RECOVERY OF ZINC FROM SULFUR CONCENTRATESField of the Invention The present invention relates to a novel class of additives useful in zinc extraction processes wherein the process involves one or more pressure leaching steps carried out at temperatures above the melting point of the sulfur.
Background of the Invention It is well known that zinc can be recovered from sulfur concentrates containing zinc by reacting the sulfides with sulfuric acid in the presence of oxygen at elevated temperatures and pressures. At temperatures above the sulfur melting point, the elemental sulfur that forms in the leaching reactions is present as finely divided liquid sulfur globules. As the reaction progresses, the quantity of these globules increases and they coat or plug the unreacted sulphurous particles, rendering them inert towards further oxidation. Additives can be added to the processes and which can prevent, or at least significantly inhibit, the fact that the molten sulfur covers the sulfur particles without leaching, thus allowing the reactions to be carried out until high extractions of water are obtained. zinc, typically above 95% and preferably above 98%, as demonstrated for example in the US patents 3,867,268 and 4,004,991. In addition, the additives contribute to the formation of a finely divided leaching residue, with good physical handling characteristics. Without additives, the extraction of zinc is typically only about 50-70%, and the liquid elemental sulfur can agglomerate, leading to the production of coarse, difficult to handle particles, which plug the pipes and vessels.
The additives used in this process must be compatible with the oxidation of the zinc sulphide, and must not introduce impurities into the zinc-containing process streams. A large number of surfactants have been used for this purpose, including organic compounds such as lignin derivatives, particularly calcium and sodium lignosulfonates; tannin compounds, particularly extracts of tree bark or tree heart wood, such as quebracho extract, American spruce and redwood extracts; ortho-phenylene diamine; and alkaryl sulfonates, particularly sodium alkylbenzene sulfonates. Sodium lignosulfonate and quebracho have been applied in commercial practice. The current status of the state of the art related to the use of surfactants soluble in zinc pressure leaching is summarized in a recent document entitled "Effect of Surfactants on Zinc and Iron Dissolution during Oxidative Leaching of Sphalerite" by G. Owusu et al. ., Hydrometallurgy 38 (1995) 315-324. Most of these state-of-the-art additives are soluble in the zinc sulfate acid solution, and rapidly decompose under the highly oxidizing and high-temperature conditions of the zinc leaching process. Consequently, it can sometimes be difficult to ensure that enough surfactant will always be available in the pressure vessel to avoid occlusion of sulfur particles without leaching by molten sulfur. In addition, the additives are usually added to the composition of the water before introduction to the autoclaves as a solution having a tendency to foam and overflow, thereby causing a potential loss of additive. In extreme cases, if the solution containing the additive fails to come into contact with the zinc sulphide / sulfuric acid suspension, the absence of surfactant can result in the agglomeration of large quantities of molten sulfur in a heavy and separate liquid phase, which causes plugging of the pipes and reaction vessels. In addition, the surfactants of the state of the art are relatively more expensive, and the in case of extracts of tree bark and tree heart wood are of limited supply. It will be appreciated therefore that there is a need for functional additives to prevent the occlusion of sulfur particles by elemental sulfur, which is not expensive, is easily obtainable, and which do not decompose so rapidly under the conditions of the zinc leaching process under pressure.
Objectives of the Invention It is an object of the present invention to provide a class of functional additives in zinc extraction processes, which involves one or more pressure leaching steps and carried out at temperatures above the melting point of sulfur, in where said additives are stable throughout the process, they are relatively cheap, they do not affect the process and finally the quality of the cathode zinc produced in continuation to the purification of the solution and the electrolysis, and which are easily separable in the form physical activity from the leaching residue. Initially, a process that has a single stage of pressure leaching is provided to recover zinc from sulphide concentrates containing zinc. The process comprises dispersing a predetermined amount of finely divided sulfide material containing zinc and iron in an aqueous solution of sulfuric acid to form a suspension. The concentration of added sulfuric acid is effective to provide predetermined concentrations of zinc and iron in the final leach solution. An effective amount of finely particulate charcoal is provided, compatible with the acid sulfate leach solution, which is functional under the reaction conditions of the next oxidation reaction step (ie, pressure leaching), to inhibit that the molten sulfur occludes the sulfur particles without leaching. The suspension and the carbon are then reacted, with stirring, with gas containing free oxygen in an autoclave at a temperature above the melting point of the sulfur to thereby effect the substantially complete extraction of the zinc values from said sulfides as soluble zinc sulfate and the concurrent conversion of sulfur sulfur associated with said zinc values to the elemental form. The product leaching solution containing the dissolved zinc values is then separated from the solid residue. The invention also extends to a pressure, countercurrent and two-step leaching process for recovering zinc from sulfur-containing materials containing zinc, and which comprises dispersing finely divided and iron-containing zinc sulphides in a Aqueous solution of sulfuric acid to form a suspension and provide an effective amount of carbon in particles, finely divided, compatible with the acid sulphate leaching solution, which is functional under the reaction conditions of the first stage of leaching, to inhibit that the molten sulfur occludes the sulfur particles without leaching. The reactants are leached under a positive partial pressure of oxygen and at a temperature above the melting point of the sulfur in two stages in countercurrent, with said sulfur containing materials containing zinc being fed into the first of said two stages of leaching and the electrolyte. spent from the subsequent electrolysis being divided between the first and the second stage of the leaching stages. The amount of said sulfur containing zinc material, which is fed into said first leaching stage is adjusted and correlated in relation to the amount of acid fed to both leaching stages, with said spent electrolyte being such that at least about one mole of zinc in the form of sulfur, it enters said leaching stage for each mole of acid entering both leaching stages and, at the same time, the amount of spent electrolyte entering both leaching stages is controlled in such a way that approximately one mole of acid enters for each mole of zinc that must be dissolved in said electrolyte to increase zinc concentration thereof to a predetermined level, suitable for the treatment of said solution in said electrolysis operation. The first stage of leaching is operated to extract zinc values from said sulfur containing material with concurrent conversion of sulfur sulfur to elemental sulfur to produce a residue from the first stage of leaching., containing unreacted sulphides, elemental sulfur and precipitated iron, such as hydrated iron oxides and jarosites, and a solution of the first leach stage having a pH above 1 and containing above about 160 g / L of dissolved zinc , less than about 10 g / L of free sulfuric acid and less than about 2 g / L of dissolved iron. The solution of the first stage of leaching is passed to iron removal and purification stages of the solution, and then to said electrolysis step. The residue from the first leach stage is passed to said second leach stage, the second leach stage being conducted to extract substantially all of the zinc from said waste from the first leach stage and to produce a residue from the second stage. leaching stage containing unreacted sulphides, elemental sulfur and precipitated iron, and a solution of the second leaching stage containing said extracted zinc, unreacted sulfuric acid and dissolved iron. The solution of the second leaching stage is then separated from said waste from the second leaching stage and said solution from the second leaching stage is passed to said first leaching stage. The waste from the second stage of leaching is passed to a waste treatment process to recover values of unreacted metal and elemental sulfur by-product. Finally, a concurrent pressure leaching process is provided to recover zinc, lead and silver from sulphide materials containing zinc, which also contain economically attractive amounts of lead and silver. The process comprises dispersing finely divided zinc-containing sulfur material with high lead and / or silver content and finely divided zinc-containing sulfur material with low lead and silver content in aqueous solutions of sulfuric acid to form sulfuric acid. separate suspensions, and treat these suspensions in stages of pressure leaching and high and low acidity, respectively. The reactants are leached under a positive partial pressure of oxygen and at a temperature above the melting point of the sulfur. The leaching solution separated from the high acidity leaching stage is added with an effective amount of finely divided carbon, which is functional under the reaction conditions of the low acidity leaching stage, to inhibit the melting sulfur clogging the sulfuric particles without leaching, towards said suspension of the leaching of low acidity. The amount of sulphurous material containing zinc that is fed to said low acidity leaching stage is adjusted and correlated in relation to the amount of residual acid and acid equivalent as iron sulphates, in the high acidity leaching liquor that is fed to said low acidity leaching stage such that at least about 1.3 moles of zinc in the form of sulfide enters said low acidity leaching stage for each mole of acid, and acid equivalent as iron sulfates, entering said stage of leaching of low acidity. Low acid leaching is operated to extract zinc values from said sulfur containing zinc material, with concurrent conversion of sulfur sulfur to elemental sulfur and produce a low acid leaching residue which contains unreacted sulphides, elemental sulfur and precipitated iron, and a low acidity leaching solution having a PH of above 1 and containing above about 160 g / L of dissolved zinc, less than about 10 g / L of free sulfuric acid and less than about 2 g / L of dissolved iron. Said low acidity leaching solution is passed through iron removal stages, purification of the solution and electrolysis. Said low acidity leaching residue undergoes a foam flotation process to separate a floating concentrate containing the sulfides without leaching, elemental sulfur and residual coal from flotation waste containing the precipitated iron and gangue minerals. Said flotation concentrate of the low acid leaching is passed together with the sulfurous material with zinc and containing lead and / or silver, towards said high acidity leaching step, where it comes into contact with spent electrolyte from said stage of leaching. electrolysis, said high acid leaching being carried out to extract substantially all of the zinc from said flotation concentrate from the leaching of low acidity and said sulphurous material with zinc and containing lead and / or silver, there being a stoichiometric excess of sulfuric acid in relation to the zinc content of the combined materials that are fed to said high acidity leaching stage of from about 50 to 100%, to produce an undissolved residue containing a greater proportion of lead and silver and a leaching solution which contains a greater part of zinc and iron. A high grade lead / silver product is separated from said high acidity leaching residue by flotation of foam, with said high grade lead / silver product being recovered as the flotation waste. The functional additives, under the conditions of the pressure leaching stage or steps, to inhibit the melting sulfur from clogging the sulfur particles without leaching, comprise a suitable carbon. Such coals would include hard coal, peat coal, peat, lignite coal, sub-bituminous coal, bituminous coal, semi-bituminous coal, semi-anthracite coal and anthracite coal. The preferred carbons are coals of low classification as opposed to coals of high classification. The preferred particle size of the carbon will vary up to about 60 microns. The preferred amount of the present carbon would be between 5 kg. and 25 kg. of carbon per ton of sulphurous material containing zinc. The total carbon content of the carbon would preferably vary between 40 to 85%. Advantageously, the coal can be recovered from the process waste and subsequently recycled for re-use. The carbon is preferably added to the suspension of zinc concentrate / sulfuric acid directly in the autoclave or can be added at any point during the feed preparation process including washing, grinding or pre-leaching.
BRIEF DESCRIPTION OF THE DRAWINGS The present invention will be better understood with reference to the following detailed description and the following figures. Figure 1 is a flowchart of a zinc recovery process of the invention, having a single step of pressure leaching.
Figure 2 is a flow diagram of a zinc recovery process of the invention, characterized in having a pressure-leaching process, of two stages to counter-current. Figure 3 is a flowchart illustrating a zinc recovery process of the invention, having a two stage co-current pressure leaching process.
Detailed Description of the Preferred Modality Referring to the accompanying figures, the processes of the invention will now be described. The invention is particularly applicable to flowcharts incorporating pressure leaching in a single stage, two-stage leaching in co- and countercurrent, and most of the flowcharts developed for the treatment of zinc sulphide concentrates that They can also contain lead and silver. In general, the invention is applicable to zinc extraction processes which are carried out at temperatures above the melting point of the sulfur. The processes are applicable either high or low grade concentrates (ie, size) containing zinc and iron sulphides. However, the starting material will usually be a sulphide concentrate containing zinc, obtained by selective flotation of sulfur-containing ore containing zinc. Said concentrates will frequently contain other non-ferrous metals in addition to zinc. Bulk sulfur concentrates that are exemplary may include significant levels of copper and lead, and lower levels of nickel, cobalt, cadmium and in many cases a small, but economically important, amount of silver and precious metals. It will be understood that the terms "concentrates or sulfur-containing zinc materials" as used herein are intended to include such materials and are also intended to include any other high or low grade material containing iron either as an aggregate constituent or as occurs naturally and economically recoverable amounts of zinc in the form of sulfur. In order to obtain optimum results, the starting sulphurous material should preferably be in a finely divided particle form. The particle size has a pronounced effect on the reaction rate and the degree of zinc extraction obtained in the pressure leaching stage (s). It is preferable, in order to achieve the full benefits of the invention, that the starting material be approximately 90% less than 44 microns (325 Tyler mesh). Material such as flotation concentrates may, in some cases, already be within the preferred range of size. The material that is not yet within the preferred size range is preferably first sprayed by, for example, wet milling. Optionally, the pulverized concentrate can be leached in dilute sulfuric acid under conditions of atmospheric pressure and low temperature, to remove easily soluble impurities, such as magnesium, manganese, chloride and fluoride. If the sulfides are deficient in iron or do not have iron, the iron content can conveniently be adjusted at this point by adding oxidizable iron to the suspension. All of the processes of the invention include the provision of an additive which is compatible with the aqueous acid sulfate solution and which is functional, under the reaction conditions of the pressure leaching stage (s), for inhibit that the molten sulfur clogs the sulfur particles without leaching. Preferably the additive would be a carbon with low carbon content, which has been found not to introduce undesirable impurities that remain in the leach liquor to interfere subsequently with the electrolysis operation. Such carbons may include hard coal, peat coal, peat, lignite coal, sub-bituminous coal, bituminous coal, semi-bituminous coal, semi-anthracite coal and anthracite coal. The preferred carbons are coals of low classification, or intermediate carbon level rather than coals of high classification or high carbon level. Preferably carbon content of the carbon would vary between 40 to 85%. Generally, carbons with high carbon content, exemplifying what anthracite carbons are, have been found to be less effective than carbons that have a low carbon content. The carbons must be finely divided and have particle sizes that vary up to approximately 60 microns. As a general rule, it is desirable to add the minimum amount of carbon that is effective to maximize the extraction of zinc in any given case. In most cases, the addition of from about 5 to 25 kg will suffice. of coal per ton of zinc-containing material, typically about 10 kg / ton. It is obvious to one skilled in the art that the amount of carbon required will be correlated to the nature thereof, that is, its classification and particle size and additionally taking into account any carbon contained in the material containing zinc. The inherent advantage of using a carbon, as opposed to the additives of the prior art, lies in the stability of the coal and in its ability to function effectively throughout the entire process without chemical decomposition taking place. In addition, unreacted carbon can be physically separated from sulfur, unreacted sulphides and iron residue during the waste treatment process and consequently can, if desired, be recovered and reused. In processes involving two stages of pressure leaching in which the solid waste from one stage is subsequently leached in another stage, the coal needs only to be added to one stage, thus representing a considerable saving over the prior art processes that use surfactants that have to be added in both stages of leaching. As background, ASTM standards use fixed caloric and carbon values calculated on a mineral-free base, as a classification criterion. The classification by class being according to the degree of matamorphism, or progressive alteration, in the natural series from lignite or anthracite. The European classifications include (a) The International Classification of hard Coals by Type and (b) The International Classification of Brown Coals, as taught in "Van Nostrand's Scientific Encyclopedia", 7th edition, Considine Volume 1, pp662-663. Generically, coal can be defined as a combustible, carbonaceous sedimentary stone formed by the compaction of partially decomposed plant materials. Certain more highly carbonized forms of carbon, examples of which are graphite, activated carbon or coke, have been found not suitable for the processes detailed here. Therefore, it is incumbent on one skilled in the art to experimentally determine compounds of the most highly charred carbon forms that are inoperative. Preferably, carbon having an aliphatic carbon content between 20 and 80%, more preferably between 25 and 55% is suitable. The leaching reaction for single leaching stages and first and second stages (in processes involving two stages of leaching) is carried out in pressure vessels equipped with agitation, such as autoclaves, at a temperature above the melting point of sulfur, that is, above about 120 ° C, but below about 175 ° C, preferably between about 135 ° C and 155 ° C. At temperatures below about 135 ° C the reaction rates are relatively slow, adversely affecting in this way the economy of the process. During the oxidation reaction, some elemental sulfur is converted to sulfuric acid and the degree of this reaction increases above about 160 ° C. The generation of some acid is desirable in most cases to compensate for mechanical losses of acid and to replace the acid consumed in irreversible secondary reactions by metal diluents such as lead and gangue material such as calcium and magnesium. At temperatures above about 175 ° C the elemental sulfur is converted to sulfuric acid at a relatively rapid rate and, therefore, more sulfuric acid than desired can be produced to maintain an acid balance. Preferably, therefore, the operating temperature should be from about 135 ° C to about 155 ° C. The leaching reactions are exothermic and produce sufficient heat to maintain the suspension within the preferred temperature range without the addition of heat from an external source once the reactions have been initiated in a continuous leach system. The total pressure at which the leaching reactions are carried out is the vapor pressure autogenously generated at the temperature of the oxidation reaction plus the overpressure of the oxidation gas. Preferably, the oxidation gas is oxygen but air or air enriched with oxygen can also be used. The reaction proceeds satisfactorily with an oxygen overpressure above about 150 kPa. However, there is an improvement in the reaction rate as the oxygen over-pressure is increased. Thus, it is preferred to use an oxygen overpressure of between about 400 to 700 kPa. The upper limit of the oxygen pressure will be that imposed by the autoclave used. As it is generally desirable, for economic reasons, to avoid the use of high pressure autoclaves, generally the upper limit will be about 1000 kPa of oxygen overpressure or air overpressure. The density of the pulp of the leaching suspension fed to the first or only leaching stage, that is, the relative amounts of sulfides and solution provided in the leaching stage in any given case, is determined in relation to the zinc content of the leach pulp. sulfides and the desired zinc or iron concentration of the final leach solution in general. The pressure leaching residue contains non-leached sulphides, elemental sulfur, iron precipitates such as hydrous iron oxides and jarosites, oxidic minerals containing lead and silver such as anglesite, gangue minerals and unreacted carbon. Sulfides without leaching and elemental sulfur are typically agglomerated into fine microspheres, which can be separated from the precipitates, oxidic and gangue minerals by flotation by foam, with the sulfides without leaching and the elemental sulfur being incorporated into the flotation concentrate. The unreacted carbon is also incorporated into the flotation concentrate but as discrete fine particles from unreacted sulfur / sulfur microspheres, and separable from unreacted sulfur / sulfur microspheres by cyclization. The unreacted charcoal fraction that is recovered can be "recycled to the leach stage (s) to decrease the requirement for the addition of fresh coal.The values of lead and silver in the leach residue are concentrated in the flotation waste, and the degree of this material depends on the amount of iron retained in the pressure lixiviation liquor, as described above, returning now specifically to the process for the extraction of zinc that involves only a leaching to pressure and in one stage, as shown in Figure 1, the amount of acid provided in the stage The shape of the suspension is determined by the target concentration of the zinc and iron that is desired in the final leaching solution. There must be enough acid available to combine as zinc sulfate with zinc values contained in the sulfides to produce the desired target concentration of dissolved zinc. It is generally desirable to produce a leaching solution containing about 140 to 180 grams per liter (g / L) of zinc since in most cases the zinc will be recovered from the solution by electrolysis and a zinc concentration within the this interval for the electrolysis stage. Typically, 10 kg. of coal containing between 40 to 85% of total coal are required per ton of material containing zinc for the reasons described above. Once the desired target concentration of zinc and iron is determined, the amount of acid required to produce this concentration can easily be calculated in relation to the stoichiometric requirements of the zinc to be extracted as zinc sulfate. In most cases, the solution that makes up the suspension will be the spent electrolyte that is obtained from the zinc electrolysis stage. The liquor normally contains in addition to the residual zinc values, regenerated sulfuric acid equivalent to the amount of dissolved zinc metal that was recovered in elemental form. Consequently, except for the acid of the initial composition, the total acid requirements for the leaching stage when carried out in conjunction with the electrolytic recovery of zinc, are limited to that amount required to compensate for mechanical losses and for that consumed by acid-reactive diluent materials that are in sulfides, such as lead and calcium, which form insoluble sulfates. The molar ratio of acid / zinc in the leach suspension should be adjusted such that there is at least one mole of acid per mole of zinc. Preferably there should be an excess of acid in the leach slurry over the amount required to combine stoichiometrically with all the zinc values in the slurry, to produce zinc sulfate. However, there is no need for a very large excess of acid since the reaction rate is not greatly increased by such excess and the amount of dissolved iron and free acid in the leach liquor will be undesirably high. Generally the molar ratio of HaSO Zn should be controlled so that it falls in the range of about 1.0: 1 to about 1.2: 1, preferably 1.1: 1.
In summary, therefore, a zinc extraction process having a single leaching stage, as exemplified in U.S. Pat. 3,867,268, the disclosure of which is incorporated herein by reference, comprises leaching the zinc concentrate with sulfuric acid and a carbon additive under the conditions described in general form supra. In commercial practice, this single-stage pressure leaching process is usually operated in conjunction with a conventional toasting-leaching-electrolytic extraction process in which most of the material fed from the zinc concentrate is roasted to form a Calcined zinc oxide, which is leached in return electrolyte in an atmospheric leaching circuit. A part of the zinc oxide calcination is advantageously used to neutralize the sulfuric acid content of the liquor from the pressure leaching. In continuation to the pressure leaching stage the leaching solution containing the zinc values extracted as soluble zinc sulfate is passed to the iron removal circuit where it is treated with the calcinated zinc oxide or lime to effect the neutralization and precipitation of iron. From there the solution is passed through conventional stages of solution purification to the electrolysis step for the recovery of zinc. The sulfur associated with the treated material that is fed in the direct pressure leaching operation is converted to its elemental form in leaching and can be separated from the iron oxide and the jarosite residue by flotation. Alternatively, the single stage pressure leaching system can be operated where the leaching stage is a high acid leaching, namely by the provision of a stoichiometric excess of sulfuric acid which is functional to ensure that the iron remains in solution to allow in this way the subsequent recovery of potential values of lead and silver contained and coming from leaching residues. Again, coal is used as the additive as taught here. With reference to Figure 2 and to the patent of the U.S.A. 4,004,991, the description of which is incorporated herein by reference, describes a zinc extraction process involving a two-stage countercurrent leaching process. This process is the preferred option when all zinc sulphide concentrate is to be treated by pressure leaching, rather than by conventional roasting to zinc oxide. The two-stage countercurrent leaching configuration produces a charged zinc sulfate solution, which contains lower concentrations of acid and iron, than what can be obtained from the single-stage process, thus reducing costs of the neutralization reagent. In the practice of the invention, sulfides, after spraying and optional acid pre-leaching to remove acid-soluble impurities, are made slurry with the recycled leach liquor from the second leach stage and a portion of the electrolyte spent and is passed to the first stage of leaching. The feed suspension is prepared using a particulate zinc concentrate, carbon additive (in an amount ranging from 5 to 25 kg / t of zinc-containing material, and preferably a low-grade carbon) and if necessary an additive of iron as described above. In general, it has now been found that it is necessary to add carbon to the second stage of leaching. The adjustment and correlation of the amount of sulfides fed to the first leach stage and the amount of free acid fed to the first and second leach stages ensures that in the first leaching stage there is an acid deficiency below that required for combined with all of the zinc and other acid-reactive constituents in the sulfides, and in the second leaching there is an excess of acid on that required to be combined with all of the acid reactive constituents in the residue from the first leaching stage. The result is that in the first stage of leaching the pH rises, for example, 2-3 higher, as the reaction progresses and this, in turn, promotes the rapid hydrolysis and precipitation of dissolved iron from the leach liquor discharged of the first stage of leaching. Meanwhile, only a part of the zinc, that is, approximately 75% will be extracted from the sulfides in the first stage of leaching, the recycled liquor from the second stage of leaching will have a high content of dissolved zinc, so that the liqueur the first stage will contain at least about 160 g / L of zinc. In the second stage of leaching, because there is an excess of acid on that required to combine with the reactive constituents to the acid in the residue of the first stage of leaching that is fed to it, strong acidic conditions prevail that will favor the rapid extraction and substantially complete of the zinc values of the residue. The retention times required to produce 98% or better zinc extraction are typically 40 to 60 minutes in the first stage and 40 to 120 minutes in the second stage of leaching. When the second stage of leaching is completed, the final leach suspension is discharged from the leach vessel into a tank and then it is quickly sent (flashed) into a container at atmospheric pressure. The leaching solution is separated from the undissolved residue in a conventional liquid-solid separation stage and is passed in its entirety to the first leaching stage. The residue from the second leaching stage contains all the elemental sulfur produced in both leaching stages, as well as a small amount of sulfides without leaching, hydrated iron oxide, jarosite, insoluble gangue materials and any precious metal present in the starting material. This waste can be disposed of in the trash, stored for further treatment or treated immediately for recovery of elemental sulfur and other economically recoverable values such as lead and / or silver.
The leach liquor from the first leach stage, after separation of the undissolved residue in the liquid-solid separation stage, is passed to the iron removal operation. Following the removal of the iron, the solution is passed to a conventional purification stage and, after the removal of the impurities precipitated in a liquid-solid separation stage, is passed to the conventional zinc electrolysis stage. A more advanced zinc recovery flow diagram, which uses the two-stage countercurrent pressure leaching configuration, in which carbon can be used to replace the first surfactants, is described in US Pat. 5,380,354, in which improved recoveries of lead, silver and iron are obtained from bulk zinc-lead-silver concentrates. The descriptions of this patent are also incorporated herein by reference. This patent describes in Figure 5, taken in conjunction with column 6, lines 10-12 and lines 40-42 of the description, the division of the spent electrolyte by feeding a portion of the spent electrolyte solution of electrolysis. to the leaching stage 10 and a portion of the spent electrolyte solution to the second leaching stage 14. In accordance with the present invention, the use of carbon as the additive to replace the surfactants of the prior art can be applied to the pressure and co-current leaching processes, for the extraction of zinc from bulk concentrates with high lead and silver contents, in conjunction with the treatment of a zinc concentrate with low contents of lead and silver, described in the US patent 4,505,744, the descriptions of which are incorporated herein by reference. Reference is made to Figure 3 for correlation with the following description. A sulphurous material containing zinc and also containing iron, lead and silver is leached under oxidation conditions in a pressure leaching and high acidity essentially as described above. In a low acidity leaching stage, a sulphurous material containing zinc and having a low content of lead and silver is treated. The finely divided carbon additive as described above, preferably a low graded coal, added in an amount ranging from 5 to 25 kg./t of metal containing zinc, and more preferably in an amount of about 10 kg / t. . Typically, it may only be necessary to add carbon to the low acid leach, since the leach residue from this leach stage is typically treated by flotation to remove the elemental sulfur and the sulfides without leaching, including residual carbon, from the oxides of iron and jarocitas, for recycling to the stage of leaching of high acidity. The pressure leaching conditions for pressure leaching and high acidity involve oxidation conditions at a temperature in the range of about 120 to 175 ° C in aqueous solution of sulfuric acid (with spent electrolyte) with significant initial stoichiometric excess of sulfuric acid in relation to the zinc content of the material or approximately 50 to 100% excess sulfuric acid. The oxygen partial pressure is preferably in the range of 400 to 1000 kPa, and the leaching temperature is preferably in the range of 135 to 155 ° C. The leaching residue produced as a result is relatively free of iron. In addition, the elemental sulfur produced in the leaching stages and consequently also present in the waste is easily separable in physical form from the remaining residue containing lead and silver. The high acidity leaching stage is continued for a period of time until more than 97% of the zinc and more than 50% of the iron have been dissolved., leaving up to about 45% of iron present as unreacted pyrite. The undissolved residue then contains little iron oxide or jarocite and contains a major part of the lead and silver that was found in the original sulfur containing zinc material, which was added to the high acidity leaching stage. In the leaching stage of low acidity, the zinc concentrate with low lead / silver content is leached with an excess thereof, the concentration of sulfuric acid being such as to effect an extraction of 65 to 75% of the zinc content. The pressure leaching is carried out under a partial pressure of oxygen of about 400 to 1000 kPa at a temperature of from about 135 to 155 ° C to obtain the zinc extraction in the zinc concentrate with low lead / silver content. concurrent precipitation of iron and neutralization of acid from the high acidity leaching solution fed to the low acidity leaching stage. The leaching solution from the leaching of low acidity and containing from about 140 to 160 g / L of zinc, from about 0.5 to about 2 g / L of iron and from about 1 to 10 g / L of sulfuric acid, it is subjected to a purification step with removal of iron and to any other necessary purification steps and subsequently to a zinc electrolysis step. Therefore, zinc is efficiently recovered from both zinc concentrates and facilitates the recovery of lead and silver from the first zinc concentrate with relatively high lead / silver content. In the following non-limiting examples, embodiments of the invention are provided for extracting zinc from sulphide concentrates containing zinc and iron.
EXAMPLESExample 1 A comparison of the various types of carbon used in zinc extraction experiments is provided in Table I below. The carbon size distribution was 100% passing 63 microns and the zinc concentrate used was Type A, the analysis of which is provided in theTable III.
TABLE Test Type of C added% of distribution% of Extraction of Coal Zinc No. Coal kg./t% Aliphatic Aromatic 20 min. 40 min 60 min.1 MC 50 84 16.6 83.4 62.0 77.5 77.82 NC 50 83 22.6 77.4 67.7 75.0 76.93 OC 50 84 23.0 77.0 65.4 77.2 80.84 PC 50 72 38.1 61.9, 92.8 98.4 98.9QC 50 56 37.2 62.8 94.3 98.4 98.96 RC 50 70 43.2 56.8 90.9 97.5 98.27 SC 50 68 44.8 55.2 93.0 98.6 99.08 TC 25 59 50.5 49.6 95.0 99.1 98.7Where "C%" represents the percentage of organic carbon present in the coal and "aggregate" refers to the amount of carbon added in relation to the amount of zinc concentrate. The "carbon distribution" refers to the relative amounts of aliphatic and aromatic carbon in the carbon as determined by NMR spectroscopy.
Example II This example is illustrative of the comparison of the amount of coal used with the zinc extraction obtained. The concentrate used was Type A (see Table III below) and Type TC coal. The distribution of carbon size was 100%, passing 44 microns. TABLE II Test Addition% Zinc Extraction No. kg / t 20 min. 40 min 60 min. 1 20 92.4 98.2 98.6 2 10 90.4 98.3 98.9 3 5 87.9 96.3 98.2 4 2 74.1 71.0 74.8 5 0 - - 60.8 Where "addition" means the amount of coal added in relation to the amount of the zinc concentrate.
EXAMPLE III The effect of the type of the zinc concentrate in the extraction of zinc using TC carbon additive is given in Table III. Typically, the zinc extraction was greater than 96%, being more of the order of 98%. This was obtained for concentrates containing 0.1 to 1.8% Cu, 2.2 to 11.8% Fe, 0.1 to 13.6% Pb, 0.4 to 3.3% Si, 26.2 to 33.6% S and 45.5. to 56.2% of Zn.
TABLE IIIAnalysis of the concentrate (%) Coal Extraction of Zn (%)ZnC "LS.Cu Fe Pb Si S Zn kg / t 60 min 120 min.
A 0.20 8.00 2.58 1.23 32.2 52.4 25 98.7 B 0.09 3.98 3.30 1.42 30.9 56.2 20 97.2 99.7C 1.63 6.02 1.69 0.51 30.5 55.8 25 98.8 98.9D 0.81 7.18 0.07 0.38 33.2 56.1 25 98.9 99.4E 1.10 10.7 0.38 0.77 31.6 52.5 25 98.0 99.0F 0.93 11.8 1.04 0.52 32.7 49.5 25 96.3 97.4G 0.50 6.28 0.63 1.23 32.6 55.2 25 98.3 99.3H 0.25 9.31 0.54 0.39 33.6 53.4 25 96.5 98.4I 1.83 2.20 13.6 3.32 26.2 45.5 25 '91.6 96.2J 2 0.39 7.05 5.81 1.35 31.0 51.0 20 - 98.7 *K 2 0.79 4.86 12.9 2.11 28.8 45.5 25 - 96.3 **Where ZnC "denotes the type of zinc concentrate, LS is the number of leaching stages," * "represents a two-stage leaching consisting of a first leach stage of 40 minutes and a second leach stage of 120 minutes, and "**" is a two stage continuous leaching test, wherein the first leach stage has a 50 minute retention and a second 100 minute leach stage.
EXAMPLE IV This example demonstrates the improved zinc extraction obtained in a batch leach and in a single stage, with decreasing size of carbon particle. As shown in Table IV, the upper limit of the carbon particle size for high zinc extraction in 60 minutes is approximately 63 microns.
TABLE IVTest Distribution of the size of the Zinc Extraction% Carbon No. 20 min. 40 min 60 min.1 - . 1 -350, +63 μm 59.7 63.4 78.3 2 -150, +44 μm 60.4 78.4 84.7 3 -63 μm 92.2 98.4 98.9 4 -44 μm 92.4 98.2 98.6Where the type of coal is TC and the zinc concentrate is Type A.
Example V This example demonstrates that the key interaction for the dispersion of the molten sulfur residues in the contact between the solid particles of carbon with the molten sulfur, more than the interaction of the organic extract of the coal with the molten sulfur. In test 1, the coal is leached by itself, the carbon leachate being then brought into contact with the zinc concentrate. The result, which is identical to not using additives, provides 61% zinc extraction. In tests 2 and 4, relatively thick carbon is held behind a wire mesh inside the autoclave, so that the solution can pass through and extract the organic species, but there is no contact of the carbon with the suspension. This examines the possibility that a short-lived organic extract is responsible for the dispersion of sulfur. Sulfur extraction was limited to approximately 63%. In tests 3 and 5 the direct contact between relatively coarse coal and the suspension of the zinc concentrate are examined and showed that zinc extraction is increased to 78% and 85% respectively. In test number 6, a smaller sample of finer carbon was used, so that the amount of organic species in solution was approximately identical to that of other tests, but it was noted that zinc extraction was increased to 99%.
TABLE V% extraction Coal distribution test Location of total C No. coal size kg / t Coal mg / l 60 min.1 - . 1 -63 μm 50 Solution * 118 60.9 2 -355, + 63 μm 60 Basket ** 30 63.3 3 -355, + 63 μm 20 Suspension *** 44 78.3 4 -150, + 44 μm 40 Basket ** 46 63.1 5 -150, + 44 μm 20 Suspension *** 44 84.7 6 -44 μm 10 Suspension *** 28 98.9Where "*" represents a test in which the coal was leached in simulated zinc pressure leaching liquor for 60 minutes under typical conditions of zinc pressure leaching (150 ° C), the carbon being filtered, adding acid to the solution to simulate the spent electrolyte, and the solution being charged to the autoclave with the zinc concentrate. "**" denotes a test in which the carbon was placed in the autoclave with the zinc concentrate, but separated from the suspension by a wire mesh. "***" represents the coal in the autoclave with concentrate using normal leaching procedures. By "total C" is meant the total organic carbon in solution, the carbon concentration being that of the sample that is being taken after 60 minutes, as determined by chemical analysis of the solution.
Example VI The zinc concentrate type J as shown in Table III was treated in continuous pilot plant autoclaves with the addition of 25 kg / t of TC type carbon with a particle size of 100% passing the 63 microns, in leaching under pressure, in two stages against the current and at 150 ° C. The coal was added to the first stage only. The overall extraction of zinc was 97.7% compared to 97.3% using calcium lignosulfonate as the additive. The K-type concentrate was treated in continuous pilot plant autoclaves with the addition of 25 kg / t of TC-type coal with a particle size of 100% passing the 63 microns, in leaching in two stages to counter-current. The coal was added to the first stage only. The overall zinc extraction was 96.3 compared to 96.0% obtained using calcium lignosulfonate.
GLOSSARY OF TERMS AND LEGENDS APPLICABLE TO THE DRAWINGSFIGURE 1 10 Zinc Concentrate 11 Coal 12 Oxygen 13 Sulfuric Acid 14 Worn Electrolyte 15 Towards Waste Treatment 16 Calcined or Cal 17 Iron Residue 18 Zinc Cathode 100 Pressure Leaching 200 Iron Removal 300 Purification 400 Electrolysis L Liquid Phase S Solid PhaseFIGURE 2 10 Zinc Concentrate 11 Coal 12 Oxygen 14 Worn Electrolyte 15 Towards Waste Treatment 16 Cal 17 Iron Residue 18 Zinc Cathode 110 First Leaching Stage 120 Second Leaching Stage 200 Iron Removal300 Purification 400 Electrolysis L Liquid phase S Solid phaseFIGURE 3 10 Zinc concentrate11 Carbon 12 Oxygen 14 Electrolyte Worn16 Cal 18 Cathode of Zinc 19 Concentrate in BulkElemental Sulfur / Pyrite21 Iron Oxide / Plaster22 Lead / Silver Product23 Iron Oxide 150 Leaching High Acidity160 Leaching of Low Acidity200 Removal of the Iron300 Purification 400 Electrolysis 500 Flotation L Liquid phase S Solid phase