Abstract
A method for recovering one or more of copper, uranium and a precious metal from an ore material, including: (a) forming a heap of the ore material; (b) during active heap irrigation, contacting the heap of the ore material with an iron containing acidic leach liquor having a high chloride content in the presence of an oxygen containing gas, and producing a pregnant leach solution; and (c) recovering one or more of copper, uranium and a precious metal from the pregnant leach solution.
Claims
1.-29. (canceled)
30. A method for recovering one or more of copper, uranium and a precious metal from an ore material, including: (a) forming a heap of the ore material; (b) during active heap irrigation, contacting the heap of the ore material with an iron containing acidic leach liquor having a high chloride content in the presence of an oxygen containing gas, and producing a pregnant leach solution; and (c) recovering one or more of copper, uranium and a precious metal from the pregnant leach solution.
31. The method of claim 30 wherein the ore material contains copper sulphides and/or uranium minerals and optionally one or more precious metals selected from gold and/or silver.
32. The method of claim 30 wherein the chloride content in the leach liquor of step (b) is a minimum of 15 g/L.
33. The method of claim 30 wherein the iron containing leach liquor contains ferric ions, which oxidize the ore material and are reduced to ferrous ions which are subsequently reoxidized to ferric ions by reaction with the oxygen containing gas, and wherein the iron containing leach liquor optionally contains cupric ions that catalyze the reaction of ferrous to ferric ions.
34. The method of claim 30 wherein the redox potential of the leach liquor exceeds 420 mV Ag/AgCl, in the absence of bioleaching microorganisms.
35. The method of claim 30 wherein the iron containing acidic leach liquor has a total iron concentration >0.1 g/L.
36. The method of claim 30 wherein the acidity of the leach liquor is such that the pH of the solution contacting the ore does not exceed pH 3.5.
37. The method of claim 30 wherein the method is carried out under ambient temperature and atmospheric pressure conditions.
38. The method of claim 30 wherein the pregnant leach solution is pretreated to adjust solution chemistry prior to recovery of one or more of copper, uranium and a precious metal therefrom.
39. The method of claim 38 wherein pretreatment of the pregnant leach solution formed in step (b) preferably comprises reduction of solution redox potential and/or consumption of at least some acid.
40. The method of claim 30 wherein prior to forming a heap in step (a), the ore material is formed into agglomerates by mixing the ore material with acidic agglomerating solution to give the agglomerates a final moisture content of up to 25 wt %.
41. The method of claim 40 wherein the acidic agglomerating solution is added to the ore material at up to approximately 20 kg/ton ore.
42. The method of claim 30, wherein, prior to contacting the ore with the acidic leach liquor in step (b), the method includes the steps of: i. treating the ore with a pre-leach solution having no, or a relatively low, chloride content, as compared with the high chloride, acidic leach liquor of step (b), in order to dissolve at least some of the uranium and gangue minerals in the ore and produce a leachate, and to deplete deleterious elements in the ore available to form salt precipitates; and ii. treating the leachate to recover uranium and remove the deleterious elements.
43. The method of claim 42, wherein the chloride content in the pre-leach solution ranges from 0 to 50 g/L.
44. The method of claim 42, wherein step (i) is conducted in the presence of an oxidant.
45. The method of claim 42, wherein the solution redox potential of the pre-leach solution is less than 450 mV Ag/AgCl.
46. The method of claim 42, wherein step (ii) includes recovering uranium by solvent extraction.
47. The method of claim 30, wherein, prior to contacting the ore with the acidic leach liquor in step (b), the method includes the steps of: i. treating the ore with a pre-leach solution having lower solution redox, as compared with the high chloride, acidic leach liquor of step (b), in order to dissolve at least some of the uranium and gangue minerals in the ore and to produce a leachate; and ii. treating the leachate to recover uranium.
48. A method of recovering one or more of a base metal, uranium and a precious metal from an ore wherein the ore is subjected to an active leach cycle in which the ore is contacted with an iron containing acidic leach liquor having a high chloride content in the presence of an oxygen containing gas, wherein: a. the ore is optionally agglomerated; b. the leach liquor has a solution potential that exceeds 450 mV Ag/AgCl, in the absence of microorganisms; c. the leach liquor has a total iron concentration >0.1 g/L; d. the ore has a moisture content in the range of 2 to 25 wt. %; e. the leach liquor has a pH that does not exceed pH 3.0; and f. the leach liquor has a Cl.sup.− ion concentration between 20 and 230 g/L.
49. A process of extracting one or more of a base metal, uranium and a precious metal from an ore material which is enhanced at solution potentials exceeding 450 mV Ag/AgCl, in the absence of any microorganism, by contacting the ore material in an active leach cycle using an acid solution at a high chloride content containing dissolved iron.
Description
BRIEF DESCRIPTION OF THE DRAWINGS
[0072] The process is further described by way of example with reference to the accompanying drawings in which:
[0073] FIG. 1 illustrates in block diagram form a first embodiment of a flowsheet for the heap leaching of copper and uranium ore.
[0074] FIG. 2 illustrates in block diagram form a second embodiment of a flowsheet for the heap leaching of copper and uranium ore.
[0075] FIG. 3 is a graph of dissolved silica concentration (mg/L) versus leach time (days) for different amounts of concentrated sulphuric acid added during ore agglomeration. Stars=no acid; squares=8 kg/ton acid; triangles=12 kg/ton acid; crosses=25 kg/ton acid.
[0076] FIG. 4 is a graph showing Cu dissolution (ie, mass Cu dissolved/mass Cu in sample, expressed as percent) versus time (days) for different concentrations of chloride in leach liquor. Circle=no chloride; square=15 g/L chloride; diamond=30 g/L chloride; triangle=45 g/L chloride.
[0077] FIG. 5 shows discharge liquor (such as pregnant leach solution, or an intermediate leach solution) redox potential (mV (Ag/AgCl)) versus time for the same chloride concentrations. Circle=no chloride; square=15 g/L chloride; diamond=30 g/L chloride; triangle=45 g/L chloride.
[0078] FIG. 6 is a plot of the discharge liquor redox potential (mV (Ag/AgCl)) versus time (days) for three different solution chloride concentrations: triangles=1-3 g/L, diamonds=15 g/L and squares=150 g/L.
[0079] FIG. 7 is a graph of the Dissolved Oxygen (DO-ppm) in discharge liquor versus time for the same chloride concentrations as in FIG. 6: triangles=1-3 g/L, diamonds=15 g/L and squares=150 g/L.
[0080] FIG. 8 is a graph showing Cu dissolution (ie, mass Cu dissolved/mass Cu in sample, expressed as percent) of predominantly chalcopyrite ore versus time (days) for three different chloride concentrations: triangles=1-3 g/L, diamonds=15 g/L and squares=150 g/L.
[0081] FIG. 9 is a graph showing Cu dissolution (ie, mass Cu dissolved/mass Cu in sample, expressed as percent) bornite/chalcocite ore versus time (days) for three different chloride concentrations: crosses=1-3 g/L, diamonds=15 g/L and squares=150 g/L.
[0082] FIG. 10 is a graph showing uraninite dissolution (ie, mass U dissolved/mass U in sample, expressed as percent) versus time (days) for chloride concentrations of 1-3 g/L (triangles) and 15 g/L (diamonds).
[0083] FIG. 11 illustrates in block diagram form a third embodiment of a flowsheet for the heap leaching of copper and uranium ore.
[0084] FIG. 12 is a graph showing the concentration of U.sub.3O.sub.8 (mg/L) extracted into solvent from a pregnant leach solution at two solution redox potentials: 475 mV (squares) and 410 mV (diamonds).
[0085] FIG. 13 is a graph showing % dissolution of uranium (as U.sub.3O.sub.8) versus time (days) for a chloride leach of uranium ore.
[0086] FIG. 14 is a graph showing % dissolution of copper versus time (days) for a chloride leach of copper ore.
[0087] FIG. 15 is a graph showing the extraction of uranium into the solvent phase at low solution ORP for a range of chloride concentrations.
[0088] FIG. 16 is a graph showing the extraction of uranium into the solvent phase at high solution ORP for a range of chloride concentrations.
[0089] FIG. 17 illustrates in block diagram form a fourth embodiment of a flowsheet for the heap leaching of copper and uranium ore.
[0090] FIG. 18 illustrates in block diagram form a fifth embodiment of a flowsheet for the heap leaching of copper and uranium ore.
DESCRIPTION OF PREFERRED EMBODIMENTS
[0091] The method is described herein with reference to the use of a high chloride mediated, high solution potential, active heap leach cycle, for crushed or run-of-mine (ROM) ore (or ore concentrate).
[0092] An advantage of the disclosed method is to significantly increase the oxidation rate of sulphide minerals or mixed sulphide and oxide minerals during active heap irrigation, and thereby improve metal recovery in a shorter leach cycle and, additionally, to lower, at least to some extent, the operational cost of a heap leach.
[0093] With reference to FIGS. 1 and 2, two flowsheet embodiments 10, 110 of the disclosed method are illustrated. In each embodiment, an ore that contained the following value minerals: [0094] Copper sulphides mostly in the form of chalcopyrite, bornite and chalcocite, but may include other copper sulphide species; [0095] Uranium in the form of uraninite, coffinite and brannerite; and [0096] Gold,
was crushed to a size distribution in the range P80 of 6 to 25 mm. Optimally, the size distribution was between 8 and 12 mm.
[0097] The crushed ore was agglomerated 12, 112 with a combination of process solution (heap leach raffinate liquor) and concentrated sulphuric acid. The concentrated sulphuric acid was added at a concentration in the range of 0 to approximately 20 kg/tonne of ore. The optimal addition was in the range of 6 to 12 kg/ton ore. The raffinate liquor was added (as required) to give a final moisture of around 3.5% (which was preferred for agglomerate formation).
[0098] Solid salt selected from one or more of the following: NaCl, MgCl.sub.2, KCl and AlCl.sub.3, was added to the ore during or post agglomeration as required to tailor the ultimate chloride concentration of the leach liquor. It was preferred to add the salt to the ore as opposed to adding it directly to the solution irrigated in the heap as the salt was then dissolved in situ and the heap acted as a fines filter for any insoluble impurities.
[0099] Salt was added onto the agglomerate conveyor.
[0100] The agglomerated ore was stacked in heaps, 14a, b, c and 114a, b, c on either multilift non-reusable pad or on a reusable pad (not shown). While diagrammatically each heap is shown as separate units a, b, and c, in practice these units are usually continuously stacked. The ripios may be removed after heap leaching and transferred to a ripios dump for storage and potentially extra Cu recovery from long term permeate collection.
[0101] Each heap unit 14a, b, c and 114a, b, c is irrigated at flux of 5-20 l/m.sup.2/h with an optimum of between 8-15 l/m.sup.2/h.
[0102] Concentrated sulphuric acid is added to the solution irrigated to each heap unit to achieve a desired concentration of acid in the discharge. This is usually a minimum to minimize overall process acid consumption.
[0103] The acid reacts with gangue minerals in each heap unit to produce ferrous ions from minerals such as siderite and chlorite. Some ferric ions may be derived from acid reaction with hematite.
[0104] The ferrous ions are converted to ferric ions by oxidation with oxygen. Oxygen is supplied into the heap by blowing air into the heap.
[0105] The acid and ferric ions react with copper sulphide minerals in the ore to release copper sulphate into solution with the ferric consequentially reduced to ferrous ions. An equivalent reaction occurs between acid, ferric ions and uranium minerals in order to release uranium into solution. Under some conditions, an equivalent reaction occurs to release precious metal, especially gold, into solution. In each case, the ferrous ions generated are re-oxidized to ferric ions by oxygen.
[0106] In each of FIGS. 1 and 2, the intermediate leach solution (ILS) produced by heap units 14a and 114a reports to an ILS pond 16, 116. It is then applied to the heap units 14b and 114b, respectively to produce the final pregnant leach solution (PLS) 15, 115 which report to the PLS ponds 18, 118, respectively.
[0107] Copper and uranium are recovered from the PLS 15, 115 through independent copper (CuSX) and uranium (USX) solvent extraction processes, 20, 120 and 22, 122, respectively.
[0108] In FIG. 1, the raffinate from the USX process reports to the heap leach unit 14c as a process solution. The leachate arising from the heap leach unit 14c reports to the heap leach unit 14a as a process solution. Saline water is added to the ILS in the ILS pond 16.
[0109] In FIG. 2, the raffinate from the USX process and the leachate arising from the heap leach unit 114c both report to the raffinate pond 124. Raffinate 126 from the raffinate pond 124 then reports to heap leach unit 114a as a process solution. Saline water is added to the heap leach unit 114c.
[0110] In FIG. 1 the ripios 28 is not washed for any other use. In FIG. 2 the ripios is washed in preparation for milling and floating residual sulphide as well as gold. The process in FIG. 2 may accordingly require a greater raffinate purge to control ions such as Fe, Al and SO.sub.4.sup.−2 because of the wash stage.
[0111] While FIGS. 1 and 2 illustrate two flowsheet embodiments, it is to be understood that there may be a variety of configurations of the same unit operations that could be considered.
[0112] Referring now to FIG. 3, the amount of soluble silica as a function of leach time is shown for different additions of concentrated sulphuric acid during agglomeration. While increased acid addition has not shown to have process improvements in regard to extent of uranium or copper dissolution, there are process improvements arising from reduced silica levels in solution. Solubilized silica becomes problematic as it preferentially precipitates at the surface of the heaps and can cause permeability problems. It is removed from solution by the addition of poly ethyl glycol (PEG) to the part of the first stage leachate. It is preferential to minimize it from the start. The precipitated silica from PEG addition is settled in a clarifier and recycled to agglomeration as a method of disposal.
[0113] FIG. 4 is a graph showing Cu dissolution % versus time (days) for different concentrations of chloride in leach liquor. FIG. 5 shows discharge liquor redox potential (mV (Ag/AgCl)) versus time for the same chloride concentrations. FIGS. 4 & 5 showing that chalcopyrite leaching passivates in the absence of chloride above 420 mV (Ag/AgCl). However this is not the case when the chloride in solution is at or above 15 g/L. This indicates that the passivation point is extended in the presence of chloride.
[0114] FIG. 6 is a plot of the discharge liquor redox potential (mV (Ag/AgCl)) versus time (days) for three different solution chloride concentrations: 1-3 g/L, 15 g/L and 150 g/L.
[0115] FIG. 7 is a graph of the DO (ppm) in discharge liquor versus time for the same chloride concentrations. FIG. 6 shows that the rate of ferrous to ferric oxidation by air is enhanced in the presence of chloride and so the overall process equilibrates at higher REDOX potentials. This is also evident in the decreased dissolved oxygen in the discharge liquor arising from the increased rate of ferric oxidation (FIG. 7).
[0116] FIG. 8 is a graph showing Cu dissolution (weight percent) of predominantly chalcopyrite ore versus time (days) for three different chloride concentrations 1-3 g/L, 15 g/L and 150 g/L. FIG. 9 is a similar graph showing Cu dissolution for bornite/chalcocite ore. It can be seen that copper sulphide minerals dissolution rates increase as a function of increasing redox.
[0117] FIG. 10 is a graph showing uraninite dissolution % versus time (days) for chloride concentrations of 1-3 g/L and 15 g/L. It can be seen that uranium mineral leach extents can also be increased by higher chloride concentration resulting in increased solution RED OX potential.
[0118] FIG. 13 shows the % dissolution of uranium versus time (days) for leaching uranium ores at varying chloride concentrations commencing at 3 to 5 g/L, increasing to between 15 and 20 g/L, then increasing again to 100 g/L. It was found that once the salinity had increased to above 15 to 20 g/l, there was little effect on the amount of uranium dissolution. The data shows that for most leach conditions, irrespective of the salinity of leaching solution, the leach is essentially complete after approximately 100 to 150 days. In contrast, FIG. 14 shows that the rate of copper dissolution is significantly slower and that a minimum of 200 days is required to achieve maximum dissolution. The rate increase observed for many of the columns is due to the addition of salt to increase the salinity of the leach solution from 25 g/L (saline) to 100 g/L (hypersaline). It is evident that an increase in salinity results in an increase in copper dissolution.
[0119] FIGS. 15 and 16 illustrate the effects of solution redox potential and salinity on extraction of uranium into the solvent phase during solvent extraction. At relatively low redox potentials (approximately 2 to 5 g/L ferric), FIG. 15 shows that the extraction of uranium is favoured at relatively low chloride concentrations, and steadily deteriorates as chloride concentration increases. FIG. 16 shows that under relatively high solution redox conditions (approx. 10 to 18 g/L ferric) uranium solvent extraction was poor for anything other than the lowest salinity (25 g/L). These graphs therefore demonstrate that high salinity and/or high solution redox adversely affect uranium solvent extraction. In the following embodiments, leaching of uranium is conducted under lowered solution redox or reduced salinity.
[0120] FIG. 11 illustrates in block diagram form a third embodiment 210 of a flowsheet for the heap leaching of copper and uranium ore, which is a first modification of the disclosed process that includes a low redox pretreatment step. This embodiment differs primarily from the previous embodiments in that the intermediate leach solution (ILS), 216, arising from an aerated heap leach reports to a “pre-leach” heap, 214b, containing acid and ferric consuming materials in the absence of forced aeration. The mineralogy of the pre-leach heap is similar to that of the other (aerated) heap/s, 214a, and primarily contains metal sulphides and uranium ore minerals. The ILS is contacted with the metal sulphides and acid consuming minerals in the pre-leach heap to reduce the ferric ion concentration and acid in solution. Reducing solution redox potential has been found to be advantageous in the subsequent extraction of uranium by solvent extraction. Acid consumption (neutralisation) has been found to be of assistance in the subsequent extraction of copper by solvent extraction.
[0121] The pretreated pregnant leach solution can then be subjected to solvent extraction to recover one or more target metals.
[0122] FIG. 12 is a graph showing the concentration of U.sub.3O.sub.8 (mg/L) extracted into solvent from a pregnant leach solution at two solution redox potentials: 475 mV (squares) and 410 mV (diamonds). This graph illustrates the advantage of reducing solution redox prior to subjecting the pregnant leach solution to uranium solvent extraction. As can be seen, the amount of uranium that loads onto the organic phase at 410 mV is more than 3 times that which loads at 475 mV. It is believed the significant difference is due to the much lower ferric present in the PLS at the lower redox- as uranium and ferric ion tend to coload onto the organic phase, the less ferric in solution, the more uranium can be loaded. The practical consequence of a higher uranium loading onto organic phase for a given uranium concentration in the PLS, is that the volume of required organic phase, and therefore the size of the required solvent extraction plant, can be correspondingly smaller, which is a saving in capital expenditure.
[0123] FIG. 17 is a fourth embodiment 310 of a flowsheet for the heap leaching of copper and uranium ore. Similarly to the third embodiment, in the fourth embodiment, part of the heap leach is conducted under lower solution redox. The intermediate leach solution (316) arising from aerated heap leach stage 314a reports to a a number of other heap leach stages, including a pretreatment stage 314b. The pretreatment stage 314b comprises treatment of the ore material with the ILS during which the heap is not subjected to forced aeration. Accordingly, the pretreatment is conducted under reduced solution redox as compared with later stages of the leach. Because the pretreatment stage occurs near the beginning of the leach cycle, there are significant unreacted sulphide minerals which react with the ferric in the ILS converting it to ferrous. The lack of aeration creates oxygen limitation and hence there is reduced subsequent conversion of the ferrous back to ferric ions. The PLS 318 produced by the pretreatment stage 314b contains mostly dissolved uranium and some copper. The PLS is subjected to USX and CuSX 320, 322.
[0124] Because of the reasonably high acid demand (40-80 kg/T) of ores deposits in the Stuart Shelf, Australia, this translates to a high concentration of dissolved salts from the acid gangue reactions. This embodiment may result in oversaturation of the dissolved salts (eg sodium iron sulphates, such as metasideronatrite (Na.sub.4Fe.sub.2(SO.sub.4).sub.4(OH).sub.2.3H.sub.2O) in the process solutions arising from the acid gangue reactions, which may in turn result in accumulation of precipitates in the heap and poor permeability. In order to address this problem, additional salt 340 may be added to the process solution/s in order to force the precipitation of oversaturated salts, 342, which can then be removed. In addition, a purge of process liquors would be required in order to keep overall salinities of process liquors within acceptable limits.
[0125] FIG. 18 illustrates a fifth embodiment 410 of a flowsheet for the heap leaching of copper and uranium ore. The fifth embodiment relates to the second modification of the disclosed process that includes a lower salinity pretreatment step. The heap includes a pretreatment stage 414b comprising treatment of the ore material with a pre-leach solution having a relatively low chloride content, as compared with the high chloride, acidic leach liquor in the presence of an oxidant, (ie, air). The chloride content may range up to 35 g/L. The pretreatment may be conducted for a period of time sufficient for at least the majority of uranium to be leached during this step. For example, the pretreatment stage 414b may be conducted for approximately 150 days of the overall heap leach. Moreover the majority of the gangue in the ore material reacts with the acid in the pre-leach solution resulting in a ripios that is depleted in such elements as iron and calcium. There would also be partial leaching of base metals, eg copper, during the pretreatment step. The ripios from the pretreatment step is then subjected to one or more high chloride leach stages 414a during which the chloride concentration in the leach solution is increased. The high chloride leach stage/s may be conducted for a sufficient period of time (eg, approximately 300 days) for at least the majority of copper to be leached during these stages. These stages are aerated. The depletion in gangue in the pretreatment ripios means that there is less ferrous available to form salt precipitates, such as metasideronatrite, and therefore a reduced risk of heap blockage. The PLS 418 from the pretreatment stage 414b is subjected to CuSX and USX, 420, 422. The raffinate 424 is treated with additional salt, 440, if necessary, in order to cause precipitation and removal of deleterious salts 442 prior to transfer to the high chloride leach stage/s 414a of the process. Although FIG. 18 shows the pretreatment and high chloride leach stages as being separate stages, there is not necessarily physical relocation of the leached ripios between the two stages.
[0126] 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.
[0127] 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.
[0128] 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.
