Patent Application
20140301919
The present invention relates to the leaching and recovery of alumina from bauxite and other aluminum-bearing ores and concentrates, using a mixed chloride lixiviant. In particular, the invention relates to a process for the leaching of aluminum-bearing ores and concentrates with a lixiviant comprising hydrochloric acid and magnesium chloride. In embodiments, the invention relates to a process for the leaching and recovery of alumina from an aluminum-bearing ore or concentrate with a lixiviant comprising hydrochloric acid and magnesium chloride, followed by solvent extraction to remove iron from the leachate liquid, and then by steps to recover aluminum in the form of alumina. Lixiviant, extractant and other materials used in the process may be recovered and recycled. As exemplified, very high levels of extraction and recovery of alumina element may be achieved.
Bauxite is an aluminum ore that is widely used in the production of metallic aluminum. Laterite or silicate bauxites and karst or carbonate bauxites are known. The aluminum may be primarily in the form of hydrated aluminum oxides e.g. gibbsite (Al(OH)3), and various forms of AlO(OH), examples of which are boehmite and diaspore. The aluminum may also be in the form of aluminosilicates e.g. argillite and nepheline. All of these minerals are typically in mixtures with iron oxides e.g. goethite and hematite.
Bauxite is usually processed into alumina and then into aluminum, especially using a process referred to as the Bayer process, in which bauxite is heated in a pressure vessel with a sodium hydroxide solution at 150-200° C. to form an aluminate. After separation of residue, gibbsite is precipitated and converted into alumina. The alumina may be converted into aluminum by electrolysis.
Processes for the extraction of aluminum and iron from aluminous ores are known. U.S. Pat. Nos. 7,837,961 of R. Boudreault et al, issued 23 Nov. 2010, and 8,241,594 of R. Boudreault et al, issued 14 Aug. 2012, describe processes for extracting aluminum and iron from aluminous ores. Extraction is carried out using hydrochloric acid. The resultant solution is extracted with an organometallic complex in hydrocarbon solvent, the complex being substantially selective with respect to ions of iron. The hydrocarbon solvent may be a C5-C12 alkane. Alternatively, the leachate may be reacted with alkali at a pH of at least 10 to precipitate iron.
In other mining industries, processes have been developed to extract value metals e.g. nickel and titanium, with recycling of solutions used in the process and with substantially reduced environmental issues. One example is the use of solutions of hydrochloric acid and magnesium chloride in the extraction of titanium from titanium-bearing ores or concentrates at low temperatures, as disclosed in U.S. Pat. No. 7,803,336 of V. I. Lakshmanan, R. Sridhar et al, issued 28 Sep. 2010. U.S. Pat. No. 7,736,606 of V. I. Lakshmanan, R. Sridhar et al, issued 15 Jun. 2010, describes extraction of base metal sulphide ores with a lixiviant of a chloride, an oxidant and hydrochloric acid. The leaching may be conducted to limit the co-dissolution of platinum group metals and gold with the base metals.
Processes that are effective for the recovery of alumina from aluminous ores in a form suitable for use in cosmetic and other industries would be beneficial. Such processes should have low or minimal environmental issues. For instance, the processes should include potential for recycling of components used in the process so as to reduce environmental issues and for economic operation of the process.
In an aspect, the present invention provides a process for the leaching and recovery of aluminum in the form of alumina from aluminum-bearing ore or concentrate in which solutions used in various stages of the process may be recycled. In particular, the process utilizes chloride chemistry in the leaching step. In a further aspect, the present invention provides a process for the leaching and recovery of alumina using chloride chemistry in the leaching step and in which the leaching step is followed by steps to remove iron prior to recovery of alumina.
Accordingly, the present invention provides a process for the extraction of alumina from an aluminum-bearing ore or concentrate, comprising the steps of
In a preferred embodiment of the process, prior to step a), the aluminum-bearing ore or concentrate has been subjected to a calcination step.
In another preferred embodiment, liquid from step b) is subjected to steps for recovery of alumina.
In another preferred embodiment, the temperature is in the range of 80-90° C.
Another aspect of the present invention provides a process for the extraction of an aluminum-bearing ore or concentrate, comprising the steps of
In a preferred embodiment, prior to step a), the aluminum-bearing ore or concentrate has been subjected to a calcination step.
In another preferred embodiment, in step e), the pregnant aluminum-bearing organic solution is subjected to stripping followed by recovery of alumina.
In a further embodiment, alumina is recovered by precipitation or crystallization.
In a still further embodiment, the temperature in step a) is in the range of 80-90° C.
The present invention relates to a process for leaching of aluminum-bearing ores or concentrates. The ore or concentrate may be in the form of bauxite for instance a laterite or silicate bauxite, or karst or carbonate bauxite. The aluminum may be primarily in the form of hydrated aluminum oxides e.g. gibbsite (Al(OH)3), and various forms of AlO(OH), examples of which are boehmite and diaspore. The aluminum may also be in the form of aluminosilicates e.g. argillite and nepheline syenite. All of these minerals are typically in mixtures with iron oxides e.g. goethite and hematite. In addition, the aluminum bearing ore or concentrate could be in the form of a hydrated aluminum oxide that has been obtained from, for example, a Bayer process.
The ore or concentrate is preferably subjected to a calcining step and then leached with a lixiviant of hydrochloric acid and magnesium chloride under conditions such that aluminum is leached into solution. After a solids/liquids separation step, steps may be taken to remove iron present in leachate solution. Subsequently, steps may be taken to recover alumina. The hydrochloric acid and magnesium chloride of the lixiviant may be recovered and recycled to the leaching step.
The present invention particularly relates to a process for leaching of aluminum-bearing ores or concentrates. Such ores typically contain iron. The process is operated at atmospheric pressure. While the concentration of hydrochloric acid may be varied, the preferred concentration of hydrochloric acid is not more than about 20% (mass ratio). Such a concentration of acid may be obtained by azeotropic distillation of hydrochloric acid solution, for example in recycle of hydrochloric acid solution in the process e.g. using pyrohydrolysis. As discussed herein, such a low concentration of hydrochloric acid provides advantages to the process, including in recycle steps in the process and in requirements for disposal of effluent.
The aluminum-bearing ore material may be ore per se, but is preferably a concentrate thereof. Sources of ore are discussed herein. Techniques for treating ore to form a concentrate or for beneficiation of the ore are known. The ore may also be in the form of tailings from another process. Flotation processes may be used to obtain concentrates for the process of the present invention. It is understood that the expression “ore” also includes any other form of the ore, and that mixtures of the various forms of the ore may be used. In particular embodiments of the invention, the process is operated on tailings, waste or other material obtained from another process. The process may also be operated on refined or other products that are relatively high in aluminum content. Other sources will be apparent to persons skilled in the art. The material fed to the process should be comminuted or ground to improve the contact between lixiviant and aluminum-bearing ore or concentrate. All of these forms are regarded as ore or concentrate to be fed to the process.
The process is preferably operated with a calcining step prior to leaching of the ore or concentrate. For instance, the ore or concentrate may be calcined at a temperature of 500-800° C. Processes for the calcination of ores or concentrates are known in the art.
In the method of the present invention, ore in a form as discussed above and especially ore that has been calcined is fed to a leaching step in which the ore is contacted and leached with a lixiviant comprising magnesium chloride and hydrochloric acid. Optionally, an oxidant may be added. Regeneration and recycling of ingredients used in the leach and in subsequent extraction and recovery steps is an important aspect of the process; use of other chlorides or mixtures of chlorides e.g. alkali metal chlorides, may adversely affect such recovery and recycling. Thus, the lixiviant is restricted to hydrochloric acid and with magnesium chloride as the only added metal chloride. While oxidant is optional, if used oxidant may be selected from the group consisting of alkali metal peroxide, alkali metal perchlorate, ammonium perchlorate, magnesium perchlorate, alkali metal chlorate, magnesium chlorate, alkali metal hypochlorite, chlorine, hydrogen peroxide and other non-sulphur containing oxidants, and mixtures thereof. Examples of alkali metal peroxide are sodium peroxide and potassium peroxide. Examples of alkali metal perchlorates are sodium perchlorate and potassium perchlorate. Ammonium perchlorate, magnesium perchlorate and magnesium chlorate may also be used. Examples of alkali metal chlorates are sodium chlorate and potassium chlorate. An example of an alkali metal hypochlorite is sodium hypochlorite. Other oxidants are non-sulphur containing oxidants; the presence of sulphur in oxidants is to be avoided. The preferred oxidants are chlorine and sodium chlorate.
The aluminum-bearing ore or concentrate is subjected to a leach with a lixiviant of hydrochloric acid and magnesium chloride. The leaching may be conducted as a co-current step, a countercurrent step or in another manner, and is most conveniently carried out at atmospheric (ambient) pressure. It is not necessary to conduct the leaching step under pressure. According to the invention, the leach is carried out at a temperature in the range of from 80° C. up to the boiling point of the leaching solution, and especially at a temperature in the range of 80-90° C. In addition, the oxidation potential (Eh) should be at least 200 mV and especially at least 400 and more particularly in the range of 400-900 mV. The leaching step is preferably carried out with a magnesium chloride concentration of at least 150 g/L, preferably 150-400 mg/L and especially in the range of 200-300 g/L. The hydrochloric acid used in the lixiviant has a maximum concentration of 18% (mass ratio). In embodiments, the concentration of hydrochloric acid is at least 12% and especially in the range of 14-18% (mass ratio); for instance the hydrochloric acid may be 3-6N. The chloride concentration in the lixiviant is most preferably in the range of 100-400 WL, and especially 200-300 g/L. The Mg/HCl (magnesium to hydrochloric acid) ratio expressed in terms of mass percentage (mlm) in the leach is preferably adjusted to optimize the leach, based on for example the particular ore or concentrate being leached and temperature. The Mg/HCl ratio in the lixiviant is generally in the range of 0.1-2.0, especially 0.4-1.0. The leach may be a single stage leach or a multi-stage leach. In addition, the leaching steps may be conducted co-current or countercurrent, in whole or in part.
Subsequent to the leach, the leaching solution is subjected to a liquid/solids separation step. In particularly preferred embodiments, the leachate obtained is treated for recovery or removal of iron. For instance, the leachate may be subjected to solvent extraction with an alkyl ketone, especially a dialkyl ketone. In particular, the alkyl groups of the dialkyl ketone may be C1-C6 alkyl groups, especially C1-C4 alkyl groups, examples of which are methyl, ethyl, propyl and butyl including isopropyl and isobutyl. The ketone may also be tridecanone. Preferred ketones are methylisobutylketone (MIBK) and diisobutylketone (DIBK). The alkyl ketone e.g. dialkyl ketone is preferably mixed with a diluent and a modifier. Examples of the diluent are organic phase diluents, for instance kerosenes e.g. CF-231 kerosene. Examples of the modifier are alcohol surfactants e.g. EXXAL™13 tridecylalcohol. In an embodiment, diisobutylketone is used in a ratio of 20/20/60 of DIBK/EXXAL™13 tridecyl alcohol/CF-2321 kerosene. Iron is stripped from the loaded organic solution obtained e.g. by addition of water or dilute HCl solution, and may be recovered by pyrohydrolysis to give Fe2O3 which may be offered for sale. The raffinate from iron pyrohydrolysis contains hydrochloric acid and magnesium chloride which may be recycled to the leaching step. Methods for purification and recycling of such liquids are known. Before stripping of iron, loaded organic may be subjected to stripping with iron containing hydrochloric acid. Remaining solution may be subjected to pyrohydrolysis to obtain hydrochloric acid.
The pregnant aqueous liquid obtained as a raffinate from the step for removal of iron is then subjected to steps to extract aluminum. Optionally, the pregnant aqueous liquid may be subjected to an acid solvent extraction step prior to steps to solvent extraction of aluminum.
In the optional step after the extraction of iron from the leach solution, and prior to solvent extraction to recover alumina, free acid may be removed from the raffinate obtained after iron removal. Addition of an alkali leads to the presence of other cations in the raffinate. Therefore, it is preferred that free acid be removed from the raffinate by including solvent extraction step, especially by using an alcohol, for example EXXAL™10 branched chain alcohol to increase the pH to greater than 1.5 prior to recovery of alumina. The raffinate may also be treated for removal and recovery of other elements e.g. niobium, uranium, thallium, copper and zinc, prior to recovery of alumina.
The pregnant aluminum-bearing solution may be subjected to organic solvent extraction to recover aluminum. In particular, the solvent used in this extraction is specific for extraction of aluminum from the solution e.g. using Cyanex™923 in a mixture of TBP (tri-n-butyl phosphate as a modifier) and a diluent; diluents have been described above. The organic solution may be scrubbed with a solution of AlCl3 in dilute hydrochloric acid e.g. 0.2N HCl and then stripped with e.g. 1-2N HCl. The aluminum-bearing product obtained is then subjected to steps to recover alumina. For instance, alumina may be precipitated by addition of a base e.g. sodium hydroxide to raise the pH of the solution and effect precipitation of alumina, followed by calcination. As exemplified, aluminum may be precipitated as aluminum hydroxide (Al(OH)3) at a pH of 4. Alternatively, aluminum may be crystallized from solution in the form of aluminum chloride (AlCl3), which may then be calcined to obtain alumina.
The invention has been described herein with respect to the leaching of ore containing aluminum using a lixiviant of hydrochloric acid and magnesium chloride, followed by solvent extraction for removal and recovery of impurities, especially iron. Alumina may be recovered from the resultant pregnant solution, especially using solvent extraction. The solvents used for such specific solvent extraction include phosphonic acids, and esters and oxides thereof, crown ethers and derivatives thereof, as well as mixtures of phosphonic acids/esters and oxides with crown ethers and their derivatives. Crown ethers are known and have the general formula (CH2CH2O—)n, where the value of n may be varied, giving the crown ether different properties with respect to solvent extraction of elements.
A particular advantage of the process of the present invention, and its embodiments, is that high rates of extraction of aluminum are obtained in a leaching step that operates at atmospheric pressure. The use of atmospheric pressure results in substantial economic advantages, especially in capital costs. The use of chloride chemistry offers advantages in operating and capital costs of the process. Leaching agent may be regenerated and recycled, especially using a pyrohydrolysis step with additional hydrochloric acid being formed from chlorine if required. Magnesium chloride may be recycled to the leaching step. Iron is separated from the liquor after the leaching step and alumina of high purity may be recovered.
The present invention is illustrated by the following examples.
A sample of a bauxite ore was analyzed for mineral content by an analytical laboratory. The results of the analysis, expressed as percentage of the ore, are shown in Table 1. The data is for ore “as received”, which contained some moisture. LOI is Loss on Ignition.
Samples of calcined ore were leached with a lixiviant of hydrochloric acid (2N) and magnesium chloride (100-250 g/L) at a temperature in the range of 80-90° C. for 5 hours. The leaching was conducted at a pulp density of 10% by weight of solids. The leach liquor was subjected to solid/liquid separation; analysis of the residue and pregnant leach solution showed a recovery of 94-98% aluminum and 78-81% iron from the samples that were leached.
Pregnant leach solution was subjected to solvent extraction for iron using a solution of ketone (DIBK) in solvent and modifier, as described herein; the concentration of ketone was 20-30% by volume. The organic solution obtained was stripped with dilute HCl solution. The raffinate from the iron extraction was subjected to solvent extraction for aluminum using Cyanex™923 (a mixture of four trialkyl phosphine oxides). The loaded organic obtained was first scrubbed using aluminum chloride solution and then stripped using 1-2N HCl. The pregnant strip liquor thus obtained had the following analysis: Aluminum 20680 ppm, iron 0.04 ppm, titanium 0.08 ppm, calcium 6 ppm, magnesium 5.7 ppm, chlorine 83.87 g/L and free acid 0.4N.
A sample of the strip liquor was subjected to an evaporation step and the product obtained was calcined. The calcined alumina product contained 11.34% aluminum, <0.03% titanium, <0.03% calcium and <0.03% magnesium.
The pH of a second sample of the strip liquor was increased to 4 by addition of sodium hydroxide. The precipitate obtained was separated from the liquid and then calcined. Analysis of the calcined alumina product obtained showed 0.003% iron, 0.002% titanium, 0.001% calcium and 0.004% magnesium. Thus, an alumina of very high purity was obtained in a process of the invention.
Based on the composition of the pregnant liquor solution, the precipitated product was calculated to have a composition of 99.95% alumina (Al2O3), 0.0001% Fe2O3, 0.003% TiO2, 0.02% CaO and 0.02% MgO, assuming no preferential precipitation from the pregnant strip solution.
| Number | Date | Country | |
|---|---|---|---|
| 61782109 | Mar 2013 | US |
