The acid consumption of the samples also varied from high to low, reflecting the variation in gangue mineralogy. The U.sub.3O.sub.8 grade of the ore was variable ranging from approximately 200 ppm to 1,000 ppm.

[0138] The dissolution of copper and uranium was measured over time. FIG. 3 shows the results of these tests. The average dissolution (%) of U.sub.3O.sub.8 and Cu are plotted against time (days). The actual Cu and U extraction values ranged from 20% to 80% and from 30% to 90%, respectively, after 250 days of leaching. It can be seen that after approximately 250 days, the average dissolution of uranium was approximately 72% and of copper was 57%.

[0139] The ripios resulting from the above Heap Leach stage was ground to a particle size between 150 and 50 microns and subjected to flotation. Flotation resulted in good copper and precious metal recoveries in the float fraction of the ripios of approximately greater than 90% such as 93% and greater than 65%, such as 73% respectively.

[0140] FIG. 4 is a bar graph showing the total copper recovery from a series of larger scaled crib heap leach and ripios flotation pilot plant tests. Each bar represents the total copper recovery from an ore sample using a combination of heap leaching and flotation of the ripios after heap leaching. The lower, lighter section of each bar represents the % copper dissolution from heap leaching the sample in a crib assembly. The upper, darker section of each bar represents the amount of copper in the ripios float fraction (ie recovered from the heap leach ripios after ripios flotation). (The float fraction would then become the feed to the potential downstream concentrate impurity leach and subsequent smelting process). Cribs 2 to 9 were floated using non-saline water (nominated as less than 5 gpl chlorides), whereas cribs 10 to 31 were floated using saline water at 10-35 gpl chlorides, but typically 15-25 gpl chlorides. As can be seen, despite variation in the amount of copper dissolved from ore samples during heap leach (ranging from about 29% to 92%), there was a high recovery of the remaining copper in the sample using flotation, resulting in an overall copper recovery of at least 90%.

[0141] FIGS. 5(a) and (b) plot copper and gold recovery, respectively, from feed (run of mine (ROM)) ore and floated ore. Each data point compares the % recovery from floated ore as compared with the % recovery from the ROM ore. Copper is partially leached during the heap leach whereas minimal or no gold is leached. The plots are approximately linear, indicating that metal recovery from the floated ore is similar to that from ROM ore. Given that flotation of aged or weathered ore is typically impeded due to the presence of oxidised or coated surfaces, it would be expected that flotation of ripios would also be impeded due to the conditions experienced during heap leaching. The results displayed by both FIGS. 5(a) and (b) were therefore surprising in that metal recovery from ripios was unexpectedly similar to that from the ROM.

[0142] FIG. 6 is a plot of copper recovery from ripios concentrate floated in saline seawater as compared with recovery from ripios floated using non-saline (tap) water. The plot is approximately linear, indicating that flotation may be as successfully conducted using saline water as using non-saline water.

Smelting of Ripios Concentrate

[0143] Table 1 shows an example of the mineralogical and chemical compositions of a conventional concentrate and of a ripios concentrate produced using the process of the present disclosure. As noted previously, the conventional concentrate was produced from flotation of run of mine ore. In contrast, the ripios concentrate was produced from flotation of the ripios remaining after heap leaching the ore in accordance with the present disclosure. The ripios concentrate tends to have lower amounts of secondary sulphides (such as Cu.sub.5FeS.sub.2, Cu.sub.2S, CuS) and a higher amount of chalcopyrite. It also includes elemental sulphur. Consequently, the ripios concentrate has lower overall copper and higher overall sulphur contents than the conventional concentrate. This translates into the ripios concentrate having a lower Cu:S ratio (0.83) than that of the conventional concentrate (1.66) in this test.

TABLE-US-00001 TABLE 1 Compositions of conventional and ripios concentrates Conventional Ripios Formula Concentrate Concentrate Mineral Secondary Cu.sub.5FeS.sub.2, Cu.sub.2S, CuS 47.4 15.6 sulphides Chalcopyrite CuFeS.sub.2 31.0 44.4 Sulphur S 9.8 Hematite Fe.sub.2O.sub.3 9.7 14.0 Pyrite FeS.sub.2 5.4 3.0 Quartz SiO.sub.2 2.1 4.6 Muscovite KAl.sub.2(Si.sub.3Al)O.sub.10(OH, F).sub.2 2.1 6.3 Magnetite Fe.sup.2+Fe.sup.3+.sub.2O.sub.4 0.4 Fluorite CaF.sub.2 0.06 0.4 Others 1.8 1.9 Element Cu 44.4 25.1 Fe 21.2 25.9 S.sub.tot 26.7 30.2 S.sup.0 9.8 SO.sub.4.sup.2 0.47 0.46 Cu:S 1.66 0.83

[0144] Three smelting tests were conducted using the conventional and/or ripios concentrates, and the results of those tests are set out in Table 2. The smelting tests were conducted in a primary matte smelter, comprising a flash smelter.

TABLE-US-00002 TABLE 2 Smelting Tests (Primary Smelter) Smelting Sample 1 2 3 Feed Material 50:50 Conventional Conventional Con + Ripios Ripios Concentrate Con Concentrate Matte Cu Grade (%) 70.8 76.2 70.7 Fe Grade (%) 7.1 2.7 7.2 S Grade (%) 19.9 19.9 20.5 Slag Cu Grade (%) 3.3 5.9 3.6 Fe Grade (%) 43.9 43.8 38.6 SiO.sub.2 Grade (%) 24.4 24.7 27 S Grade (%) 0.6 0.2 0.6 Fe: SiO.sub.2 1.80 1.77 1.42

[0145] Each of the tests in Table 2 produced high grade copper matte, ie in excess of 70%.

[0146] The matte produced from the primary smelter was granulated and fed to a secondary smelter for conversion to blister copper. The secondary smelter in this case was a Hybrid DBF. While conventionally a DBF would only treat concentrate, in the present process, the furnace treats a combination of both matte and concentrate products simultaneously, thereby avoiding the need for a converting furnace to treat the matte alone. Table 3 shows the results of three smelting tests in the Hybrid DBF with different feed materials. In Sample 4, the feed material comprised a combination of conventional matte (produced according to Sample 1) and ripios matte (produced according to Sample 3). The feed material for Sample 5 was conventional concentrate (see Table 1) and conventional matte (produced according to Sample 1). The feed material for Sample 6 was conventional concentrate (see Table 1) and conventional/ripios matte (produced according to Sample 2). In each Sample, the resulting blister copper exhibited a grade of >99% copper, and contained low sulphur and controlled Fe:SiO.sub.2, indicating that high quality blister can be produced according to the present process.

TABLE-US-00003 TABLE 3 Smelting Tests (Secondary Smelter) Smelting Sample 4 5 6 Feed Material Conventional Conventional Conventional Matte Concentrate + Concentrate + (Sample 1) + Conventional Conventional/ Ripios Matte Matte Ripios Matte (Sample 3) (Sample 1) (Sample 2) Blister Cu Grade (%) 99.9 99.9 99.7 Fe Grade (%) 0.008 0.011 0.206 S Grade (%) <0.05 <0.05 <0.05 Slag Cu Grade (%) 34.3 31.6 25.9 Fe Grade (%) 22.8 31.2 34.2 SiO.sub.2 Grade (%) 12.8 14.7 14.6 S Grade (%) <0.1 <0.1 <0.1 Fe:SiO.sub.2 1.8 2.1 2.3

[0147] 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.

[0148] 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.

[0149] Those references are also not intended to limit the application of the process as disclosed herein.

[0150] 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.