MXPA97009729A - Hydrometalurgical extraction of nickel and cobalt assisted by chloride, from sulf minerals - Google Patents

Hydrometalurgical extraction of nickel and cobalt assisted by chloride, from sulf minerals

Info

Publication number
MXPA97009729A
MXPA97009729A MXPA/A/1997/009729A MX9709729A MXPA97009729A MX PA97009729 A MXPA97009729 A MX PA97009729A MX 9709729 A MX9709729 A MX 9709729A MX PA97009729 A MXPA97009729 A MX PA97009729A
Authority
MX
Mexico
Prior art keywords
process according
solution
cobalt
extractant
magnesium
Prior art date
Application number
MXPA/A/1997/009729A
Other languages
Spanish (es)
Other versions
MX9709729A (en
Inventor
L Jones David
Original Assignee
Cominco Engineering Services Ltd
Priority date (The priority date is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the date listed.)
Filing date
Publication date
Priority to US08/488,128 priority Critical patent/US5650057A/en
Priority to US08488128 priority
Application filed by Cominco Engineering Services Ltd filed Critical Cominco Engineering Services Ltd
Publication of MX9709729A publication Critical patent/MX9709729A/en
Publication of MXPA97009729A publication Critical patent/MXPA97009729A/en

Links

Abstract

A process for extracting the Ni / Co values from a mineral or concentrate, comprises the steps of subjecting the ore or concentrate to oxidation under pressure in the presence of oxygen and a solution containing halide, copper ions , and sulfate, to obtain a liquor containing Ni / Co values, from the resulting pressure oxidation paste. The liquor is subjected to a selective precipitation treatment to obtain a solid that is subjected to a Ni / Co leaching stage and separated by solvent extraction to produce suitable solutions for the electrolytic extraction of Ni and Co. The process also provides recovery of precious metals and other metals such as cob

Description

HYDROMETALURGICAL EXTRACTION OF NICKEL AND COBALT ASSISTED BY CHLORIDE. FROM SULFIDE MINERALS FIELD OF THE INVENTION This invention relates to the hydrometallurgical treatment of minerals or metal concentrates. In particular, it refers to the extraction of metals from sulphide and laterite minerals, in the presence of halogen ions, such as chloride ions.
BACKGROUND OF THE INVENTION Nickel sulfide ores are currently treated in commercial practice by a variety of processes, wherein the first step is almost always a physical concentration by flotation to refine the Ni content, typically from a scale of 0.5 per cent. percent to 2.0 percent, and up to 7 to 25 percent of Ni, as a concentrate. The subsequent treatment of this concentrate is usually pyrometallurgical (smelting), to produce a Ni matte or an artificially high-grade sulfide with about 20 to 75 percent Ni. Then the matte is refined in general to obtain nickel products by hydrometallurgical techniques. This combination of pyrometallurgical / hydrometallurgical processing of the Ni concentrates is now well established commercially with a number of variations, particularly in the hydrometallurgical portion. Most processes recover some portion of the associated metal values where they are present, such as copper and cobalt. In addition, a leaching residue containing precious metals, such as gold and silver, as well as elements of the platinum group, for example platinum and palladium, is often produced for the subsequent recovery of the contained values. This treatment scheme has some inherent drawbacks. Those associated with the pyrometallurgical step include: (i) the production of smelting gases, including SO2, which must now be treated in an acid plant to produce the byproduct of sulfuric acid, which is often difficult to trade from a remote place. (The capital and operating costs of these acid plants have an impact on the overall economies of the process). ii) Losses of nickel, and particularly cobalt, to the slag produced during smelting, are often greater than 50 percent of the cobalt input. (iii) The high costs of smelting in general, particularly for low grade concentrates (<10 percent Ni). (iv) The difficulty to treat certain concentrates with harmful elements, such as magnesium (Mg) and arsenic (As). The hydrometallurgical steps for the treatment of Ni matte vary considerably, but all known commercial processes have one or more of the following drawbacks: (i) high costs for reagents, such as caustic soda or ammonia, required for neutralization . (ii) A large byproduct production, such as ammonium sulfate or sodium sulfate, which are difficult to trade. (iii) High energy costs, due to large temperature changes during the process. (iv) A complex and costly flow diagram of the process, which leads to high capital and operating costs. As an alternative to the established pyrometallurgical / hydrometallurgical route illustrated above, there is a known process using completely hydrometallurgical steps, which treats the concentrates without melting. This uses a pressure leaching technique with an ammoniacal solution. This eliminates most of the drawbacks associated with the smelting processes, but unfortunately, it still suffers from all the aforementioned drawbacks of the known hydrometallurgical routes, and in fact, is not even as globally efficient as the best of the pyrometallurgical / hydrometallurgical routes. Copper or nickel sulfide ores often also contain other metal values, such as cobalt, as well as precious metals, such as gold and silver, and platinum group metals. Since these minerals are typically low grade minerals, in terms of copper / nickel, and also have a high proportion of sulfur to copper / nickel, the economic extraction of copper, nickel, and cobalt values has been problematic . Some sulfide minerals contain these copper / nickel values so low that the recovery of precious metals must be high in order to make the process economical. Due to the pyrite content of some minerals, the recovery of gold by conventional cyanidation is often difficult, which also makes the treatment of the ore uneconomical. The present invention provides a process for the hydrometallurgical extraction of copper, nickel, and cobalt, as well as other metals, from sulfide minerals. It also provides a process for the hydrometallurgical extraction of nickel and cobalt from laterite ores.
SUMMARY OF THE INVENTION In accordance with the invention, a process is provided for the extraction of a non-cuprous metal, from a metal ore or concentrate, which comprises the step of subjecting the ore or concentrate to oxidation under pressure. , in the presence of oxygen and an acid solution containing halogen ions, and a source of bisulfate or sulfate ions, to form a solution of the non-cuprous metal, wherein the sulfate or sulfate ion source is selected from from the group consisting of sulfuric acid and a metal sulfate that is hydrolysed in the acid solution. In addition, according to the invention, a process is provided for the extraction of Ni / Co values from a mineral or concentrate, which comprises the steps of subjecting the ore or concentrate to oxidation under pressure, in the presence of oxygen and an acid solution containing halide, copper, and sulfate ions, to obtain a liquor containing Ni / Co values from the resulting pressure oxidation paste; subjecting the liquor to a selective precipitation treatment to obtain a solid containing Ni / Co hydroxide; and subjecting the solid to a Ni / Co leaching stage with an ammonium solution to produce a leaching solution containing Ni / Co values and a residue. The process may further comprise the steps of subjecting the waste to an acid wash step, to producing a wash solution containing Ni / Co values, and a disposable waste, and recycling the wash solution to the selective precipitation treatment, or alternatively, treating the washing solution for the recovery of Ni / Co values therefrom. The selective precipitation treatment may comprise the steps of subjecting the liquor to a precipitation step at a pH of about 5 to 6 to precipitate the iron, copper, or zinc present in the liquor; and subjecting the resulting solution to a precipitation step at a pH of about 7 to 8, to obtain the solid containing Ni / Co hydroxide. The liquor containing the Ni / Co values can be obtained by subjecting the oxidation paste under pressure to neutralization, at a predetermined pH, where the copper, iron, or zinc present in the paste are in solid form , and the Ni / Co values are in solution, or in an alternative way, the liquor containing the Ni / Co values can be obtained by subjecting the oxidation paste under pressure to neutralization, at a predetermined pH, where the Ni / Co values are in solid form, and subject to the values of solid Ni / Co to an acid leaching to obtain the Ni / Co values in solution. The process may further comprise the step of controlling the concentration of nickel in the leach solution to a maximum value of about 3 to 25 grams / liter, preferably 8 to 10 grams / liter, and more preferably about 10 grams / liter . The Ni / Co leaching stage can be carried out with an ammonium sulfate solution. The concentration of the ammonium sulfate solution can be about 150 to 250 grams / liter, preferably about 200 grams / liter. Also, according to the invention, there is provided a process for the recovery of the Ni / Co values from a concentrate containing Ni / Co hydroxide, which comprises the steps of subjecting the concentrate to a leaching stage of Ni / Co with an ammonium solution, to produce a leaching solution containing Ni / Co values and a residue; and controlling the concentration of nickel in the leaching solution to a maximum value of about 3 to 25 grams / liter. The term "concentrate" in this specification refers to any material in which the content of the value of the metal has increased to a higher percentage by weight, compared to the naturally occurring mineral, and includes the artificial sulfide mineral. made by man, as it kills it, and metal values precipitated as solids, such as hydroxides and sulfides. Other objects and advantages of the invention will become clearer from the description of the preferred embodiments of the invention that follow.
BRIEF DESCRIPTION OF THE DRAWINGS The invention will now be described by way of examples, with reference to the accompanying drawings, in which: Figure 1 is a flow chart of a hydrometallurgical metal extraction process according to the invention. Figure 2 is a flowchart that gives more details about the solvent extraction steps of the process of Figure 1. Figures 3A and B show a flow diagram of an additional embodiment of the process according to the invention, for the recovery of precious metals. Figure 4 is a flow chart of a hydrometallurgical metal extraction process according to another embodiment of the invention.
DETAILED DESCRIPTION OF THE PREFERRED MODALITIES The process according to the invention is suitable for the treatment of copper ores, particularly copper sulphide ores, which also contain nickel and / or cobalt values, or nickel / cobalt sulphide ores. significant copper values, as well as nickel / cobalt oxide (laterite) ores. In addition, the process can treat nickel / cobalt minerals with other elements often considered as harmful, sas magnesium, arsenic, and zinc, or items that are valuable or whose recovery is valuable, sas precious metals, gold and silver. , and the metals of the platinum group. The mineral or feed concentrate for the process may contain one or more sulphide minerals of the base metals Cu, Ni, Co, and Zn, often combined with Fe, and sometimes with other elements sas As, Sb, Ag , etc. The sulfide minerals typical of the base metals mentioned above are: Copper: Cu2S - Calcocite, CuFeS2 - Chalcopyrite Nickel: NiS - Milerite, (Ni, Fe) 9S8 - Pentlandite Cobalt: C03S4 - Linaeite, (Co, Fe) AsS - Cobaltite Zinc: ZnS - Sphalerite (Zn, Fe) S - Marmatite The proportion of the metal: sulfur in this context is the ratio of the total base metals (Cu, Ni, Co, Zn) to the sulfur in the concentrate, and this is a measure of the degree of the concentrate. Typically, the metal: sulfur ratio varies from 1.5 for high grade concentrates, down to 0.2 for low grade concentrates. For concentrates that are predominantly Ni / Co, the metal: sulfur ratio is most frequently in the lower part of the range, from 0.2 to 0.8 (the Fe is specifically excluded from this calculation, even when present in virtually all sulfur concentrates). The significance of the proportion of sulfur metal to the process is that it affects the metallurgy that occurs during the initial operation of the pressure oxidation. The different embodiments of the process according to the invention can be used to treat a range of Ni / Co concentrates, where the proportion of metal: sulfur varies from low to high, as illustrated above. However, in addition to this proportion, there is another important feature that must be taken into account. The degree of oxidation of sulfur (in sulphate) during oxidation under pressure. The sulfur contained in the concentrate is converted during oxidation under pressure, either in elemental sulfur (S °) (there is no oxidation of sulfur), or it is oxidized in sulfate (SO 4"). 70 to 95 percent of the sulfur, and is produced as elemental sulfur, expressed otherwise, the oxidation of sulfur (in sulfate) normally varies from 5 to 30 percent.It is considered beneficial to minimize the oxidation of sulfur, and is an objective This is facilitated by the introduction of a sulphate or bisulfate source, sas H2SO4, in the oxidation stage under pressure.The meaning of sulfur oxidation, is that it produces acid, which eventually neutralize, and this affects the distribution of Cu, Fe, and other elements in the pulp product of pressure oxidation.The higher acid pastes (low pH) contain Cu in solution, while the lower acid pastes (high pH) have Cu in solid form, such as basic copper sulfate. For concentrates with a low proportion of sulfur metal, and / or a high sulfur oxidation, the general case is the process flow diagram shown in Figure 1. This is referred to as Mode C. Sufficient acid is produced during Oxidation under pressure 12, so that it is necessary to neutralize this acid by means of slaked lime in the last stages of the autoclave. This is indicated as the neutralization 501 in Figure 1. Without this neutralization, the product paste would have a low pH, resulting in significant Fe in solution, and almost all Cu as well. It is an important feature of the process that pasta product contains Fe minimum in solution (less than 100 ppm), and approximately 1 to 5 grams / liter in solution. By adjusting the amount of slaked lime added in the 501 neutralization, these objectives can be achieved, even with concentrates having a low metal: sulfur ratio, and exhibiting a relatively high sulfur oxidation, for example, 15%. at 30 percent. A typical example of this type of concentrate is a type of pentlandite / pyrite mineral assembly. However, for concentrates having a high proportion of metal: sulfur, and / or a low sulfur oxidation, the total amount of acid produced during pressure oxidation 12 is less, and 501 neutralization may not be required to achieve a paste product with a low content of Fe and Cu in the desired range. This mode of the process is called Mode A, and is described below with reference to Figure 4. A typical example of this type of concentrate is a type of pentlandite / chalcopyrite / pyrrhotite mineral assembly. In Mode A, the amount of acid consumed during oxidation under pressure by other chemical reactions is more than sufficient to use all the acid produced by the oxidation of sulfur. The examples of Mode A and Mode C, required for two different concentrates, are shown in the following table: DB TYPE CONCENTRATE TEST1,% OXIDATION PROPORTION OF METAL PROCESS: SULFUR SULFUR,% Cu Ni Co S MODE A 6.3 14 0.6 34 0.61 6 MODE C 0.1 22 0.6 29 0.78 15 Accordingly, the first concentrate with 14 percent Ni exhibited only 6 percent oxidation of S in the oxidation under pressure, and therefore, it was treated by Mode A, while the second concentrate required the C, due to the higher oxidation of S (15 percent). Now Process Mode C will be described with reference to Figure 1. First, the ore or concentrate is subjected to oxidation under pressure 12 in an autoclave, in the presence of an acid oxidation containing sulfate, chloride, and copper ions. . In the present example, the amount of H2SO4 introduced into the autoclave is about 40 grams / liter, and the concentration of the chloride in solution is about 10 to 12 grams / liter. Typically the temperature is from about 90 ° C to about 'i * 160 ° C, under a partial pressure of oxygen of approximately 200 to 2,000 kPa. The retention time is approximately 0. 5 to 5.0 hours, depending inversely on the temperature, and the process is normally carried out in a continuous manner in the autoclave. However, the process can also be done in a batch form, if desired. Neutralization 501 is effected by pumping slaked lime towards the last one or two compartments, on the outlet side in the autoclave, with approximately 10 to 20 percent solids in water. After oxidation under pressure 12, the paste produced in the autoclave is discharged through one or more evaporation tanks 22, to reduce the pressure to atmospheric pressure, and the temperature to 90-100 ° C. The slurry is then further cooled, and subjected to filtration 24 to produce a pressure oxidation filtrate 29, and a solid residue (filter cake of oxidation under pressure). The neutralization step 501 is used to precipitate the soluble copper in the oxidation filter cake under pressure, which would otherwise be reported to the oxidation filtrate under pressure 29. Accordingly, the neutralization 501 can be used to minimize the copper in filtrate 29, typically down to 1 to 5 grams / liter of copper, which makes the subsequent removal of the copper from the solution easier. In addition, the neutralization 501 helps to minimize the Fe in the oxidation filtrate under pressure 29. However, when adding slaked lime, it is preferable not to add too much to precipitate the Ni / Co. Typically, the addition of slaked lime so that the pressure oxidation filtrate 29 has a pH of between about 3 and 4, has been found to be adequate to remove most of the copper, and yet minimize the Ni precipitation. /Co. The pressure oxidation filtrate 29 is generally subjected to a solvent extraction 50, particularly if there are significant copper values present in the original concentrate, to recover the copper values and reduce [Cu2 +] in the raffinate 63 as much as possible , typically up to less than 100 ppm. In addition, the filter cake of the oxidation under pressure is subjected to an atmospheric leaching 14 to recover the copper in solution, which solution is subjected to a solvent extraction of Cu 16. The leaching 14 is carried out with the raffinate 120 from the solvent extraction of Cu 16, which is acid diluted to approximately 3 to 20 grams / liter of H2SO4. In addition, the leaching 14 helps to wash the pooled solution containing any Ni / Co values, out of the filter cake of the oxidation under pressure. These values, which will accumulate in stream 51, can be recovered on a flea basis (say 1 to 10 percent of the flow, depending on the concentration), by precipitation at a pH of 7 to 8 with slaked lime as hydroxides of Ni / Co, similar to 506 precipitation conditions, described later. The mixed Ni / Co hydroxide can then be filtered and recycled to a purification step 500, described below. The paste 31 resulting from the leaching 14 is difficult to filter, and the liquid / solid separation is effected by means of a series of thickeners in a counter-current decanting (CCD) configuration 34. The washing water is provided for a portion of the raffinate from the extraction with solvent 16, which is divided into 36, and neutralized in 46 using limestone to remove the acid. The paste from the neutralization 46 is filtered at 48, to produce a gypsum residue, and the liquid 51 is recycled as wash water. The extractant charged from the extractions with solvent 50 and 16, is subjected to separation 44, and then sent to the electrolytic extraction of copper 20. The solvent extractions of Cu 50 and 16, are operated with a common extractant. This is shown in Figure 2, where the dotted line indicates the organic extractant that is circulating after separation 44. The separation 44 is carried out with spent acid or electrolyte 55 from the electrolytic extraction 20, to obtain a sulfate solution of pure copper, or pregnant electrolyte 59, which is then passed to the electrolytic extraction stage 20. Any suitable copper extractant that can selectively remove Cu from an acid solution that also contains Ni / Co / Zn / Fe can be used. . A suitable extractant is a hydroxyoxime, such as the LIX 84MR or LIX 864MR reagents from Henkel Corporation. If there are no significant copper values present in the ore or in the concentrate, however, it is beneficial to perform oxidation under pressure 12 in the presence of copper ions (for example, from 5 to 10 grams / liter of Cu). Copper ions can be added in the form of a copper salt, such as CuSO4 or CuCl2. Subsequently, the solvent extraction of Cu and the separation are still carried out, but the electrolytic extraction 20 will be omitted, and the pregnant copper liquor resulting from the separation 44 of the organic extractant will be recycled towards oxidation under pressure 12. In an alternative way , a copper concentrate can be added, in which case, the copper can be recycled after the extraction with Cu solvent and the separation, or it can be sent to the electrolytic extraction for the recovery of the copper. This will also be the case if a laterite mineral is being processed. The raffinate 63 is subjected to a purification step 500, to prepare a Ni / Co solution free of elements such as Fe, Zn, and Cu, which cause difficulty in the following steps of the solvent extraction process and the electrolytic extraction. of Ni and Co. The purification step 500 is a precipitation step wherein residual Cu, Fe, and Zn are precipitated by the addition of slaked lime and recycled Mg (0H) 2. Typically, the feed solution to purification stage 500 will contain copper and iron, as well as any zinc and magnesium present in the concentrate. The precipitation 500 is carried out at a pH of about 5 to 6, such that, ideally, no more than about 1 ppm Zn, 1 ppm Cu, and 1 ppm Fe remain in the solution. It is also important not to precipitate too much Ni / Co. This is achieved by careful control of the pH, that is, by not allowing the pH to rise too high. It has been found that recycled Mg (0H) 2 is beneficial in this respect. The product from precipitation 500 is subjected to a liquid / solid separation 502. Cu, Fe, and Zn, which precipitate as hydroxides, can be further processed by dilute acid washing or 503 leaching, particularly for the recovery of Ni / Co. The product from the 503 acid wash is subjected to a liquid / solid separation 505, which mainly leaves Cu, Fe, and Zn hydroxides, which provides an outlet for the zinc from the system. The liquid 504 from the liquid / solid separation 505, it is recycled towards oxidation under pressure 12.
If the Zn content is sufficiently high, the Cu / Fe / Zn hydroxide can be further leached with dilute acid to selectively recover the zinc. In an extreme case, a case of extraction with zinc solvent may be included if desired. The concentrations of Ni, Co, and Mg in solution after precipitation 500, will depend on the composition of the concentrate. Depending on the mineralogy, it is possible that most of the magnesium in the concentrate is leached, during oxidation under pressure 12. Therefore, for the Mi concentrate containing, say, 20 percent nickel and 5 percent of magnesium, the typical solution after precipitation 500 will be about 30 grams / liter of nickel, and about 6 grams / liter of magnesium. The magnesium content will be higher in the case of a laterite mineral. The solution resulting from the liquid / solid separation 502 is subjected to a selective precipitation step 506, where Ni and Co are precipitated as hydroxides or carbonates, with a suitable neutralizing agent, such as slaked lime (Ca ( 0H) 2), soda ash (Na2C03), ammonia or caustic soda (NaOH). This is effected at a pH of about 7 to 8, while the precipitation of Mg (OH) 2 is minimized. A preferred neutralizing agent is slaked lime, "because of its relatively low cost, and because the reaction does not introduce new cations, such as Na + and NH4 +, into the liquor.
Neutralization with Lime Off NiS04 (aqueous) + Ca (0H) 2 (s) + CaS04 »2H20 (s) (1) (gypsum) A similar reaction occurs with CoS04, and MgSO4, producing Co (0H) 2 and Mg (OH) 2, respectively.
Neutralization with Caustic Soda (NaOH) NiS04 (aqueous) + NaOH? Ni (OH) 2 (s) '+ NaS04 (aqueous) (2) However, it is important to have some Mg present in the precipitated solid, which facilitates the separation of Ni and Co, as will be described later. A sequence of two-stage countercurrent precipitation has been found beneficial. In some circumstances, precipitation with caustic soda or ammonia is desirable, for example, that does not produce a solid by-product (gypsum), so that the Ni precipitate is of a higher degree, and is free of calcium.
The product of the precipitation step 506 is subjected to a liquid / solid separation 508. The liquid from the liquid / solid separation 508 is subjected to a precipitation step 510, preferably again with slaked lime, for the same reasons above , to precipitate additional Mg, if necessary, to prevent in this way the accumulation of Mg in the system. The product from the precipitation step 510 is subjected to a liquid / solid separation 512. The solid from the separation 512 is a byproduct of magnesium hydroxide 514. As indicated above, some of the byproduct of magnesium hydroxide 514 is recycled to used in the precipitation 500. The liquid from the separation 512 is recycled to the oxidation under pressure 12, as indicated by the recycle stream 516. The solid hydroxide cake from the separation step 508, which contains the values of Ni and Co, is subjected to a leaching 518 with an ammonium solution, at a pH of about 6 to 8. The ammonium solution can be ammonium sulfate or ammonium carbonate, but it has been found that the former is superior, because has a lower pH, thus allowing a better separation of Co to Ni in solution. In addition, ammonium sulfate has a lower ammonia vapor pressure (gas), and also, Ni / Co extractions are higher with ammonium sulfate. In the present example, an ammonium sulfate solution of 200 grams / liter is used. The reactions that take place during leaching 518, where nickel and cobalt diamine sulphates are formed, are as follows: (NH4) 2S04 + Ni (OH) 2? Ni (NH3) 2S04 + 2H20 (3) (NH4) 2S04 + Co (0H) 2? C (NH3) 2S04 + 2H20 (4) The Mg present in the solid also dissolves, as follows: (NH4) 2S04 + Mg (0H) 2? MgS04? 2H20 + 2NH3 (5) In the embodiment of leaching 518, it is not a matter of leaching 100 percent of the Ni / Co values in the solid, but of about 90 to 99 percent. This makes it possible for leaching 518 to be carried out at a low pH instead of a pH higher than about 9, which would be what would be otherwise required. This higher pH requires the addition of ammonia to the leach as a second reactant with ammonium sulfate. A further problem that arises is that the Co-known or commercially available extractant does not function effectively at this high pH value. The extractant degrades and is not selective against Ni. As a result, it is necessary to first perform Ni extraction, rather than Co extraction, which would then require the reduction of the pH by the addition of an additional reagent such as an acid, which in turn would result in the production of the by-product. ammonium sulfate, and the consumption of the ammonia reagent. Another problem that arises is that, in order to carry out the extraction with Ni solvent first, it is necessary to first oxidize all the Co up to the +3 oxidation state, to avoid the extraction of Co with Ni. This oxidation is difficult to achieve in a quantitative way. Consequently, this results in more complications of the process. It is also necessary to reduce Co3 + back to Co2 +, followed by Ni extraction, and this is equally difficult to achieve. To avoid the above difficulties, the process according to the present invention provides for leaching 518 at a pH of about 6 to about 8, and then subjecting the resulting solid to a subsequent washing step with a solution of dilute ammonium sulfate, as will be described later. An additional aspect of the process is that the concentration of nickel ions in solution during leaching 518 is controlled to remain at a relatively low value of about 10 grams / liter at most. It has been found that this results in a better recovery of Ni during leaching 518. With the amount of Ni present in the known solid, the appropriate volume of liquid required to reach the desired Ni concentration can be calculated. The product from the leach 518 is subjected to a liquid / solid separation 522. The liquid from the separation 522 is subjected to a solvent extraction of Co 534, to provide an extractant laden with Co, and a raffinate which is then subjected to to a solvent extraction of Mg 536, to provide an extractant loaded with Mg, and a raffinate which is subjected to extraction with Ni 538 solvent, to provide an extractant loaded with Ni and a raffinate. Raffinate from Ni 538 solvent extraction is recycled to leach 518. The solid product from liquid / solid separation 522 is subjected to re-wetting or washing step 520 as indicated above, where the solid is washed with a solution of ammonium sulfate. This is a weak ammonium sulfate solution of about 10 percent of the concentration of the 518 leach solution. It results from washing the ammonium sulfate solution tucked out of the solid in the washing step 520.
The product from re-boost step 520 is subjected to a liquid / solid separation 524, and the solid is washed with water. The washing water and the liquid from the liquid / solid separation 524 are subjected to a solvent extraction of Co 526, to again provide an extractant loaded with Co and a raffinate, which is subjected to solvent extraction of Mg 527, to provide an extractant loaded with Mg and a raffinate, which is subjected to an extraction with Ni 528 solvent to provide a Ni-charged extractant and a final raffinate which is recycled to the re-boost step 520. To compensate for the water added during the washing with water in separation 524, there is a purge of the final raffinate until the strong ammonium sulfate raffinate that comes from the solvent extraction of Ni 538. For this purpose, the strong ammonium sulfate circuit includes an evaporation step 539 to compensate the purge of the raffinate from the weak ammonium sulfate raffinate. The solvent extractions of Co 534, 526, solvent extractions of Mg 536, 527, and solvent extractions of Ni 538, 528, respectively, are all operated with a common extractant, as is the case with solvent extractions. of Cu 50, 16. An extractant which has been found to be suitable for both Co and Mg extraction, is an organic phosphorus acid extractant, more specifically an organic phosphinic acid based extractant, such as Cyanex 272MR, Cyanamid Inc., which comprises bis-2,4,4-trimethylpentylphosphinic acid. For the extraction of Ni, it has been found that an extractant based on hydroxyoxime, such as LIX 84MR, from Henkel Corp is suitable. The extractants loaded with Co, Ni, and Mg respectively, are purified with suitable aqueous solutions, to remove the solution of ammonium sulfate methylated, and then separated with dilute acid to produce pure Co and Ni pregnant solutions, and a Mg pregnant liquor containing small amounts of Co and Ni. The solutions of Co and Ni are sent to the electrolytic extraction stages of Co and Ni 530 and 532, respectively. Before separation, the extractant loaded with Co is purified with a concentrated solution of Co that is separated from the pregnant solution of Co that goes to the electrolytic extraction of Co, and / or a concentrated solution of Mg that is separated from the pregnant liquor of Mg. This is to facilitate the removal of the Ni values that may be present in the extractant loaded with Co. In the same way, the extractant loaded with Mg can be purified with a concentrated Mg solution that is separated from the Mg pregnant liquor. For a good separation of Co from Ni during Co solvent extraction and Ni solvent extraction, it has been found that it is beneficial to have some Mg present in the feed in solution towards solvent extraction from Co. Typically, the Analysis of the solution has the same ratio of Co to Ni found in the original feed concentrate (commonly 1:30). Therefore, for 10 grams / liter of Ni, 0.33 grams / liter of Co. is typical. The same extractant is used for both solvent extractions of Co and Mg 534 and 536. The extractant is more selective for Co than for Mg , and more selective for Mg than for Ni. During the solvent extraction of Co 534, the amount of extractant used is limited to occupying all the available sites with Co ions, to a greater degree, and with Mg ions, to a better degree, which counteracts the extraction of Ni . During solvent extraction of Mg 536, the available sites are filled mainly with Mg ions, and to a lesser degree with some Co ions, and possibly also a small amount of Ni ions. The Ni and Co ions are then recovered by recycling Mg pregnant liquor to the precipitation of Ni / Co 506, as indicated by arrow 543. In addition, it has been found that it is beneficial to maintain an Mg concentration approximately equal to the concentration of Co, although this can vary very widely from say, 1: 5 to 5: 1. The benefit of having Mg present is that: (i) it minimizes the amount of Ni that is extracted during the extraction with Co solvent, while it allows (ii) an extraction of a high percentage of Co, that is, greater than 90%. percent, and (iii) a high proportion of Co to Ni in the product of Co, that is, Co: Ni > 1000: 1 Without Mg present, some compromise must be reached in the extraction with Co solvent, where: (i) some Co is extracted from Ni with Co, and (ii) Co extraction is incomplete, or (iii) the proportion of the Co to Ni in the Co product is too low. With Mg present, some Co (ie, 5 to 10 percent) can be left without extracting during Co solvent extraction, and instead, it will be extracted during extraction with Mg solvent. The products of Mg solvent extraction are: (a) the pregnant liquor from the separation containing some Mg, Ni, and Co, which is recycled and not lost, -and (b) the Mg raffinate with very low levels of Co, that is, approximately 1 ppm, which allows the following extraction with Ni solvent to produce a very good proportion of Ni to Co in the Ni pregnant liquor going to the Ni electrolytic extraction. Consequently, Ni cathodes and very pure Co cathodes are obtained. The solid from the liquid / solid separation 524 is washed (540) with dilute acid, to recover the Ni / Co met, which is recycled to the precipitation 500. The solid residue after the liquid / solid separation 542 is discarded . A suitable temperature scale for the leaching of Ni / Co 518 and solvent extractions of Ni / Co, it has been found to be from about 30 ° C to 60 ° C, preferably from about 40 ° C to about 50 ° C. Turning now to Figures 3A and B, the recovery of precious metals, such as gold and silver, will be described. This process involves the treatment of the stream of the final residue 35 of Figure 1. Precious metals do not leach during the oxidation stage under pressure., but remain in the solid residue that remains after the atmospheric leaching step 14. In order to facilitate the recovery of the precious metal, evaporation 22 is carried out from the oxidation step under pressure 12 in two stages. The first stage is a temperature slightly higher than the freezing point of the elemental sulfur, ie, from about 120 ° C to 130 ° C, with a corresponding vapor pressure of about 50 to 150 kPa. The preference process is carried out in a continuous mode, the retention time being in the first evaporation stage of approximately 10 to 30 minutes. The second evaporation stage is at atmospheric pressure, and at approximately 90 ° C to 100 ° C, with a retention time of at least 10 minutes. This allows the elemental sulfur, which is still melted in the first evaporation stage, to become one of the solid phases, such as the stable orthorhombic crystalline phase. This procedure facilitates the production of clean crystals of elemental sulfur, which is important for the recovery of precious metals from the residue of leaching. The leaching residue 35 now produced by atmospheric leaching stage 14 contains, in addition to the precious metals, hematite, crystalline elemental sulfur, unreacted sulfides (pyrite), and any additional products that may result from the particular concentrate being using, for example, gypsum, and iron hydroxides. It is believed that the gold in residue 35 is largely untouched by the process, since it is most likely in the native state. However, the silver is oxidized in the oxidation step under pressure 12, and is probably present as a silver salt, such as silver chloride or silver sulfate.
It has been found that conventional cyanidation does not leach the gold well from residue 35. It is believed that this is due to the encapsulation of gold into mineral particles, such as pyrite. However, gold can be released by the oxidation under pressure of these minerals, referred to as "total oxidative leaching". In order to effect this leaching without oxidizing the elemental sulfur also contained in the residue 35, the process comprises the step of removing as much elemental sulfur as possible. First, by virtue of the evaporation of two stages, sulfur crystals of good quality are produced. Second, the leach residue 35 is subjected to a flotation of foam 402, to produce a sulfur-rich flotation concentrate 404 and a depleted flotation tail of sulfur 406. The tail 406 is subjected to a solid / liquid separation 408 , to produce a liquid, which is recirculated to a conditioning tank 410 upstream of the flotation passage 402, and a solid 412 that is sent to the total oxidative leaching stage 414. The flotation concentrate 404 is filtered (416) , and dried to a low humidity in a dryer 418. Then the product is subjected to a leaching step of sulfur 420 with a sulfur extractant. Any suitable sulfur extractant, such as perchlorethylene (PCE) or kerosene, can be used. In the present example, hot perchlorethylene is used. The slurry from the leach 420 is filtered 422, and the resulting liquid is subjected to cooling 424 to produce S ° crystalline, and then filtered (425). The cooled sulfur can be subjected to an optional sulfur purification step (not shown), to remove impurities, such as selenium and tellurium, from it. The solid sulfur is dried in a dryer 426 to produce a product of sulfur 428. The liquid from filtration 425 is recycled to the leaching of hot perchlorethylene 420. The solid residue from filtration 422 is dried in a dryer 430. The resulting product , which is a low sulfur residue 432, is sent to the total oxidative leaching 414. The perchlorethylene vapors from cooling 424 and dryers 426 and 430 are recycled to the leaching of hot perchlorethylene 420 by means of a condenser 434. A test was conducted where 100 grams of atmospheric leaching residue 14 containing 25.1 percent elemental sulfur (S °) and 3 percent sulfur were processed through flotation 402 and 420 leaching. 73.8 grams of sulfurized waste (feed material for total oxidation leaching 414) containing 1.9 percent S °, and 4.1 percent sulphide, and That is to say, a total of 6 percent of total sulfur. The desulphurized residue containing 5.9 percent of the elemental sulfur (S °) in the original leaching residue, that is, 94.1 percent, was recovered in a pure elemental sulfur product. Total oxidative leaching 414 is carried out at approximately 200 ° C-220 ° C, and a partial oxygen pressure of 200 to 2000 kPa, sufficient to completely oxidize all of the sulfur and metal compounds to the highest valencies, respectively. Therefore, all the sulfur and pyrite are oxidized to obtain sulfate. Oxidation is conducted under acidic conditions, such as with the acid produced at the site. If sufficient pyrite is present, the reaction is highly exothermic, and in general, the desired operating temperature can be achieved. Typically, approximately 10 percent of the total oxidizable sulfur will be sufficient, with a normal percentage of solids in the feed pulp. After total oxidative leaching 414, the paste is subjected to neutralization 437 at a pH of 2 to 3, with limestone, and then subjected to a liquid / solid separation 438 by means of a countercurrent settling circuit. (CCD), to obtain a solid containing precious metals and a liquid 13 that may contain base metal values, such as copper. The liquid 13 can be combined with the liquid (stream 33) going to the extraction with solvent 16, for the recovery of copper, as indicated in Figure 1. A portion of the neutralized stream 51 (Figure 1) of the raffinate from solvent extraction of Cu 16, is divided into 49, and the resulting stream 53 is partially used (approximately 80 percent) as wash water in the liquid / solid separation 438, and partially recycled (approximately 20 percent). one hundred) to total oxidative leaching 414, as indicated in Figure 3B. The precious metal recovery circuit of Figures 3A and B is indicated by block 155 in Figure 1. Prior to cyanidation 444, solids from separation 438 may be subjected to an optional slaked lime boiling step 443 , to facilitate the recovery of silver during 444 cyanidation, by decomposing the silver jarosite compounds formed during total oxidative leaching 414. Precious metals are in the remaining solids after separation 438. Now that they have decomposed pyrite and other encapsulating minerals of the original concentrate, the precious metals are susceptible to cyanidation 444. In the cyanidation step 444, the solids are leached with NaCN under alkaline conditions. In order to do this, the solids are formed in a paste with a cyanide solution, to form a paste with a solids content of 30 to 40 percent. Additional NaCN and slaked lime are added as required to maintain a minimum concentration of NaCN from about 0.2 to about 0.5 grams / liter of NaCN, with a pH of about 10. The temperature is ambient, and a time of Retention of approximately 4 to 8 hours in continuous operation mode. Both gold and silver are reported in a high yield to the cyanide solution, and are typically recovered through the established pulp carbon circuit process, by which activated carbon is added to the cyanide paste to absorb the precious metals , without the need for filtration. Charged coal, now rich in precious metals, is separated by sifting (445), and the pulp of the auger is discarded towards the tail. Charged coal is treated by established methods to recover the content of precious metals by means of a leaching / electrolytic extraction / smelting process (447). The product is usually Dore metal, which contains both gold and silver, which is sent to a 449 gold refinery for the final separation of gold from silver. The coal of auger from a step of regeneration of coal 451, after the recovery of precious metals, is recycled to the carbon circuit in pulp 444. The global recovery of precious metals through the total process, is generally greater than 90 percent, and under optical conditions, approaches 99 percent. A test was performed where a desulfurized residue was processed in a total oxidative leaching 414 at 200 ° C for 2 hours under an oxygen pressure, and then depressurized and cooled to room temperature. The resulting paste was neutralized to a pH of 3 with limestone, and then filtered. Then the filter cake was leached with a cyanide solution under conventional conditions to leach the gold and silver. The extraction of gold after total oxidative leaching 414 and cyanidation 444, was 97 percent, with only a consumption of 1.0 kilogram / ton of NaCN. In comparison, the extraction of gold on a residue that had not been oxidized in total oxidative leaching 414 was only 34 percent, and the cyanide consumption was extremely high at 19.0 kilograms of NaCN / ton. Figure 4 is a flow diagram of Mode A. The steps that correspond to those of the modality of Figure 1, have the same reference numerals. The process comprises a pressure oxidation step of 12, where the sulphide minerals of the concentrate or ore are oxidized by high pressure oxygen, followed by a liquid / solid separation (e.g., filtration) 24, which produces a solid (pressure oxidation filter cake) 25, and oxidation filtration under pressure 29. The solid 25 contains all or almost all of the copper content of the feed concentrate, and is treated for the recovery of copper by leaching of acid, extraction with solvent, and electrolytic extraction, as in the embodiment of Figure 1, thus producing high quality copper cathodes, and a residue 35 which may contain precious metals. The residue 35 can be treated for recovery of the precious metal, as described with reference to Figures 3A and B above. This is indicated by block 155 of Figure 4. Filtrate 29 is purified at 500 to remove the detrimental elements, such as Cu, Fe, and Zn, by neutralization with slaked lime to a pH of about 6, as described with reference to Figure 1, producing a purified solution 36, after filtration, containing Ni, Co, and other elements such as Mg, which may be present in the feed concentrate. Solution 36 is treated for the recovery of Ni / Co, as described with reference to Figure 1. This is indicated by block 38 in Figure 4. Solution 39 produced at 38 is recycled back to oxidation under pressure 12, to complete the cycle, as before (stream 516 in Figure 1). Although only preferred embodiments of the invention have been described in detail herein, the invention is not limited thereto, and modifications may be made within the scope of the appended claims.

Claims (75)

  1. CLAIMS 1. A process for extracting Ni / Co values from a concentrated mineral, which comprises the steps of: subjecting the ore or concentrate to oxidation under pressure in the presence of oxygen and an acid solution containing halide ions , copper, and sulfate, to obtain a liquor containing Ni / Co values, from the resulting pressure oxidation paste; subjecting the liquor to a selective precipitation treatment to obtain a solid containing Ni / Co hydroxide; and subjecting the solid to a leaching stage of Ni / Co with an ammonium solution, to produce a leaching solution containing Ni / Co values and a residue. A process according to claim 1, which further comprises the steps of: subjecting the waste to an avid washing step to produce a wash solution containing Ni / Co values, and a disposable waste; and recycling the wash solution towards the selective precipitation treatment. 3. A process according to claim 1, which further comprises the steps of: subjecting the waste to an acid wash step to produce a wash solution containing Ni / Co values, and a disposable waste; and treating the washing solution for the recovery of the Ni / Co values therefrom. A process according to any of the preceding claims, wherein the selective precipitation treatment comprises the steps of: subjecting the liquor to a precipitation step at a pH of about 5 to 6 to precipitate iron, copper, or the zinc present in the liquor, - and subjecting the resulting solution to a precipitation step, at a pH of about 7 to 8, to obtain the solid containing Ni / Co hydroxide. 5. A process according to any of the preceding claims, wherein the liquor containing the Ni / Co values is obtained by subjecting the oxidation paste under pressure to a neutralization, at a predetermined pH, where the copper , the iron, or the zinc present in the paste, are in solid form, and the Ni / Co values are in solution. 6. A process according to any of claims 1 to 4, wherein the liquor containing the Ni / Co values are obtained by subjecting the oxidation paste under pressure to a neutralization, at a predetermined pH, where the Ni / Co values are in solid form, and subject to the values of solid Ni / Co to an acid leaching to obtain the Ni / Co values in solution. A process according to claim 2 or 3, which further comprises the step of subjecting the residue to a washing step, before the acid washing step, to produce a second washing solution containing Ni / values. Co, and a residue that is subjected to the acid wash step. A process according to claim 7, which further comprises the step of subjecting one or both of the leaching solution and the additional washing solution containing Ni / Co values, to a solvent extraction, to recover the Ni / Co values of the same. 9. A process according to claim 8, wherein the extraction with solvent is carried out with a nickel extractant to produce a solution containing nickel. 10. A process according to claim 8, wherein the extraction with solvent is carried out with a cobalt extractant to produce a solution containing cobalt. 11. A process according to claim 8, wherein the solvent extraction comprises the steps of: effecting a cobalt solvent extraction in the presence of magnesium ions, using a cobalt extractant, to produce a charged cobalt extractant with cobalt ions, and a first raffinate containing nickel and magnesium ions in solution, - make a solvent extraction of magnesium on the first raffinate, using a magnesium extractant, to produce a magnesium extractant loaded with magnesium ions and cobalt, and a second raffinate; and extracting with nickel solvent on the second raffinate, using a nickel extractant, to produce an extractant loaded with nickel and a third raffinate. A process according to claim 11, wherein each of the leaching solution and the additional washing solution is subjected to solvent extraction, the third raffinate being recycled from the solvent extraction of the leaching solution to the Ni / Co leaching stage, and the third raffinate being recycled from the solvent extraction of the additional washing solution, to the washing step, before the acid washing step. 13. A process according to claim 12, wherein the solvent extractions of Co, Mg, and Ni, are carried out with common extractants of Co, Mg and Ni. A process according to claim 11, which further comprises the step of separating extractants charged with cobalt and nickel, to produce solutions of cobalt and nickel, respectively. 15. A process according to claim 14, wherein, before separation, the extractant charged with cobalt is purified with a concentrated solution of cobalt which is recycled from the separation. 16. A process according to claim 11, wherein the Ni / Co leaching step, and solvent extractions of cobalt, magnesium, and nickel, are performed at a temperature of about 30 ° C to about 60 ° C. . 17. A process according to claim 16, wherein the Ni / Co leaching, and the solvent extractions of cobalt, magnesium, and nickel, are carried out at a temperature of about 45 ° C to about 50 ° C. 18. A process according to claim 4, which further comprises the step of subjecting the cobalt and nickel solutions to electrolytic extraction, to recover the cobalt and nickel thereof. 19. A process according to claim 11, which further comprises the step of: separating the magnesium extractant to produce a pregnant solution containing magnesium and cobalt ions; and recycling the pregnant solution towards the selective precipitation treatment. 20. A process according to claim 19, wherein, before separation of the magnesium extractant, it is purified with a concentrated magnesium solution comprising the pregnant solution, which is recycled from the separation of the magnesium extractant. 21. A process according to claim 11, wherein the cobalt extractant is similar to the magnesium extractant, the extractant being more selective for cobalt than for magnesium. 22. A process according to claim 8, wherein the solvent extraction comprises the steps of: performing a solvent extraction of cobalt at a pH of about 6 to 8, to produce a cobalt solution and a first raffinate; and effecting a nickel solvent extraction on the first raffinate at substantially the same pH as extraction with cobalt solvent, to produce a nickel solution and a second raffinate. 23. A process according to any of the preceding claims, which further comprises the step of controlling the concentration of nickel in the leaching solution to a maximum value of about 3 to 25 grams / liter. 24. A process according to claim 23, wherein the maximum value is from about 8 to 15 grams / liter. 25. A process according to claim 24, wherein the maximum value is about 10 grams / liter. 26. A process according to any of the preceding claims, wherein the selective precipitation treatment is carried out in the presence of magnesium ions, to produce a solid containing magnesium in addition to the Ni / Co hydroxide, and a resulting solution It contains magnesium. 27. A process according to claim 26, which further comprises the step of subjecting the resulting solution to a precipitation stage., to produce a precipitate of magnesium hydroxide and a magnesium-depleted raffinate. 28. A process according to claim 27, which further comprises the step of recycling the spent magnesium raffinate to oxidation under pressure. 29. A process according to claim 27 or 28, which further comprises the step of recycling at least part of the magnesium recycle to the selective precipitation treatment. 30. A process according to any of the preceding claims, wherein the Ni / Co leaching stage is carried out with an ammonium sulfate solution. 31. A process according to claim 30, wherein the ammonium sulfate solution has a concentration of about 150 to 250 grams / liter. 32. A process according to claim 31, wherein the ammonium sulfate solution has a concentration of about 200 grams / liter. 33. A process according to any of claims 1 to 29, wherein the Ni / Co leaching stage is carried out with an ammonium carbonate solution. 34. A process according to any of claims 1 to 29, wherein the Ni / Co leaching stage is carried out with a mixture of ammonium sulfate and ammonium carbonate in solution. 35. A process according to any of the preceding claims, wherein a copper salt or a concentrate is added to the acid solution in oxidation under pressure, to provide the copper in the acid solution. 36. A process according to claim 35, wherein the copper is recycled toward oxidation under pressure. 37. A process according to any of the preceding claims, which includes the step of subjecting the liquor from oxidation under pressure to extraction with copper solvent, before the selective precipitation treatment. 38. A process according to claim 1, wherein the halide is chloride. 39. A process according to any of claims 1 to 34 or claim 38, wherein the ore or concentrate comprises a sulfide concentrate that also contains copper values, and which further comprises the steps of leaching the pulp from the oxidation under pressure or the solid obtained therefrom, with an acid sulfate solution, to produce a leaching liquor containing copper sulfate in solution and a leaching residue; subjecting the leach liquor to a solvent extraction of copper to produce a concentrated copper solution and a copper-depleted raffinate; and recycling at least part of the spent copper raffinate to the acid sulfate leaching step. 40. A process according to claim 39, which further comprises the step of subjecting the liquor to a copper solvent extraction prior to the selective precipitation treatment, to produce an additional concentrated copper solution. 41. A process according to claim 40, which further comprises the step of subjecting the concentrated copper solutions from extractions with copper solvent, to electrolytic extraction, to produce a copper metal. 42. A process according to any of the preceding claims, wherein the ore or concentrate comprises a sulphide concentrate that also contains precious metals, and which further comprises the steps of: removing the elemental sulfur from the pressure oxidation paste or a solid obtained from it, to obtain a low sulfur residue, and to subject the low sulfur residue to an oxidative leaching at an elevated temperature and pressure, to oxidize the sulfur and the precious metal compounds present in the low residue. in sulfur, to produce a waste for the extraction of precious metals from it. 43. A process according to claim 42, wherein the removal of sulfur comprises the steps of: subjecting the oxidation paste under pressure or the solid obtained therefrom, to a flotation of foam, to produce a flotation concentrate rich in sulfur, and a float tail depleted of sulfur, - and subjecting the solid residue to a sulfur extraction with a suitable extractant, to produce the low sulfur residue. 44. A process according to claim 43, wherein the spent sulfur flotation tail is subjected to a solid / liquid separation, to produce a liquid, which is recirculated to foam flotation, and a solid that is subjected to oxidative leaching. 45. A process according to claim 43 or claim 44, wherein the sulfur extraction is carried out at a temperature of about 90 ° C to 150 ° C. 46. A process according to claim 45, wherein the sulfur extractant is selected from the group consisting of keracene and perchlorethylene. 47. A process according to claim 46, wherein the oxidative leaching is carried out at a temperature of about 200 ° C to 220 ° C, and a partial pressure of oxygen of about 500 to 1200 kPa under acidic conditions. 48. A process for the recovery of Ni / Co values from a concentrate containing Ni / Co hydroxide, which comprises the steps of: subjecting the concentrate to a Ni / Co leaching stage with an ammonium solution , to produce a leaching solution containing Ni / Co values and a waste; and controlling the concentration of nickel in the leaching solution to a maximum value of about 3 to 25 grams / liter. 49. A process according to claim 48, wherein the maximum value is from about 8 to 15 grams / liter. 50. A process according to claim 49, wherein the maximum value is about 10 grams / liter. 51. A process according to any of claims 48 to 50, wherein the leaching step of Ni / Co is carried out with an ammonium sulfate solution. 52. A process according to claim 51, wherein the ammonium sulfate solution has a concentration of about 150 to 250 grams / liter. 53. A process according to claim 52, wherein the ammonium sulfate solution has a concentration of about 200 grams / liter. 54. A process according to any of claims 48 to 50, wherein the Ni / Co leaching stage is carried out with an ammonium carbonate solution. 55. A process according to any of claims 48 to 50, wherein the Ni / Co leaching stage is carried out with a mixture of ammonium sulfate and ammonium carbonate in solution. 56. A process according to one of claims 48 to 55, which further comprises the steps of: subjecting the residue to an acid wash step, to produce a wash solution containing Ni / Co values, and a residue disposable; subjecting the wash solution to a selective precipitation treatment, to obtain a solid containing Ni / Co hydroxide, and recycling the solid to the Ni / Co leaching stage. 57. A process according to any of claims 48 to 55, which further comprises the steps of: subjecting the residue to an acid wash step to produce a wash solution containing Ni / Co values, and a disposable waste; and treating the washing solution for the recovery of the Ni / Co values therefrom. 58. A process according to claim 56 or claim 57, which further comprises the step of subjecting the residue to a washing step before the acid washing step, to produce a second washing solution containing values of Ni / Co, and a residue that is subjected to the acid wash step. 59. A process according to claim 58, which further comprises the step of subjecting one or both of the leaching solution and the additional washing solution containing Ni / Co values, to a solvent extraction, to recover the values of Ni / Co from it. 60. A process according to claim 59, wherein the extraction with solvent is carried out with a nickel extractant, to produce a solution containing nickel. 61. A process according to claim 59, wherein the extraction with solvent is carried out with a cobalt extractant, to produce a solution containing cobalt. 62. A process according to claim 58, wherein the solvent extraction comprises the steps of: effecting a solvent extraction of cobalt in the presence of magnesium ions, using a cobalt extractant, to produce a charged cobalt extractant with cobalt ions, and a first raffinate containing nickel and magnesium ions in solution, - make a solvent extraction of magnesium on the first raffinate, using a magnesium extractant, to produce a magnesium extractant loaded with magnesium ions and of cobalt, and a second raffinate; and extracting with nickel solvent on the second raffinate, using a nickel extractant, to produce an extractant loaded with nickel, and a third raffinate. 63. A process according to claim 62, which further comprises the step of separating extractants charged with cobalt and with nickel, to produce cobalt and nickel solutions, respectively. 64. A process according to claim 63, which further comprises the step of subjecting the cobalt and nickel solutions to an electrolytic extraction, to recover the cobalt and the nickel thereof. 65. A process according to claim 63, which further comprises the step of: separating the magnesium extractant to produce a pregnant solution, which contains magnesium and cobalt ions; and recycling the pregnant solution towards the selective precipitation treatment. 66. A process according to claim 62, wherein the cobalt extractant is similar to the magnesium extractant, the extractant being more selective for cobalt than for magnesium. 67. A process according to claim 59, wherein the solvent extraction comprises the steps of: effecting a solvent extraction of cobalt at a pH of about 6 to 8, to produce a cobalt solution and a first raffinate; and effecting a nickel solvent extraction on the first raffinate at substantially the same pH as extraction with cobalt solvent, to produce a nickel solution and a second raffinate. 68. A process for extracting the Ni / Co values from a solution, which comprises the steps of effecting a solvent extraction of cobalt in the presence of magnesium ions, using a cobalt extractant, to produce a cobalt extractant loaded with cobalt ions, and a first raffinate containing nickel and magnesium ions in solution, - extract with magnesium solvent on the first raffinate, using a magnesium extractant, to produce a charged magnesium extractant with magnesium and cobalt ions, and a second raffinate; and extracting with nickel solvent on the second raffinate, using a nickel extractant, to produce an extractant loaded with nickel and a third raffinate. 69. A process according to claim 68, which further comprises the step of separating extractants charged with cobalt and with nickel, to produce cobalt and nickel solutions, respectively. 70. A process according to claim 69, wherein, before separation, the extractant charged with cobalt is purified with a cobalt concentrated solution, which is recycled from the separation. 71. A process according to claim 70, which further comprises the step of subjecting the cobalt and nickel solutions to an electrolytic extraction, to recover the cobalt and the nickel thereof. 72. A process according to claim 71, which further comprises the step of: separating the magnesium extractant to produce a pregnant solution containing magnesium and cobalt ions. 73. A process according to claim 72, wherein, before separation of the magnesium extractant, it is purified with a concentrated solution of magnesium, which comprises the pregnant solution which is recycled from the separation of the magnesium extractant. 74. A process according to claim 68, wherein the cobalt extractant is similar to the magnesium extractant, the extractant being more selective for cobalt than for magnesium. 75. A process according to claim 68, wherein the solvent extraction comprises the steps of: performing a solvent extraction of cobalt at a pH of about 6 to 8, to produce a cobalt solution and a first raffinate; and effecting a nickel solvent extraction on the first raffinate, substantially at the same pH as the cobalt solvent extraction, to produce a nickel solution and a second raffinate.
MXPA/A/1997/009729A 1993-07-29 1997-12-05 Hydrometalurgical extraction of nickel and cobalt assisted by chloride, from sulf minerals MXPA97009729A (en)

Priority Applications (2)

Application Number Priority Date Filing Date Title
US08/488,128 US5650057A (en) 1993-07-29 1995-06-07 Chloride assisted hydrometallurgical extraction of metal
US08488128 1995-06-07

Publications (2)

Publication Number Publication Date
MX9709729A MX9709729A (en) 1998-10-31
MXPA97009729A true MXPA97009729A (en) 1999-01-11

Family

ID=

Similar Documents

Publication Publication Date Title
CA2356050C (en) Process for the solvent extraction of nickel and cobalt values in the presence of magnesium ions from a solution
US6054105A (en) Process for the solvent extraction of nickel and cobalt values in the presence of magnesium ions from a solution
US5855858A (en) Process for the recovery of nickel and/or cobalt from an ore or concentrate
US5902474A (en) Chloride assisted hydrometallurgical extraction of metal
US6171564B1 (en) Process for extraction of metal from an ore or concentrate containing nickel and/or cobalt
US5874055A (en) Chloride assisted hydrometallurgical extraction of metal
US5869012A (en) Chloride assisted hydrometallurgical extraction of metal
AU2003233283B2 (en) Chloride assisted hydrometallurgical extraction of metals
AU2014270210B2 (en) Method for recovering metals
CA2854778A1 (en) Recovery of zinc and manganese from pyrometalurgy sludge or residues
AU728941B2 (en) Process for the recovery of nickel and/or cobalt from a concentrate
MXPA97009729A (en) Hydrometalurgical extraction of nickel and cobalt assisted by chloride, from sulf minerals
AU731780B2 (en) Chloride assisted hydrometallurgical extraction of metal
AU3878201A (en) Process for the recovery of nickel, and/or cobalt from a concentrate
MXPA97009727A (en) Hydrometalurgical extraction of metal assisted porclor