MXPA97007161A - Mineral lixiviation process at atmospheric pressure - Google Patents

Abstract

The present invention relates to a method of processing a sulfide mineral composition, which at least partially comprises an iron-containing mineral, the method being characterized in that it comprises the steps of: (a) grinding the composition to a particle size P80 of 20 microns or less, (b) leaching the composition with a solution comprising sulfuric and ferric ions at ambient pressure, while being sprayed with an oxygen-containing gas in an open tank reactor at a temperature up to about boiling of the solution, so that at least some of the acid and at least some of the ferric ions are obtained from the dissolution of the iron-containing mineral and the ferrous ions generated by the leaching reaction are substantially re-oxidized to ferric ions in the leaching solution, (c) precipitate the iron and separate the iron and solid materials from the leaching solution n) (d) extract the desired metal ions from the leaching solution by solvent extraction with an organic solvent to form an organic phase and the refining comprising sulfuric acid and ferric ions, e) return the raffinate to the leach tank reactor open and mix with more ground composition, f) separate the metal from the organic phase obtained in step (d) by washing with electrolyte from an electrolytic extraction cell and perform the electrolytic extraction.

Description

MINERAL LEACHAGE PROCESS AT ATMOSPHERIC PRESSURE DESCRIPTION OF THE INVENTION This invention relates to a method that allows a sulfide mineral composition to be leached at atmospheric pressure instead of above atmospheric pressure, which has hitherto been required to achieve stable leaching rates. Sulfur minerals such as copper, nickel, zinc, gold and the like are recovered from their mines by many well-known processes. One such process uses the mineral in solution to allow the ore to be leached from the ore. Conventional leaching processes require expensive equipment and a high level of technical expertise to maintain and use the equipment. In this way, it is not uncommon for the oxidative hydrometallurgy leaching plant to be located at some distance separated from the body of the ore and still in another state. This in turn significantly increases transportation costs and realizes that the transportation of the ore or only the partially enriched ore containing perhaps only a little percent of the desired mineral is extremely wasted and undesirable, but in the absence of being able to recover the value metal of the site where the ore is, there is little real alternative.

Oxidative hydrometallurgy processing methods are commonly used in many different applications. Due to the refractory nature of many of the mineral species treated in such processes, these applications generally require high temperature and pressure leaching conditions and require substantial oxygen supplies. For example, base metals such as copper, nickel and zinc can be recovered by hydrometallurgical processes, which usually include pretreatment, oxidative pressure leaching, solid / liquid separation, solution purification, metal precipitation or solvent extraction and electrolytic extraction. According to conventional technology, oxidative leaching processes usually require very aggressive conditions to achieve acceptable oxidation rates and / or final metal recoveries. Under these conditions, which often mean temperatures in excess of 150 ° C or alternatively temperatures in the range of 150-200 ° C and total pressures in excess of 1500 kPa, chemical reactions which occur using large amounts of oxygen, both stoichiometric considerations and in practice, where quantities in excess of the stoichiometric requirements are used due to inefficient processes.

An example of oxidative hydrometallurgy is the treatment of refractory gold ores or concentrates. The refractory gold ores are those gold ores from which gold can not be easily leached by the practice of conventional cyanidation. The refractory nature of these gold ores is essentially due to very fine encapsulated gold (sub-icroscopic) within the sulfide minerals. This gold can often only be released by chemical destruction (usually oxidation) of the sulfide structure, before gold recovery, which is usually done by dissolving in cyanide solution. Of course, other leachates such as thiourea and halogen compounds and the like can also be used. Many of the processing options are available for the treatment of refractory gold ores, which contain sulfur minerals such as pyrite, arsenopyrite and others. Oxidation under pressure, typified by the so-called Sherritt process, is one such process which typically consists of the steps of feed preparation, pressure oxidation, solid / liquid separation, liquid neutralization and gold recovery from solids. oxidized usually by cyanidation. A cryogenic oxygen plant is usually required to supply the substantial levels of oxygen demand during the oxidation stage with pressure, which is the center of the Sherriff process. Typically, the conditions for the oxidation stage with pressure require temperatures in the range of 150 ° C to 210 ° C, a total pressure of 2100 kPa, a density of the paste equivalent to 20% to 30% solids by mass, a Retention time from two hours to three hours. The typical oxidative hydrometallurgical processing methods mentioned in the foregoing generally have oxidation reactions that are carried out in autoclaves of multiple compartments adapted with agitators. To withstand the generally aggressive conditions of the reactions, autoclaves are very expensive, both to install and maintain. These containers must be capable of withstanding high pressure and acid coatings and bricks resistant to the acid needed to be used. The agitators are made of titanium metal and the pressure release systems used are also expensive and require high maintenance. These high costs and sophistication of technology (operators with skill are generally required) take away from the wider acceptance of high pressure / high temperature oxidation, particularly for use in remote areas or by small to medium operators.

The cooling of the agitators also presents a problem and expensive cooling coils and heat exchange jackets are required to maintain the leaching temperature at optimum conditions. The aggressive leaching conditions described for the recovery of metal values from base metal concentrates are required to achieve acceptable leaching rates of the minerals. Under conditions of atmospheric pressure, the leaching rates of the mineral species are too low to support economically viable leaching processes. Attempts have been made to reduce aggressive conditions and to reduce pressures to reduce the cost of construction and operation of a leaching plant. For example, it is known initially to grind fine ore or ore concentrate (which is known to be used in flotation as an initial step for minerals concentrated in ore), before oxidative hydrometallurgy to leach ore. Fine grinding increases the surface area to volume ratio in the ore particles to improve extraction. Fine milling at 80% size that exceeds 15 microns or less is used. The initial fine grinding results in acceptable leaching rates that are observed with less aggressive conditions and the leaching can be carried out at temperatures of 95-110 ° C and at a pressure of about 10 atmospheres or about 1000 kPa. In this way, although some progress has been made in reducing the operating parameters and thus the cost of the system to leach to date, leaching still has to be carried out under pressurized conditions. Pressurized leach systems are expensive to build. Due to the high capital and processing costs of pressure leaching systems, these systems are only economical for a high degree of concentrates. High grade concentrates are required because - (1) the operating cost per unit of contained metal considerations (2) less heat generation / exchanger problems with high grade concentrations (3) capital costs per unit of metal content decreasing the rolling with great initial capacity that is spent for metal recovery. It is also known oxidatively to leach species of sulfur ore with ferric ions. The ferric ion is a relatively effective oxidizing agent, which allows oxidation to be carried out at a pressure lower than that normally required, when oxygen is the oxidant. However, there are many practical deficiencies associated with the use of ferric ions as the oxidant. First, at an ambient pressure, the reaction is inherently slow. Also during the leaching reaction, the ferric ions are reduced to ferrous ions. The formation of ferrous ions in the leach solution adversely affects the rate of leaching. Also the ferrous ions must be removed normally from the leach liquor before another processing which is difficult. Leaching solutions are generally recycled. However, before a ferrous leach solution can be recycled, the ferrous ions must be re-oxidized to ferric ions. This is because it is important for the maximum effectiveness of leaching that most of the iron in iron form. Leaching solutions can be regenerated by electrolytic oxidation, use of strong oxidants such as permanganate, oxidation under high oxygen pressure, or oxidation by bacteria. Each of these methods experiences disadvantages, which limit its application. For example, high pressure oxidation is limited by the costs of the autoclaves involved. Oxidation by oxygen under ambient pressure may occur, but only at an inherently slow rate. The catalysts can be used to increase but such catalysts are expensive and are not economical for the recovery of low grade ores.

Each of the above processes either requires expensive autoclaves or other equipment and / or the addition of expensive reagents used for the oxidation or regeneration of the ferric ions. This means that it is only economically viable for processes with high grade ores by these methods. Another disadvantage of these processes is that they generate significant amounts of waste products, such as gypsum, sulfuric acid and jarosite. These products must be disposed of in an environmentally acceptable manner which is also added to the cost. Many ores that carry valuable copper or zinc were found in association with iron-containing ores such as pyrite. Pyrite is of little value and is effectively a diluent of valuable ores. In addition, the leaching of pyrite produces the iron species, which interferes with the extraction of the desired metals. Pyrite is therefore usually removed from other ores before processing. Pyrite can be removed by methods such as flotation. Such a separation also adds significantly to the cost and in some cases it is not economically feasible to process some low grade ores at all. It is an object of the present invention to provide a method of processing a mineral composition, which can be carried out under mild conditions of temperature and pressure and which is economical when compared to the existing processes. Surprisingly it has been discovered that by subjecting the sulfide compositions to fine grinding prior to leaching under conditions in which the chemistry of the solution is controlled in a particular form, such compositions must be processed under ambient conditions in open reactors without the need of the addition of expensive reagents and a separate step for the regeneration of the solution of the leaching solution. According to a first embodiment of the present invention, there is provided a method of processing the sulfide mineral composition, which at least in part comprises an iron-containing mineral, the method comprising the steps of; (a) milling the composition to a particle size P80 of 20 microns or less, (b) leaching the composition with a solution comprising sulfuric acid and ferric ions at its ambient pressure, while being sprayed with a gas containing oxygen in a reactor open at a temperature up to above the boiling point of the solution, so that at least some of the acid and at least some of the ferric ions are obtained from the dissolution of the iron-containing ore, and the ferrous ions generated by the leaching reaction substantially reoxidized to ferric ions in the leaching solution; (c) precipitate the iron and separate the iron and solid materials from the leaching solution; (d) extracting the desired metal ions from the leaching solution by solvent extraction with an organic solvent to form an organic phase and the raffinate comprises sulfuric acid and ferric ions; (e) return the raffinate to the open tank reactor and mix with another ground composition; (f) separating the metals from the organic phase obtained in step (d) by washing with electrolyte from an electrolytic extraction cell and performing the electrolytic extraction. According to a second embodiment of the present invention, it provides a method of processing a sulfide mineral composition, which at least partially comprises an iron-containing mineral, the method comprising the steps of: (a) grinding the composition at a particle size of P80 of 20 microns or less and (b) leaching the composition with a solution comprising sulfuric acid and ferric ions at ambient pressure, while being sprayed with an oxygen-containing gas in an open tank reactor at a temperature up to about the boiling point of the solution, so that at least some of the acid and at least some of the ferric ions are obtained from the dissolution of the iron-containing mineral, and the ferrous ions generated by the Leaching reaction are substantially reoxidized to ferric ions in the leach solution. According to a third embodiment of the present invention, there is provided a method of processing a metal sulfide flotation concentrate, comprising the steps of; (a) grind the ore at P80 of 5 microns and (b) leach the ore with a solution comprising sulfuric acid and ferric ions, at ambient pressure while being sprayed with an oxygen-containing gas in a tank reactor open to a temperature up to about the boiling point of the solution. The method of the present invention is applicable to any type of sulfide mineral composition. Such compositions include ores and concentrates. The method of the present invention is especially suitable for the processing of the concentrate. Examples of suitable materials include chalcopyrite, bornite, enargite, pyrite, covelite, sphalerite, chalcocite, pentlandite, cobaltite, pyrrhotite or mixtures of any two or more thereof. The metals, which can be extracted from the mineral compositions according to the method of the first embodiment, include copper, zinc, nickel and cobalt. The degree of the concentrate may be in the range of very low, such as for example with copper-containing materials of 7-8% by weight of copper to high-grade concentrates having approximately 26% by weight of copper. The ore that contains iron can be any mineral, which under the conditions of leaching will produce ferrous ions or ferric ions by dissolution. Especially preferred is pyrite, FeS2 »or pyrite ore, which produces ferric ion and some sulfuric acid according to the following: FeS2 + 2O2 - > FeSO4 + S ° 2FeS2 + 7 / 2O2 - > 2FeS04 + 2H2SO4 4FeS04 + 02 + 2H2SO, - > 2Fe2 (SO,) 3 + H2O FeS2 + 15/402 + l / 2H2O --- > l / 2Fe2 (SO4), + l / 2H2SO4 Preferably, the ore containing sufficient iron is present in such a way as to provide substantially all of the ferric ions in the leaching solution. The relative amounts of a mineral that contains iron will, of course, depend on the types and amounts of the other components in the ore. Typically about 20 to about 60% by weight of pyrite is present. If desired, additional pyrite or other materials can also be added. Alternatively, additional ferric sulfate can be added. It can be seen that materials containing iron can also provide a source of sulfuric acid. It may be necessary to provide additional acid, if required. Sulfuric acid is typically generated in associated processes, such as electrolytic extraction and solvent extraction. Preferably the sulfuric acid produced in this form is recycled to the leaching step. A preferred type of apparatus, which may be suitable for producing fine or ultrafine sulfide mineral compositions in finely divided form is a stirring mill. However, it will be appreciated that other types of tritutation apparatus may also be used, such as wet and dry vibrating mills or planetary mills to provide fine or ultrafine milling of the invention. In a preferred form, vertical or horizontal agitator mills generally consist of a tank filled with small diameter grinding media (eg 1-6 mm diameter steel balls or ceramic), which are agitated by means of a Vertical or horizontal arrow usually adapted with arms or perpendicular discs. The sulfide minerals (usually contained in the form of a concentrate) are milled by the bending action produced by the ball-to-ball contact, or between the balls and the agitator or the balls and walls of the tank. The grinding can be carried out dry or wet. These vertical or horizontal mills with agitation have been found to be satisfactory in providing the required degree of fineness and in satisfying the energy and milling consumption requirements. In addition, the activity of the ground product as measured by its response to subsequent oxidation has also been found to be satisfactory. In this regard, the ore is ground to an average particle size, maximum of 80% which passed the size of 20 microns as measured with a laser-sized meter. In the present specification and claims, the term P80 is used to describe the size at which 80% of the mass of the material will pass. Preferably, the particle size is less than P80 of 5 microns. The desired particle size may vary with the type of mineral species used. Especially, the preferred particle sizes for different concentrates, expressed as P80, are chalcopyrite / bornite - 4.5 microns; enargite - 3 microns; pyrite - 3 microns; covelita - 20 microns; chalcocite - 20 microns; pentlandite - 5 microns and cobaltite -5 microns. The mild pressure and temperature conditions in oxidative leaching, which follows milling, are low when compared to the relatively high pressure and temperature conditions of known pressure oxidation techniques such as the Sherritt process or the Activox process. As indicated in the above, the Sherritt process typically requires temperatures in the order of 150 to 210 ° C and the total pressure in the order of 2100 kPa. The Activox process is designed to operate at a pressure between 9 and 10 atmospheres and temperatures within the range of 90 - 110 ° C. However, the acceleration of the leaching response of the mineral species according to the present invention allows the oxidative leaching to be carried out at temperatures below about 100 ° C and at atmospheric pressure in an open tank reactor. cheap. With the preferred operating conditions that are approximately 60 ° C up to the boiling point of the solution and at a total pressure of 1 atmosphere, a low cost reactor, such as the open tank is sufficient to serve as the leach vessel. There is also no need for the use of expensive titanium metal stirrer, due to the less corrosive nature of the leaching solution. In addition, abrasion problems are substantially reduced due mainly to the fine nature of the feed. Importantly, the heat exchange and pressure of the complex leave the systems necessary for the operation of a pressure vessel, they are not necessary, since the reactor operates at atmospheric pressure, the excess heat is removed from the system by means of the operation -of the solution and this eliminates the need for expensive heat exchange accessories. Also, the reaction becomes autogenous above about 60 to about 70 ° C. If additional heating is required, this can easily be done by known methods, such as steam injection. In addition, suspensions with much higher percent solids can be treated by the method described, due to the relaxation of the requirement to decrease the sulfur levels in the feed to an autoclave necessary for heat control purposes. Typically, the density of the leaching suspension ranges from about 10 to about 20% by weight. In the method of the first embodiment, the leaching solution is refined typically recycled from the solvent extraction stage. In this case, the ferric ion and the sulfuric acid can be more preferably generated by electrolytic leaching / solvent extraction / extraction processes. If leaching is not part of a continuous process, sulfuric and ferric acid may be added if required. Typically, the raffinate comprises 30 -40 g / l of H2SO4 and 10-20 g / l of Fe. The Fe will normally be present as a mixture of ferric and ferrous ions.

The leaching solution is sprayed with a gas containing oxygen, the gas can be air or preferably oxygen or air enriched with oxygen. The gas flow is dependent on the amount of oxygen required to sustain the leaching reaction and the regeneration of the ferric ions. Typically, the gas flow from about 400 to about 1000 kg of 02 per ton of metal produced. If desired, a surfactant or the like can be added to minimize the foaming of the leaching solution. Under the conditions of the leaching reaction, the metals can be oxidized by the ferric ion according to the following general reaction: MS + 2FeJ * > Ma * + 2Fe2 * + s ° Another oxidation of the sulfur element to sulfate according to the reaction: S ° + 3/202 + H20 --- > H2S04 it requires high temperature and pressure and does not occur to any significant extent under the leaching conditions of the present invention. For example, at 90 ° C at atmospheric pressure, in the absence of bacterial catalysis, less than 5% of the elemental sulfur is oxidized to sulfate. By comparison, at 180 ° C and atmospheres and partially pressurized oxygen, most of the sulfur is oxidized to sulfate. Oxidation to sulfate has several disadvantages since additional neutralization reagents are required during the post-leaching neutralization steps. Another additional advantage of the formation of elemental sulfur is that the gaseous emission such as sulfur dioxide is reduced to a minimum, which causes an environmental hazard. In addition, by not carrying out the oxidation completely to sulphate, the oxygen consumption is reduced significantly which saves on the operating costs. For example, the conventional PSA oxygen plant may be sufficient for oxygen supply without the need for cryogenic oxygen plants. This in turn reduces the cost of capital and operating costs by the simple use to operate the equipment. Ferric ions are regenerated by the reaction of ferrous ions with oxygen according to: 4Fe 2 »4H * 0 > 2, > 4Fe J * 2H, O Typically, ferrous ion oxidation occurs at a rate of 2-5 g of oxidized ferrous ion per liter of suspension per reaction hour. After, substantially all of the ore has been oxidized, the leached suspension can be further processed according to known methods. Preferably, the suspension is filtered to remove the solids and the clear liquid subjected to solvent extraction, followed by electrolytic extraction. Typically, the leaching suspension is neutralized before any further processing. As described above, the production of sulfatto is reduced under the conditions of the present invention, thereby minimizing the amount of neutralizing agents, which need to be added. Typically, the suspension is neutralized by the addition of lime or the like. This also precipitates excess iron, arsenic and other impurities generated in leaching. Tests carried out under the conditions of the present invention have also indicated that the iron can be selectively precipitated and remains in the leaching residue as goethite, jarosite or some form of hydrous oxide, while valuable minerals similar to nickel , copper or zinc remain in solution. If desired, the precipitated solids can be further filtered and any remaining liquid can be returned to the leaching solution for further processing. As described above, substantially sulfur-sulfur is oxidized to elemental sulfur during the leaching reaction. Elemental sulfur is present as finely dispersed granules. Because the leaching is carried out at temperatures below the sulfur melting point, ie 120 ° C, the agglomeration of molten sulfur is prevented. The granulated sulfur is normally removed from the leaching solution with the goethite residue and / or other iron.

The solids which are separated from the leaching solution can also be treated - to extract any of the precious metals such as gold, platinum or silver. These extraction methods such as cyanidation for gold are well known in the art. The steps of solvent extraction and electrolytic extraction are well known in the art and need not be described in detail. Typically, the neutralized suspension can be filtered and extracted with an organic solvent, which recovers the metals such as copper, nickel or zinc. The metals can then be washed out of the organic phase by known methods. The metals are then separated from the electrolyte by electrolytic extraction. The spent electrolyte can then be returned to the washing stage. BRIEF DESCRIPTION OF THE DRAWINGS Figure 1 is a flow chart of the method according to a preferred embodiment of the invention. The present invention will now be described in relation to the following examples. However, it will be appreciated that the generality of the invention as described in the foregoing should not be limited by the following description.

Example one - Leaching of enargite and chalcocite The copper flotation concentrate containing 19.5% copper, 4.0% arsenic, 23% iron, 2.35 g / t gold and sulfur sulfur is ground to a size of 80% passes 5 microns in a one-liter, horizontal, agitated ball mill. Mineralogically, the concentrate was composed of 11.9% chalcocite (Cu2S), 20.9% enargite (Cu3AsS4), 50% pyrite (FeS2) and the rest were siliceous gangue minerals. The ground pulp was leached at 90 ° C in an open reactor using the leaching solution of sulfuric-ferric acid with either oxygen or sprayed air. The solids were recovered by filtration and the resuspension was washed with 5% v / v of sulfuric acid solution, before being dried and tested. The post-leaching solutions were analyzed for copper, arsenic and iron by conventional atomic absorption spectroscopic analytical methods. The ferrous and ferric levels were determined by titrations with potassium permanganate, while the acid levels were determined by a neutralization method. More than 92% copper was obtained in the solution from the ground concentrate at a P80 particle size of 3.5 microns in 10 hours, using leaching conditions of 10% density of the paste, 30 g / l of ions ferrous, 50 g / l of sulfuric acid, 90 ° C, oxygen sprinkled and 2.0 kg / ton of lignosol. Lignosol was used to reduce the amount of foaming in the initial stages of leaching. The leaching residue is then subjected to leaching in aerated sodium cyanide solution to recover the gold. The dominant test is thought to be occurring in the previous system, it involves the ferric ions that act as the oxidant, although it is conceivable that an acid / oxygen mechanism is also functioning. The predominant reactions that occur in this leaching system are presented in the following. Cu2S + H2S04 + 5/202? 2Cu04 + H2O Cu2S + 2Fe2 (SO,? 2CuS04 + 4FeS04 + Sß 2CUjAsS4 + llFe2 (S04) 3 + 8H20 - * 6CuS04 + 2H3As04 + 5H2S04 + 8S ° + 22FeS04 FeSj + 202? FeS04 + S ° FeS2 + 15/40, + 1 / 2H, 0? l / 2Fe2 (SO 3 + 1 / 2H2S04 The ferric iron was regenerated in the leaching solution by the action of oxygen on the ferrous iron according to the reaction: 2FeS04 + H5S0 + l / 202 --- > Fe2 (S0). + H20 In this way, the ferric oxidant was continuously regenerated during the leaching process. The density of the pulp in the reactor appeared to be limited by the solubility of the copper sulfate in the resulting iron / acid electrolyte. The use of air instead of oxygen increases the residence time of leaching from 10 hours (oxygen) to 14 hours (air) without loss of total copper recovery. Example Two - Leaching from Chalcocite Concentrate in copper flotation containing 8.1% copper, 0.2% of arsenic, 13.8% of iron and 18% of sulfur sulfur was ground to a size of 80% that passes from 5 microns in a one-liter, horizontal, agitated ball mill. This concentrate had 9.4% of chalcocite (Cu2S) of enargite (Cu3AsS4), 29.6% of pyrite (FeS2) and the rest was a silicia gangue. The ground pulp was leached at 90 ° C in an open reactor using a ferric ion leaching solution - sulfuric acid with either oxygen or sprayed air. Over 95% of the copper solution was achieved in 10 hours using leaching conditions of 10% density of the paste, 30 g / l of ferric ions, 50 g / l of sulfuric acid, 90 ° C, oxygen sprayed and 2.0 kg / ton of lignosol. Example Three - Chalcopyrite leaching Copper flotation concentrate containing 18.0% copper, 25.5% iron and 18.6% sulfur sulfide was milled to a size of 80% to 5 microns in a ball mill in agitation, 1 liter, horizontal. This concentrate had 51.8% chalcopyrite (CuFeS2), 20.8% pyrite (FeS2) and the rest was a silicia gangue. The milled pulp was leached at 80 ° C in an open reactor using a leaching solution of ferric ion ion sulfuric acid with oxygen spray. Over 95% copper of the solution was achieved in 10 hours using the leaching conditions of 10% density of the paste, 5 g / l ferric ions, 20 g / l of ferrous iron, 50 g / l of sulfuric acid , 90 ° C, sprinkled with oxygen and 2.0 kg / ton of lignosol. The method of the invention can be used in association with other processes upstream or downstream. For example, before fine grinding and leaching, the ore can be treated in one or more flotation stages. Downstream processes can include solvent extraction and electrolyte extraction steps. Example Four - Nickel Leaching The nickel-containing concentrate containing 1.7% nickel, 0.03% cobalt, 11% iron and 16% sulfur sulfur was ground to a size of 80% which passes 5 microns in a mill stirring balls, 1 liter, horizontal. Mineralogically the concentrate was composed of pentlandite, pyrite and the rest were silicia gangue materials.

The ground pulp was leached at 90 ° C in an open reactor using a leaching solution of sulfuric acid - ferric ion with oxygen spray. The solids were recovered by filtration and the resuspension was washed with water, before being dried and tested. The post-leaching solutions were analyzed for nickel, cobalt and iron by conventional atomic absorption spectroscopic analytical methods. The ferrous and ferric levels were determined by titrations with potassium permanganate, while the acid levels were determined by neutralization method. More than 92% of nickel and 86% of cobalt were obtained from the solution from the ground concentrate at a P80 particle size of 5 microns in a leaching time of 8 hours using the leaching conditions of 10% density of the paste, 5 g / l ferric ions, 20 g / l of ferrous iron, 80 g / l of sulfuric acid, 90 ° C, oxygen sprayed and 2.0 kg / ton of lignosol. Lignosol was used to reduce the amount of foaming in the initial stages of leaching. Example Five - Cobalt Leaching The cobalt-bearing concentrate containing 0.309% cobalt, 8.5% iron and 0.66% arsenic was milled to a size of 80% that passes 3 microns in a shaking ball mill, of 1 liter, horizontal. Mineralogically, the concentrate was composed of cobalt and pyrite cobaltoferrosa, pyrite and the rest were silicia gangue minerals. The ground pulp is leached at 90 ° C in an open reactor using a leaching solution of sulfuric acid - ferric ion with oxygen spray. The solids were recovered by filtration and the resuspension was washed with water, before drying and testing. The post-leaching solutions were analyzed for cobalt and iron by conventional atomic absorption spectroscopic analytical methods. The ferrous and ferric levels were determined by titrations with potassium permanganate, while the acid levels are determined by a neutralization method. Over 79% of the cobalt solution was obtained from the ground concentrate at a particle size of P80 of 3 microns in a leaching time of 8 hours using leaching conditions of 10% density of the paste, 10 g / l of ferric ions, 50 g / l of sulfuric acid, 90 ° C, sprinkled with oxygen and 2.0 kg / ton of lignosol. Example Six - Zinc Leaching Zinc concentrate containing 46.6% zinc, 10% iron and 2.8% lead is ground to a size of 80% that passes 3 microns in a horizontal 1 liter agitated ball mill . Mineralogically, the concentrate is composed of sphalerite, galena, pyrite and the rest was silicious gangue materials. The ground pasta is leached at 90 ° C in an open reactor using a leaching solution of sulfuric-ferric acid with oxygen spray. The solids were recovered by filtration and the resuspension was washed with water, before being dried and tested. The post-leaching solutions were analyzed for zinc and iron by conventional atomic absorption spectroscopy analytical methods. The ferrous and ferric levels were determined by titrations with potassium permanganate, while the level of the acid was determined by a neutralization method. More than 97% zinc extraction was obtained from the ground zinc concentrate at a particle size of P80 of 3 microns in a leaching time of 8 hours, using leaching conditions of 10% density of the pulp, g / l of ferric ions, 50 g / l of sulfuric acid, 90 ° C, oxygen spray and 2.0 kg / ton of lignosol. Example Seven - Chalcopyrite leaching as a continuous process The following example describes the operation of a fully continuous pilot plant designed to produce 8 kg per day of copper from the A LME cathode from a copper concentrate. The pilot plant is run for 21 days treating the feed described in the following. With reference to Figure 1, a concentrated sample of the composition listed in the following was suspended in tap water at a suspension density of 60% w / w. The suspension is then milled in step 1 to a particle size of 80% which passes 10 microns in a horizontal stirring bead mill. Table 1 Composition of the concentrated sample: CuFeS2 37% p / p FeS2 44% p / p Si02 11% p / p Other 8% p / p Then the sample of the suspension is mixed in step 2 with raffinate 3 of the solvent extraction plant to dilute the density of the suspension of 15% w / w. The raffinate had 35 g / l of H2SO4, 9 g / l of ferric iron and 10 g / l of ferrous iron. The diluted suspension is then pumped through a three-vessel leaching train 4,5,6 at a flow rate designed to give a residence time in the leach train for 20 hours. The leaching train consisting of three tanks with diverting plates of 100 liters, in agitation. The suspension is flowed by gravity from one tank to the next. The tanks are maintained at 90 ° C by a combination of the exothermic nature of the leaching reaction and the injection of active vapor into the suspension. Oxygen was injected into the suspension by air sprinklers 7,8,9 located under the leach agitator. Oxygen was added at a rate of 600 kg per tonne of copper produced. The copper extraction by means of the leaching circuit was 97% w / w. No acid or iron sulfate was added to the leach circuit for the duration of the 21-day run of the pilot plant. The spill of the slurry leached to the tank the final leaching in a neutralization tank 10 and then to a 10A thickener. The composition of the leaching solution was typically 17-19 g / l of copper and 35-45 g / l of iron. The suspension was neutralized to pH 2.0 with the lime slurry 12 to precipitate the iron from the leaching slurry as goethite 13. The finely granulated elemental sulfur was removed with the goethite. The suspension comprising the goethite, sulfur and leach residue was filtered 11 and any liquid was returned to the leaching solution 14. The neutralized suspension is then pumped through a plate and a pressure filter of structure 15. The filtrate had 17-19 g / l of copper, 20 g / l of iron and 5 g / l of H2S04. The filtrate was then pumped through the three-stage solvent extraction plant 16,17,18 to recover the copper from the leach liquor. The refining of the solvent extraction stage had 0.3 g / l of copper, 20 g / l of iron and 35 g / l of H2SO4 and 20 g / l of iron, and was transferred back to the leaching circuit and mixed with more concentrate ground in 2. The organic material charged from the solvent extraction plant is then washed 19 with spent electrolyte containing 180 g / l of H2SO4. The washed organic material 20 is pumped back to the extraction stage. The rich electrolyte 21 is then pumped through a 200 l electrolytic extraction cell 22 containing two cathodes and three anodes. Each surface of the cathode has an area of 0.25 m2. The copper was electrodeposited outside the rich electrolyte at a current density of 280 A / m 2 to produce a cathode plate. The spent electrolyte of the cell had 180 g / l of H2SO4 and 32 g / l of copper. The copper cathode was analyzed and the requirements for the Grade A LME were found. The residue from the goethite leachate was leached in sodium cyanide to determine the amount of gold, which can be recovered from the oxidized concentrate. The leached residue was leached at 45% w / w solids at a pH of 10, with a free cyanide level of 300 ppm maintained throughout the leaching. The suspension was leached for 24 hours in a discontinuous test. The gold recovery of the leaching residue was 92.5% w / w. It may be noted that the method of the present invention offers many advantages over existing methods. The need to adjust control over concentrate grade is relaxed due to lower leaching operation costs relative to pressurized leaching and control of excess heat generation through evaporative cooling of open tank reactors . The leaching can be carried out in cheap open tanks instead of the expensive pressure vessel. A pressurized leach fitting, of equivalent size can cost approximately 6 to 8 times as much as the open tank leaching system of the present invention. This also allows a leaching circuit to be built practically on the site. This avoids transportation costs, which can be considerable. In some cases these costs can make it inexpensive for transportation and the process of low grade ores. The leaching circuit is less sensitive to the metal grade. Therefore, this will allow greater recovery of the metal in the operations of the unit upwards.

The method of the present invention is also capable of producing a high-grade electrolytic metal. For example, in some cases it may be possible to produce a product by means of solvent extraction / electrolytic extraction, which can be sold directly. The molten metal generally requires other refining. This allows substantial cost savings as well as production of a product, which attracts a premium price. The reaction of li is self-sustaining, when coupled with a solvent extraction - electrolytic extraction plant, as the only reagent that needs to be added to the leach is air / oxygen and a neutralizing agent such as lime. The need for the addition of expensive reagents was eliminated. The present invention is ideal for mineralogically complex ores, which are finely disseminated from base metal sulfide minerals with another sulfide (e.g., chalcocite, sphalerite, enargite that covers pyrite ores) that have traditionally been difficult to treat metallurgically. It should be appreciated that various other changes and modifications can be made to the described modalities, without departing from the spirit and scope of the invention.

Claims (29)

1. CLAIMS 1. A method of processing a sulfide mineral composition, which at least partially comprises a ferrous-containing mineral, the method is characterized in that it comprises the steps of; (a) grinding the composition to a particle size P80 of 20 microns or less, (b) leaching the composition with a solution comprising sulfuric acid and ferric ions at ambient pressure, while being sprayed with a gas containing oxygen in a tank reactor open at a temperature up to approximately the boiling point of the solution, so that at least some of the acid and at least some of the ferric ions are obtained from the dissolution of the iron-containing mineral and the ferrous ions generated by the leach reaction are substantially reoxidized to ferric ions in the leach solution; (c) precipitate the iron and separate the iron and solid materials from the leaching solution; (d) extracting the desired metal ions from the leaching solution by solvent extraction with an organic solvent to form an organic phase and raffinate comprising sulfuric acid and ferric ions; (e) return the raffinate to the open leach tank reactor and mix with more ground composition; (f) separating the metal from the organic phase obtained in step (d) by washing with electrolyte from an electrolytic extraction cell and performing the electrolytic extraction.
2. 2. The method of compliance with the claim 1, characterized in that the composition is a flotation concentrate.
3. 3. The method of compliance with the claim 2, characterized in that the concentrate is a low grade concentrate.
4. 4. The method according to claim 1, characterized in that the ore containing iron is a pyrite ore.
5. 5. The method according to claim 4, characterized in that the mineral composition comprises 20 to 60% by weight of pyrite.
6. 6. The method according to claim 5, characterized in that substantially all ferric ions are generated by dissolving the pyrite.
7. 7. The method of compliance with the claim 1, characterized in that the composition is milled to a particle size of P80 of 10 microns or less.
8. The method according to claim 1, characterized in that the temperature is from about 60 ° C to about the boiling point of the leaching solution.
9. 9. The method according to claim 1, characterized in that the gas is oxygen.
10. 10. The method of compliance with the claim 9, characterized in that the oxygen is sprayed at an index of 400 to 1000 kg per tonne of metal produced.
11. 11. The method according to the claim I, characterized in that the solid of step (c) is further leached to recover any of the precious metals.
12. 12. The method in accordance with the claim II, characterized in that the precious metals are selected from the group consisting of gold, platinum or silver.
13. The method according to claim 1 or claim 2, characterized in that the metal is selected from the group consisting of copper, zinc, nickel or cobalt.
14. 14. A method of processing a sulfide mineral composition comprising from about 30 to about 40% by weight of chalcopyrite, from about 40 to about 50% by weight of pyrite and up to 20% by weight of silicon gangue, the method is characterized in that it comprises: (a) grinding the composition to a particle size P80 of 10 microns, (b) leaching the composition with a solution comprising sulfuric acid and ferric ions at ambient pressure, while being sprayed with oxygen at a rate of about 600 kg per ton of copper produced in an open reactor at a temperature of about 90 °, (c) neutralize the leach solution with lime to precipitate the excess iron as goethite and remove the goethite and any other solids from the solution leaching; (d) filtering the leaching solution and extracting the dissolved copper from the leaching solution by solvent extraction, with an organic solvent in such a manner that the raffinate comprises sulfuric acid and ferric ions; (e) return the raffinate to the leach tank and mix with another ground composition; (f) separating the copper from the organic phase obtained in step (c) by washing with electrolyte from an electrolytic extraction cell and performing the electrolytic extraction.
15. 15. A method of processing a sulfide mineral composition which at least partially comprises a ferrous-containing mineral, the method is characterized in that it comprises the steps of: (a) grinding the composition to a particle size P80 of 20 microns or less, and (b) leaching the composition with a solution, which comprises sulfuric acid and ferric ions at ambient pressure, while being sprayed with an oxygen-containing gas in an open tank reactor at a temperature up to about boiling of the solution, so that at least some of the acid and at least some of the ferric ions are obtained from the dissolution of the pyrite and the ferrous ions generated by the leaching reaction are substantially reoxidized to ferric ions in the leaching solution.
16. 16. The method according to claim 15, characterized in that the composition is a flotation concentrate.
17. 17. The method according to claim 15, characterized in that the iron-containing mineral is a pyrite ore.
18. 18. The method of compliance with the claim17, characterized in that the composition comprises from 20 to 60% by weight of pyrite.
19. 19. The method according to the claim 18, characterized in that substantially all ferric ions are generated by the dissolution of pyrite.
20. 20. The method according to claim 15, characterized in that the composition is milled at a particle size P80 of 5 microns.
21. 21. The method according to claim 15, characterized in that the temperature is about 60 ° C up to the boiling point of the leaching solution.
22. 22. The method of compliance with the claim 15, characterized in that the gas is oxygenated.
23. 23. The method according to claim 22, characterized in that the oxygen is sprayed at a rate of about 400 to about 1000 kg per ton of metal produced.
24. 24. The method according to claim 15, characterized in that a surfactant is added in step (a) to minimize the foaming of the solution.
25. 25. A method of processing a sulfur metal flotation concentrate characterized in that it comprises the steps of; (a) grind the ore at P80 of 5 microns and (b) leach the concentrate with a solution comprising sulfuric acid and ferric ions, at ambient pressure while being sprayed with an oxygen-containing gas in an open tank reactor to a temperature up to about the boiling point of the solution.
26. 26. The method according to claim 25, characterized in that the temperature is about 60 ° C up to the boiling point of the leaching solution.
27. 27. The method of compliance with the claim 26, characterized in that the gas is oxygen.
28. 28. The method of compliance with the claim 27, characterized in that oxygen is sprayed at an index of about 400 to about 1000 kg per tonne of metal produced.
29. 29. The method according to claim 28, characterized in that a surfactant is added in step (a) to minimize the foaming of the solution. SUMMARY OF THE INVENTION The present invention relates to a method of processing a sulfide mineral composition, which at least partially comprises an iron-containing mineral, the method comprising the step of: (a) grinding the composition to a size of particle P80 of 20 microns or less; (b) leaching the composition with a solution comprising sulfuric acid and ferric ions at ambient pressure, while being sprayed with an oxygen-containing gas in an open tank reactor at a temperature up to about the boiling point of the solution, whereby at least some of the acid and at least some of the ferric ions are obtained from the dissolution of the iron-containing mineral and the ferrous ions generated by the leaching reaction are substantially re-oxidized to ferric ions in the leaching solution; (c) precipitate the excess iron and separate the iron together with any of the solid materials from the leaching solution; (d) extracting the desired metal ions from the leaching solution by solvent extraction with an organic solvent, such that the raffinate comprises sulfuric acid and ferric ions; (e) return the raffinate to the leach tank and mix with another ground composition; (f) separating the metals from the organic phase obtained in step (c) by washing with electrolyte and electrolytic extraction.